Santana Pea (Sedar Copy)

Amazon GeoServices Ltda ABN 52 065 481 209 Preliminary Economic Assessment - Santana Phosphate Project, Para State, Brazil – MBAC Page: 1 Effective Date – 22 September 2011

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Preliminary Economic Assessment - Santana Phosphate Project, Para State,

Brazil

On Behalf of - MBAC Fertilizer Corp

Effective Date – 22 September 2011

Qualified Persons: Beau Nicholls Consulting Geologist (Amazon Geoservices Ltda) BSc (Geo) MAIG

Carlos Guzman Consulting Mining Engineer (NCL Brasil Ltda) Mining Engineer, Registered Member of the Chilean Mining Commission

Version / Status: FINAL

Path & File Name: D:\P Drive\Projects\MBAC\Reports\43101 Sanatana Res est July 2011\PEA_MBAC_43-101 Santana_22Sept2011_FINAL.docx

Amazon GeoServices Ltda ABN 52 065 481 209 Preliminary Economic Assessment - Santana Phosphate Project, Para State, Brazil – MBAC Page: ii Effective Date – 22 September 2011

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Table of Contents

1 Summary ............................................................................................................................... 11

1.1 Introduction................................................................................................................. 11

1.2 Location...................................................................................................................... 11 1.3 Ownership .................................................................................................................. 11

1.4 Geology and Mineralization ......................................................................................... 11

1.5 Mineral Processing and Metallurgical Testing .............................................................. 11

1.6 Mineral Resources ...................................................................................................... 12

1.7 Mining Operations....................................................................................................... 13

1.8 Indicative Economics .................................................................................................. 13 1.9 Conclusions and Recommendations ........................................................................... 16

2 In troduction ........................................................................................................................... 17

2.1 Scope of Work ............................................................................................................ 17

2.2 Forward Looking Information....................................................................................... 17

2.3 Principal Sources of Information.................................................................................. 18 2.4 Qualifications and Experience ..................................................................................... 18

2.5 Units of Measurements and Currency ......................................................................... 18

2.6 Abbreviations .............................................................................................................. 19

3 Reliance on Other Experts .................................................................................................... 20

4 Property Des crip tion and Location ...................................................................................... 21

4.1 Project Location .......................................................................................................... 21

4.2 Tenement Status ........................................................................................................ 22

4.3 Royalties and Agreements .......................................................................................... 23 4.4 Environmental Liabilities ............................................................................................. 23

4.5 Permitting ................................................................................................................... 23

5 Acc es s ib ility, Climate , Local Res ources , In fras tructu re and Phys iography ...................... 24

5.1 Project Access ............................................................................................................ 24

5.2 Physiography and Climate .......................................................................................... 24 5.3 Local Infrastructure and Services ................................................................................ 24

6 His to ry ................................................................................................................................... 25

7 Geological Setting and Mineraliza tion ................................................................................. 25

7.1 Regional Geology ....................................................................................................... 25

7.2 Local and Property Geology ........................................................................................ 26

7.3 Mineralization ............................................................................................................. 28

8 Depos it Types ....................................................................................................................... 29

9 Explo ration ............................................................................................................................ 30

9.1 Ground Magnetic Survey ............................................................................................ 30

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9.2 Ground Penetrating Radar .......................................................................................... 30 9.3 Auger Drilling .............................................................................................................. 31

9.4 Bulk Density Determinations ....................................................................................... 31

10 Drilling ................................................................................................................................... 32

10.1 RC Drilling .................................................................................................................. 32 10.1.1 RC Drilling Results and Quality ............................................................................. 32

10.2 DC Drilling .................................................................................................................. 34 10.2.1 DC Drilling Results and Quality ............................................................................. 34

10.3 Drilling Results............................................................................................................ 34

11 Sample Preparation , Analys es and Security ....................................................................... 35

11.1 MBAC Sampling Method ............................................................................................. 35 11.1.1 Diamond Core ....................................................................................................... 35 11.1.2 Reverse Circulation ............................................................................................... 35

11.2 Sample Security ......................................................................................................... 36 11.3 ALS Laboratory Sample Preparation and Analysis ...................................................... 36

11.4 Adequacy of Procedures ............................................................................................. 36

12 Data Verifica tion .................................................................................................................... 37

12.1 Geological Database .................................................................................................. 37

12.2 QAQC ........................................................................................................................ 37 12.2.1 Certified Standards and Blanks ............................................................................. 37 12.2.2 Field Duplicates ..................................................................................................... 40 12.2.3 Data Quality Summary .......................................................................................... 42

13 Mineral Proces s ing and Metallu rg ical Tes ting..................................................................... 43

13.1 Samples Preparation and Methods ............................................................................. 43 13.1.1 Outcrop Samples ................................................................................................... 43 13.1.2 Drill Core Samples ................................................................................................ 44

13.2 Results and Discussion ............................................................................................... 47 13.2.1 Outcrop Samples ................................................................................................... 47 13.2.2 Drill Core Samples ................................................................................................ 50

13.3 Concentrate Characteristics ........................................................................................ 54

13.4 Recovery Evaluation ................................................................................................... 54

13.5 Solubility Tests ........................................................................................................... 55

13.6 Conclusion.................................................................................................................. 56

14 Mineral Res ource Es timates ................................................................................................. 57

14.1 Introduction................................................................................................................. 57

14.2 Geological Modeling ................................................................................................... 57

14.3 Block Model Development .......................................................................................... 60

14.4 Statistical Analysis ...................................................................................................... 61

14.5 Variography ................................................................................................................ 61

14.6 Grade Estimation ........................................................................................................ 61

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14.7 Model Validation ......................................................................................................... 62 14.8 Mineral Resource Reporting ....................................................................................... 62

15 Mineral Res erve Es timates ................................................................................................... 66

16 Mining Methods ..................................................................................................................... 66

16.1 Mineral Resource Block Model .................................................................................... 66

16.2 Operating Parameters................................................................................................. 67

16.3 Parameters for Mine Design ....................................................................................... 67 16.3.1 Base Parameters ................................................................................................... 67 16.3.2 Lerch-Grossman Economic Shells Results ........................................................... 69

16.4 Pit and Mine Design Phase ......................................................................................... 71 16.4.1 Final Pit Design ..................................................................................................... 71 16.4.2 Tabulation of Pit Contained Resources ................................................................. 72

16.5 Mine Production Schedule .......................................................................................... 73 16.6 Mine Equipment .......................................................................................................... 75

16.7 Mine Personnel ........................................................................................................... 76 16.7.1 Salaried Staff ......................................................................................................... 76 16.7.2 Hourly Labor .......................................................................................................... 78

17 Recovery Methods ................................................................................................................ 80

17.1 Production Plants ....................................................................................................... 80 17.1.1 Mine Site ............................................................................................................... 80 17.1.2 Industrial Site ......................................................................................................... 85

18 Pro ject In fras tructu re ............................................................................................................ 92

18.1 Background ................................................................................................................ 92

18.2 Project Scope ............................................................................................................. 92 18.2.1 Production Units .................................................................................................... 92 18.2.2 Mine Site ............................................................................................................... 93 18.2.3 Industrial Site ......................................................................................................... 93

18.3 Project Location Characteristics .................................................................................. 93 18.3.1 Mine Site ............................................................................................................... 93 18.3.2 Industrial Site ......................................................................................................... 95

18.4 Access........................................................................................................................ 96 18.4.1 Road Access ......................................................................................................... 96 18.4.2 Waterway Access .................................................................................................. 97

18.5 Infrastructure Aspects ................................................................................................. 98 18.5.1 Mine Site ............................................................................................................... 98 18.5.2 Industrial Site ......................................................................................................... 99

18.6 Industrial Utilities ........................................................................................................ 99 18.6.1 Electrical Power Networks ..................................................................................... 99 18.6.2 Water Supply ....................................................................................................... 102

18.7 Logistic Aspects ........................................................................................................ 102 18.7.1 Raw Materials and Products ............................................................................... 102

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18.8 Project Implementation Schedule .............................................................................. 105

19 Market S tud ies and Contracts ............................................................................................ 106

19.1 Brazilian Market Overview ........................................................................................ 106 19.1.1 Agribusiness ........................................................................................................ 106 19.1.2 Fertilizer ............................................................................................................... 106

19.2 Region of Project Influence ....................................................................................... 109

19.3 Competitors .............................................................................................................. 110

19.4 Project Santana Competitive Advantage ................................................................... 111

20 Enviromen tal S tud ies , Permittings and Social o r Community Impact.............................. 112

20.1 Environmental Requirements .................................................................................... 112

20.2 Protected Areas ........................................................................................................ 113

20.3 Physical Environment ............................................................................................... 113 20.3.1 Climate and Weather ........................................................................................... 113 20.3.2 Soils ..................................................................................................................... 113 20.3.3 Hydrography ........................................................................................................ 113 20.3.4 Hydrology ............................................................................................................ 114 20.3.5 Flora .................................................................................................................... 115 20.3.6 Pasture Area ....................................................................................................... 117 20.3.7 Fauna .................................................................................................................. 118 20.3.8 Mammals ............................................................................................................. 119 20.3.9 Avifauna .............................................................................................................. 119 20.3.10 Herpetofauna (Reptiles and Amphibians) ........................................................... 120 20.3.11 Biota .................................................................................................................... 120 20.3.12 Anthropic - Social Indicators ................................................................................ 121 20.3.13 HDI ...................................................................................................................... 121

20.4 Infrastructure ............................................................................................................ 122 20.4.1 Electricity ............................................................................................................. 122 20.4.2 Sanitation ............................................................................................................ 122 20.4.3 Trade ................................................................................................................... 122 20.4.4 Public Safety ....................................................................................................... 124 20.4.5 Anthropic considerations ..................................................................................... 124

21 Capita l and Operating Cos ts .............................................................................................. 126

21.1 Capital Costs ............................................................................................................ 126 21.1.1 Mine Capital Costs .............................................................................................. 126 21.1.2 Mine & Industrial Sites Capital Costs .................................................................. 128

21.2 Operating Costs (OPEX) ........................................................................................... 129 21.2.1 Mine Operating Costs .......................................................................................... 129 21.2.2 Mine & Industrial Sites Operating Costs .............................................................. 132 21.2.3 Extra Operating Costs ......................................................................................... 135 21.2.4 Total SSP Annual Cost ........................................................................................ 135

22 Economic Analys is ............................................................................................................. 136

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22.1 SSP Long term price ................................................................................................. 136 22.2 Valuation model ........................................................................................................ 136

22.2.1 Assumptions ........................................................................................................ 136 22.2.2 Economics ........................................................................................................... 137 22.2.3 Sensitivity Analysis .............................................................................................. 138

23 Adjacen t Properties ............................................................................................................ 140

24 Other Relevan t Data and In fo rmation ................................................................................. 140

25 In terp re ta tion and Conclus ions .......................................................................................... 140

26 Recommendations .............................................................................................................. 141

26.1 Exploration and Resources ....................................................................................... 141

26.2 Mining ...................................................................................................................... 141

26.3 Mineral resource and Evaluation Budget ................................................................... 142

27 References .......................................................................................................................... 143

28 Date and S ignatu re Page .................................................................................................... 144

29 Certifica tes o f Qualified Pers ons ....................................................................................... 145

List of Tables Table 1_1 – Grade Tonnage Report 12

Table 1_2 – Indicative Economics 15

Table 1_3 – Santana Project - Proposed Resource and Evaluation Budget 16

Table 2.6_1 – List of Abbreviations 19

Table 4.2_1 – Summary of MBAC’s Permit Status in the Santana Region 22

Table 9_4– Summary Bulk Density 31

Table 10_1 – Summary Drilling Statistics 32

Table 12.2.1_1 – Standards Utilized by MBAC 37

Table 13.1.1_1 – Exploratory flotation tests with 1.5 kg outcrop samples 43

Table 13.1.2_1 – Criteria used to classify the drill core samples. 44

Table 13.1.2_2 – Expected Chemical Characteristics of the several types of material 44

Table 13.1.2_3 – Composite samples tested 45

Table 13.1.2_4 – Main conditions of the flotation tests 46

Table 13.2.1_1 – Mineralogical composition of three outcrops samples 47

Table 13.2.1_2 – Mineralogical identification of three outcrops samples 48

Table 13.2.1_3 – Chemical Characterization of three outcrops samples 48

Table 13.2.1_4 – Weight and P2O5 distribution in selected sizes of the outcrop samples crushed to minus 2mm 49

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Table 13.2.1_5 – Exploratory flotation results with sample RK 72833 49

Table 13.2.1_6 – Bond Work index determinations in outcrop samples 49

Table 13.2.2_1 – Chemical composition of selected drill core samples for mineralogical identifications 50

Table 13.2.2_2 – Mineralogical identification of selected drill core samples 50

Table 13.2.2_3 – Size distributions of several types of materials crushed to minus 2mm 51

Table 13.2.2_4 – Summary of flotation tests results with drill core samples 53

Table 13.3_1 – Chemical composition of typical concentrate. Fluorine analysis by Celqa. Other elements by pro-

solo 54

Table 13.4_1 – Estimation of the overall deposit recovery 54

Table 13.5_1 – SSP characteristics with outcrop sample and flotation concentrates after 15 days 55

Table 13.5_2 – Chemical characterization of a phosphoric acid produced with a flotation concentrate 55

Table 14.2_1 – Domains Wireframes 60

Table 14.3_1 – Block Model Summary 60

Table 14.4_1 – Summary Statistics – 3m Composites 61

Table 14.6_1 – Grade Estimation Parameters 61

Table 14.8_1 – Confidence Levels of Key Categorisation Criteria 63

Table 14.8_2 – Grade Tonnage Report 64

Table 16.3.1_1 – Lerch-Grossman Optimization Parameters 68

Table 16.3.2_1 – Lerch-Grossman Optimization Results 69

Table 16.4.1_1 – Pit Design Parameters 71

Table 16.4.2_1 – Resources Contained in Final Pit at Various Cut-off Grades 72

Table 16.5_1 – Waste dump design parameters 74

Table 16.5_2 – Mine Production Schedule 74

Table 16.6_1 – Peak Fleet Requirements for Commercial Production 75

Table 16.7.1_1 – Salaried Staff Requirements 77

Table 16.7.2_1 – Hourly Labour Requirement 79

Table 19.1.1_1 – Soybean export evolution 106

Table 21.1.1_1 – Summary of Mine Capital Costs (US$’000) 127

Table 21.1.1_2 – Equipment Unit Costs (US$’000) 127

Table 21.2.1_1 - Summary of Mine Capital Costs (US$’000) 129

Table 21.2.1_3– Summary of Mine Operating Costs - Total Dollars 131

Table 21.2.1_4– Summary of Mine Operating Costs – US$ per tonne 132

Table 21.2.2_1– Reagent Consumption and Vendor 134

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Table 21.2.2_2– Consumables Summary 134

Table 21.2.2_3– Raw Material Summary 134

Table 21.2.2_4– CAPEX factor for each project area 135

Table 22.1_1 – Real SSP Prices 136

Table 22.2.3_1 – NPV at Q3 2013 Sensitivity Analysis by SSP and Sulphur prices 138

Table 22.2.3_2 – NPV at Q3 2013 Sensitivity Analysis by SSP Price and Total CAPEX 139

Table 22.2.3_3 – NPV at Q3 2013 Sensitivity Analysis by SSP Price and Total OPEX 139

Table 22.2.3_4 – NPV at Q3 2013 Sensitivity Analysis by SSP Price and WACC 139

Table 26.3_1 – Santana Project - Proposed Resource and Evaluation Budget 142

List of Figures Figure 4.1_1 – Project Location 21

Figure 4.2_1 – Permit Location 22

Figure 7.1_1 – Regional Geology Map 26

Figure 7.2_1 – Project Geology Map 27

Figure 7.2_2 – Outcrop showing main lithological units 27

Figure 7.2_3 – Diamond Core Hole 28

Figure 7_1 – Apatite mineralization along fracture 28

Figure 9_1 – Ground Magnetic Survey 30

Figure 10.2.1_1 – Wet RC samples 32

Figure 11.1.2_1 and 2 – Recommended riffle splitters 35

Figure 12.2.1_1 – Blank material review 38

Figure 12.2.1_2 – Standard PFA review 39

Figure 12.2.1_3 – Standard PFM review 39

Figure 12.2.1_4 – Standard PFB review 40

Figure 12.2.2_1 – Field Duplicate Data 41

Figure 14.2_1 – Plan of Vertical Sections 57

Figure 14.2_2 – Supergene Saprolite Domain 59

Figure 14.7_1– Block Model Visual validation 62

Figure 14.8_1– Grade Tonnage Curve 65

Figure 16.3.2_1 – Tonnes and Grades versus optimization price 69

Figure 16.3.2_2 – Final pit shell – General Review 70

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Figure 16.4.1_1– Plan View Final Pit and Dump Designs 71

Figure 17.1.1_1– Process Flow Sheet 80

Figure 17.1.1_2– Beneficiation Flow Chart 83

Figure 17.1.1_3– Dam Project 84

Figure 17.1.2_1– Sulphuric Acid Flow Sheet 87

Figure 17.1.2_2– Acidulation Plant Flow Sheet 89

Figure 17.1.2_3– Granulation Flow Sheet 91

Figure 18.2.3_1– Industrial Site 93

Figure 18.3.1_1– Mine Site 94

Figure 18.3.2_1– Industrial Site 95

Figure 18.4.1_1– Location Plan 96

Figure 18.4.2_1– Water Access 98

Figure 18.6.1_1– High Tension Electricity 100

Figure 18.6.1_2– Electrical Power 101

Figure 18.7.1_1– Freight Route 103

Figure 18.8_1– Project Implementation Schedule 105

Figure 19.1.2_1– Brazilian P2O5 Market Scenario 106

Figure 19.1.2_2– Domestic Nutrients Consump 108

Figure 19.2_1 Project Area Immediate Market 109

Figure 19.3_1 Competitors 110

Figure 19.4_1 Competitors Distances to Santana Project Area 111

Figure 20.3.4_1– Farms in the project area 114

Figure 20.3.5_1– Area Comprised of small individuals 115

Figure 20.3.5_2– Vegetation Composed by pioneer species 116

Figure 20.3.5_3– Palm species identified in the study area 116

Figure 20.3.6_1– Pasture Area 117

Figure 20.3.6_2– Cleaning Areas of pastures 117

Figure 20.3.6_2– Streams in Permanent Preservation Areas 118

Figure 20.3.9_1– Avifauna 119

Figure 20.3.13_1– Health in São Félix do Xingu 121



Figure 20.4.3_3– Streets in São Félix do Xingu 123

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Figure 20.4.3_4– Roads in São Félix do Xingu 124

Table 21.3.2_1 – Capital Cost Breakdown 128

Figure 21.2.1_2 – Hourly Cost per Equipment 129

Figure 21.2.4_1– SSP Annual Costs 135

List of Appendices No table of figures entries found.


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1 SUMMARY

1.1 Introduction

Amazon Geoservices Ltda (Amazon Geoservices) and NCL Brasil Ltda (NCL) have been commissioned by MBAC Fertilizer Corp (MBAC) to prepare a preliminary economic assessment (PEA) for the Santana Phosphate Project, in Para State, Brazil.

Amazon Geoservices and NCL are collectively referred to in this report as “Amazon/NCL”

The PEA has been prepared under the guidelines of Canadian Institute of Mining (CIM) National Instrument 43-101 and accompanying documents 43-101.F1 and 43-101.CP (“NI43-101”).

1.2 Location

The Santana Phosphate Project is located in the southeast of Pará State near the state border of Mato Grosso and Para. Santana is located approximately 200km from the cities Santana do Araguaia, (Para State) and Vila Rica (Mato Grosso State). The topographical coordinates of the project are 417882 East and 8967784 North (Datum SAD69 Zone 22 South).

1.3 Ownership

MBAC, through its 100% owned Brazilian subsidiary Itafos Mineracao Ltda (Itafos), is the sole registered and beneficial holder of eight exploration properties with an additional three exploration permits under application for a total of 97,949ha.

1.4 Geology and Mineralization

The Santana property is located within the Iriri Group of hydrothermally-altered volcaniclastic and carbonate rocks of Precambrian age. The hydrothermally-altered volcaniclastic rock consists of lapilli and crystal tuff, with clasts varying considerably in size from very coarse sand-size clasts to cobble-sized particles (agglomerate, volcanic breccia).

The phosphate mineralization occurs in a hydrothermally altered (or more correctly metasomatic) lapilli tuff, consisting mainly of carbonate and/or silica. A higher grade supergene enriched zone has been interpreted to lay sub-horizontally within close proximity to the saprolite / fresh rock boundary. Lower grade mineralization has been intercepted in the deeper fresh rock drilling but has not been interpreted due to lack of data.

1.5 Mineral Processing and Metallurgical Testing

MBAC has commenced preliminary metallurgical testwork which has defined a preferred process route.

Testwork has focussed on 3 outcrop samples along with 9 composite samples from 22 diamond core holes. All testwork has been undertaken at the MBAC’s process facilities located at their Itafos phosphate mine in Brazil.

Mineralogical identification of outcrop and drill core samples indicated that main phosphate mineral is apatite and the main gangue minerals are quartz, iron oxides and clay minerals.

Flotation tests showed that the Santana samples responded well to the process with high grade concentrates easily obtained.

The following preferred process route has been defined by MBAC:


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Primary crushing of the ROM material;

Classification of the ore into two size fractions;

Grinding of the coarse product to 80% passing 100 microns;

Flotation of the grinded product;

Regrind of the rougher concentrate to 80% passing 44 microns;

Flotation of the regrinded rougher concentrate;

High intensity magnetic separation of the cleaner concentrate; and

Concentrate dewatering.

1.6 Mineral Resources

The Santana Inferred mineral resource estimate is based on 53 diamond holes (2,386m) and 6 RC holes (238m) drilled at a spacing of approximately 200m by 200m. Only data received as at 31 July 2011 has been used in this estimate. MBAC have an ongoing infill and extensional drillhole program underway.

The mineral resource estimate has focused on the main supergene enriched oxide mineralization with 2 flat lying mineralized domains defined using the saprolite and fresh geological boundary along with a 3% P2O5 grade cut-off to guide the wireframing process. An independent inferred mineral resource has been estimated comprising 33.5Mt with an average P2O5 content of 12.39% (using a preferred 3% P2O5 Cut-off). (Table 1_1 below)

The statement has been classified by Qualified Person Beau Nicholls (BSc (Geo) MAIG) in accordance with the Guidelines of NI 43-101 and accompanying documents 43-101.F1 and 43-101.CP. It has an effective date of 31 July 2011.

Table 1_1

Santana Deposit

Inferred Mineral Resource Grade Tonnage Report – 31 July 2011 Inverse Distance Weighted Power 2 (IDW2)

12.5E x 12.5mN x 3mRL

Lower Cutoff Grade (%P2O5)

Million Tonnes P2O5 %

CaO %

Fe2O3 %

Al2O3 %

SiO2 %

3 33.5 12.39 16.7 17.6 8.8 27.5 10 17.2 17.6 23.8 15.32 6.42 22.18 20 5.8 24.3 35.1 10.6 4.1 11.8


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1.7 Mining Operations

An open pit mine plan was developed by NCL for the Santana Phosphate Project to produce 300 k tonnes of phosphate concentrate at a grade of 34% P2O5 per year, with a peak total material movement rate of 4.2 M tonnes per year. This scenario results in the processing of approximately 1.5 M tonnes per year of ore with an average grade of 12.4 % P2O5. The expected mine life for this project is 24 years.

The reader is cautioned that the open pit mining study is a preliminary economic assessment that is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary assessment will be realized. No Mineral Reserves have been estimated. Mineral resources that are not mineral reserves do not demonstrate economic viability.

1.8 Indicative Economics

This document provides a report on the results of resources, mining, processing and a preliminary economic assessment of the potential project development. The preliminary economic assessment is based on the mineral resource estimate for the Santana Target as of 31 July 2011.

The indicative economics for the production of 300Ktpa (Base Case) of phosphate concentrate at a grade of 34% P2O5 is presented in Table 1_2 below. This demonstrates encouraging economics for the Santana Phosphate Project based on the cost projections and price assumptions as presented in this PEA.

However, readers are cautioned that the preliminary economic assessment is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary assessment will be realized. No Mineral Reserves have been estimated.

Based on project assumptions preliminary financial model indicates robust project economics with a NPV of US$423 million as at Q3 2013 (the estimated start of construction). This figure does not include US$20 million required for exploration and engineering development during 2011 and 2012.

The following economic results were obtained from the financial analysis:

• Capital cost of US$ 385 M, including US$ 73 M of contingencies and US$ 23 M of working capital.

• Sustaining capital of US$78 M, including:

Sustaining Capital – US$ 1 M each year and additional of US$ 2 M every 3 years;

Mine Equipments - US$ 23.3 M starting in 2016 and expend during the whole mine life;

Closure Cost - US$ 20 M, being US$ 10 M in 2037 and US$ 10 M in 2038.

• Operating cash costs - of $127 per tonne SSP.

• Internal Rate of Return (IRR) - 26.27% pa.

• Net Present Value: US$ 423M @ 10% WACC (Weighted Average Cost of Capital).


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Table 1_2 Indicative Economics


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1.9 Conclusions and Recommendations

Amazon/NCL considers that the proposed exploration and development strategy is entirely appropriate and reflects the potential of the Santana Phosphate Project.

Amazon/NCL has made a number of recommendations within this report to increase the mineral resource confidence and project development.

The Santana Phosphate Project demonstrates encouraging economics based on the scoping study concepts, cost projections and price assumptions as presented in this PEA.

Amazon/NCL recommends that the Project be advanced in one phase of work to the prefeasibility level of evaluation and design.

The cost estimate for the recommended work program is shown in Table 1_3 below

Table 1_3

Santana Phosphate Project

Proposed Resource and Evaluation Expenditure

Activity Total (US$) DC and RC drilling $ 3,000,000 Assaying and Characterization $ 200,000 Geology $ 50,000 Pre-Feasibility Work $ 500,000 Travel and accommodation $ 30,000 Field supervision and support $ 150,000 Administration $ 70,000 Sub-total $4,050,000


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2 INTRODUCTION

2.1 Scope of Work

Amazon Geoservices Ltda (Amazon Geoservices) and NCL Brasil Ltda (NCL) have been commissioned by MBAC Fertilizer Corp (MBAC) to prepare a preliminary economic assessment (PEA) for the Santana Phosphate Project, in Para State, Brazil.

Amazon Geoservices and NCL are collectively referred to in this report as “Amazon/NCL”

The PEA has been prepared under the guidelines of Canadian Institute of Mining (CIM) National Instrument 43-101 and accompanying documents 43-101.F1 and 43-101.CP (“NI43-101”).

2.2 Forward Looking Information

This report contains "forward-looking information" within the meaning of applicable Canadian securities legislation. Forward-looking information includes, but is not limited to, statements related to the capital and operating costs of the Santana Phosphate Project, the price assumptions with respect to phosphate materials, production rates, the economic feasibility and development of the Santana Phosphate Project and other activities, events or developments that the MBAC expects or anticipates will or may occur in the future. Forward-looking information is often identified by the use of words such as "plans", "planning", "planned", "expects" or "looking forward", "does not expect", "continues", "scheduled", "estimates", "forecasts", "intends", "potential", "anticipates", "does not anticipate", or "belief", or describes a "goal", or variation of such words and phrases or state that certain actions, events or results "may", "could", "would", "might" or "will" be taken, occur or be achieved.

Forward-looking information is based on a number of factors and assumptions made by the authors and management, and considered reasonable at the time such information is made, and forward-looking information involves known and unknown risks, uncertainties and other factors that may cause the actual results, performance or achievements to be materially different from those expressed or implied by the forward-looking information. Such factors include, among others, obtaining all necessary financing, licenses to explore and develop the project; successful definition and confirmation based on further studies and additional exploration work of an economic mineral resource base at the project; as well as those factors disclosed in MBAC’s current Annual Information Form and Management's Discussion and Analysis, as well as other public disclosure documents, available on SEDAR at www.sedar.com.

Although MBAC has attempted to identify important factors that could cause actual actions, events or results to differ materially from those described in forward-looking information, there may be other factors that cause actions, events or results not to be as anticipated, estimated or intended. There can be no assurance that forward-looking information will prove to be accurate. The forward-looking statements contained herein are presented for the purposes of assisting investors in understanding MBAC’s plan, objectives and goals and may not be appropriate for other purposes. Accordingly, readers should not place undue reliance on forward-looking information. MBAC does not undertake to update any forward-looking information, except in accordance with applicable securities laws.

http://www.sedar.com/�


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2.3 Principal Sources of Information

In addition to site visits undertaken by Mr. Beau Nicholls (Amazon Geoservices) and Mr. Carlos Guzman (NCL) to the Santana Phosphate Project between 25th and 27th July 2011, the authors of this report has relied extensively on information provided by MBAC along with discussions with MBAC technical personnel. A full listing of the principal sources of information is included in Section 27 of this report and a summary of the main documents is provided below:

MBAC (June 2011) - Internal Technical Report on Santana Phosphate Project, Pará State, Brazil.

MBAC (15th August 2011) - Internal Metallurgy Report on Santana Phosphate Project, Pará State, Brazil.

Letter dated 20 September 2011 - Magma Servicos de Mineracao Ltda, of Brasilia, Brazil

Geomma Consulting (2011) - Preliminary Environmental Assessment

ATS Engineering (2011) - Infrastructure review

Amazon/NCL have made enquiries to establish the completeness and authenticity of the information provided and identified. Amazon/NCL have taken all appropriate steps in their professional judgement, to ensure that the work, information or advice contained in this report is sound and Amazon/NCL do not disclaim any responsibility for this report.

2.4 Qualifications and Experience

The “qualified persons” (as defined in NI 43-101) for this report are Mr. Beau Nicholls and Mr. Carlos Guzman.

Mr. Nicholls is the principal consulting geologist for Amazon Geoservices with 17 years experience in exploration and mining geology. Mr. Nicholls is also a Member of the Australian Institute of Geosciences (MAIG) and is responsible for sections 4 to 12, 14 and jointly responsible for sections 1 to 3 and 23 to 29.

Mr. Guzman is the principal consulting mining engineer for NCL with 17 years experience in Mining engineering. Mr. Guzman is a registered Member of the Chilean Mining Commission and is responsible for Sections 13 and 15 to 22 and jointly responsible for Sections 1 to 3 and 23 to 29.

Neither Amazon/NCL nor the authors of this report has or has had previously any material interest in MBAC or related entities or interests. Our relationship with MBAC is solely one of professional association between client and independent consultant. This report is prepared in return for fees based upon agreed commercial rates and the payment of these fees is in no way contingent on the results of this report.

2.5 Units of Measurements and Currency

Metric units are used throughout this report unless noted otherwise. Currency is United States dollars ("US$").


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2.6 Abbreviations

A full listing of abbreviations used in this report is provided in Table 2.6_1 below.

Table 2.6_1 List of Abbreviations

Description Description $ United States of America dollars l/hr/m2 litres per hour per square metre “ Inches M million µ Microns m metres 3D three dimensional Ma thousand years AAS atomic absorption spectrometry Mg Magnesium Au Gold ml millilitre bcm bank cubic metres mm millimetres CC correlation coefficient Mtpa million tonnes per annum cm Centimetre N (Y) northing Co Cobalt Ni nickel CRM certified reference material or certified standard NPV net present value Cu Copper NQ2 Size of diamond drill rod/bit/core CV coefficient of variation ºC degrees centigrade DDH diamond drillhole OK Ordinary Kriging DTM digital terrain model P80 -75µ 80% passing 75 microns E (X) Easting Pd palladium EDM electronic distance measuring ppb parts per billion Fe Iron ppm parts per million G Gram psi pounds per square inch g/m3 grams per cubic metre PVC poly vinyl chloride g/t grams per tonne of gold QC quality control HARD Half the absolute relative difference QQ quantile-quantile HDPE High density poly ethylene RC reverse circulation HQ2 Size of diamond drill rod/bit/core RL (Z) reduced level Hr Hours ROM run of mine HRD Half relative difference RQD rock quality designation ICP-AES inductivity coupled plasma atomic emission spectroscopy SD standard deviation ICP-MS inductivity coupled plasma mass spectroscopy SG Specific gravity ISO International Standards Organisation Si silica kg Kilogram SMU selective mining unit kg/t kilogram per tonne t tonnes km Kilometres t/m3 tonnes per cubic metre km2 square kilometres tpa tonnes per annum kW Kilowatts UC Uniform conditioning kWhr/t kilowatt hours per tonne w:o waste to ore ratio


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3 RELIANCE ON OTHER EXPERTS

Amazon/NCL has relied on the independent lawyers Magma Servicos de Mineracao Ltda, of Brasilia, Brazil for their opinion on the title for the Santana mineral permits and Amazon/NCL have received a memorandum from them dated 20 September 2011, supporting MBAC’s claims.

MBAC is utilizing a number of experts in respects to the Santana Phosphate Project. They are listed below:

Metallurgy undertaken by Jose Ignacio – Senior Metallurgist MBAC (Section 13 of this report).

Environmental review– Geomma Consulting (Section 20 of this report).

Engineering Review -– ATS Engineering Projetos (Section 18 of this report).


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4 PROPERTY DESCRIPTION AND LOCATION

4.1 Project Location

The Santana Phosphate Project is located in the southeast of Pará State near the state border of Mato Grosso and Para. Santana is located approximately 200km from the cities Santana do Araguaia, (Para State) and Vila Rica (Mato Grosso State) (Figure 4.1_1). The topographical coordinates of the project are 417882 East and 8967784 North (Datum SAD69 Zone 22 South). Palmas city in Tocantins is the most convenient major city with commercial flights and light aircraft from Palmas takes approximately 1 hour to reach Santana.

Figure 4.1_1 Project Location


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4.2 Tenement Status

MBAC, through its 100% owned Brazilian subsidiary Itafos Mineracao Ltda (Itafos), is the sole registered and beneficial holder of eight exploration properties with an additional three exploration permits under application for a total of 97,949ha. Details of MBAC permits in the Santana region are found in Table 4.2_1 and Figure 4.2_1.

The exploration permits are valid for three years and are renewable for an equal period. MBAC are required to pay $ 2.02 Brazilian Reais per hectare per year to the DNPM for permit annual maintenance fee.

Table 4.2_1

Summary of MBAC Permits Status in the Santana Region

Permit Type Permit No. Date Granted Holder Area

(Ha) Comments

Exploration 851090/2008 24 September 2010 Itafós Mineração Ltda 9,828.91 Exploration

850335/2010 25 November 2010 Natanael Rodrigues da Silva 9,932.69 Transf. to Itafos.

Exploration 850381/2010 Under application Itafós Mineração Ltda 8,667.90 priority Exploration 850382/2010 Under application Itafós Mineração Ltda 9,731.39 priority Exploration 850383/2010 15 December 2010 Itafós Mineração Ltda 9,850.34 Exploration 850384/2010 15 December 2010 Itafós Mineração Ltda 9,946.74 Exploration 850385/2010 15 December 2010 Itafós Mineração Ltda 9,900.64 Exploration 850386/2010 15 December 2010 Itafós Mineração Ltda 9,995.82 Exploration 850387/2010 15 December 2010 Itafós Mineração Ltda 9,857.65 Exploration 850388/2010 15 December 2010 Itafós Mineração Ltda 9,862.03 Exploration 850567/2010 Under application Itafós Mineração Ltda 374.72 priority

Total 97 948.83

Figure 4.2_1 Permit Location


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4.3 Royalties and Agreements

Based on information supplied to Amazon/NCL by independent lawyers Magma Servicos de Mineracao Ltda, of Brasilia, MBAC has the following royalties and agreements:

Brazilian government royalty for phosphate is 2%

MBAC has a verbal agreement with three landowners in which MBAC has been granted full access to the properties in exchange for ongoing minor maintenance including fences, gates and repairs to a farm house

MBAC has exercised an option agreement with Natanael Rodrigues da Silva (dated 2 June 2010) to acquire 100% of permit numbers 851090/2008 and 850335/2010. Ongoing obligations are:

• Biannual progress reports describing work completed and results obtained

• 1% Net Smelter Royalty (NSR)

4.4 Environmental Liabilities

MBAC filed an Environment Drilling Permit (“Relatório de Controle Ambiental- RCA”) at the Secretaria Estadual de Meio Ambiente – SEMA although it should be noted that drilling is currently only being undertaken in privately owned farms with no clearing of access required.

Amazon/NCL is unaware of any environmental liabilities to which the Santana Phosphate Project is subject.

4.5 Permitting

No additional permits are required at the current stage of exploration.

Amazon/NCL is unaware of any other factors risking the development of the exploration works.


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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Project Access

Access to the Santana Phosphate Project is via approximately 125km of all weather gravel roads and 80km of paved roads from the regional city centres of Santana do Araguaia, in Para State (northeast of Santana) or via Vila Rica city in Mato Grosso State (southwest of Santana)

The Project is located on a large cattle station called Fazenda Santa Luzia. A nearby cattle station has a well maintained gravel airstrip which MBAC utilise with small airplanes requiring approximately a one hour flight from Palmas city in Tocantins state.

5.2 Physiography and Climate

The climate is tropical with an annual rainfall of around 2,000mm and seasonal variations with a drier period between June and November and a wetter period between December and May. The average annual temperature is approximately 27.5ºC with minimal month to month variations.

Exploration is undertaken at the Project all year round. Drilling activities are intensified during the drier season, from May to November, and continues with adequate equipment during the rainy season. At the peak of the rainy season drilling and mapping are paralysed for two to three weeks, when floods cover the bridges.

The region dominated by igneous massifs separated by alluvial lowlands. The lowlands lie at about 300 m ASL, with the tops of the massifs and tablelands at about 480 m ASL.

5.3 Local Infrastructure and Services

Santana is an exploration camp utilising the Fazenda Santa Luzia cattle station basic infrastructure which includes a house and sheds. All electricity is supplied by diesel generators and all consumables are brought in from the nearest major cities located over 200km away.

MBAC is in the process of moving the sample preparation and geological base to the small city of Vila Mandi which has electricity and medical services available.

The local economy consists mainly of cattle ranches.

A comprehensive review of local infrastructure and services is included in section 18 of this report.


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6 HISTORY

In June 2010, MBAC through its wholly owned Brazilian subsidiary, Itafos Mineração Ltda. (Itafos), acquired, via an option agreement, two Exploration Licenses numbers 850.335/2010 and 851.090/2008 from Mr. Natanael Rodrigues da Silva. These claims have been transferred to MBAC by the Federal Mines Department (DNPM).

In the same year, MBAC applied for nine Exploration Licenses, of which six were granted by the DNPM, and the remaining three have priority.

Mr. da Silva initially claimed the first area for lateritic nickel, but had identified the area as prospective for phosphate via discussions with local farmers who had returned anomalous phosphate levels in agricultural soil sampling.

Prior to MBAC involvement there had been no material exploration or mining activities undertaken on the project or the region.

7 GEOLOGICAL SETTING AND MINERALIZATION

7.1 Regional Geology

The Santana Phosphate Project is located within the Tapajós District (Figure 7.1_1) situated in the south-central portion of the Amazon Craton which became tectonically stable at the end of the Late Proterozoic period. The Craton is generally divided into the Guyana Shield north of the Amazon River and the Brazil Shield south of the Amazon River. The provinces have a northwest trend across the shields. The Brazil Shield has, as its nucleus, the Archaean granitoid - greenstone terranes of the Carajás-Imataca province in the east. The structural provinces become younger towards the west and are dominantly granitic rocks of Paleoproterozoic age. There is a general agreement that in this region, initial oblique collision tectonism was associated with crustal shortening linked to subduction and or accretion of magmatic arcs and early continental nucleation.

The main units that form the basement of the Tapajós Gold Province are the Paleoproterozoic Cuiú-Cuiú Metamorphic Suite (2.0 to 2.4Ga old), and the Jacareacanga Metamorphic Suite, also of possible Paleoproterozoic age (>2.1Ga), regionally mapped as Xingu complex. The Cuiú-Cuiú Suite comprises gneisses, migmatites, granitoid rocks and amphibolites. The Jacareacanga Suite comprises a supra-crustal sedimentary-volcanic sequence, which has been deformed and metamorphosed to greenschist facies. Both Suites are intruded by granitoids of the Parauari Intrusive Suite consisting of a monzodiorite dated at 1.9 to 2.0Ga. These form the basement of the extensive felsic to intermediate volcanic rocks of the Iriri Group, dated at 1.87 to 1.89Ga, including co-magmatic and anorogenic plutons of the Maloquinha and Rio Dourado Suite with intrusive events dated at 1.8 to 1.9Ga. The Iriri-Maloquinha igneous event is associated with a strong extensional period. Regional structural analysis in the Tapajós area has identified important lineaments that trend mainly northwest to southeast with a less well defined transverse east to west set.

The Santana Phosphate Project represents the first phosphate occurrence in the region.


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Figure 7.1_1 Regional Geology Map

7.2 Local and Property Geology

The Santana property is located within hydrothermally-altered volcaniclastic and carbonate rocks, possibly belonging to the Iriri Group. The hydrothermally-altered volcaniclastic rock consists of lapilli and crystal tuff, with clasts varying considerably in size from very coarse sand-size clasts to cobble-sized particles (agglomerate, volcanic breccia).

The rock has undergone weak to very strong carbonatization. Where carbonate has not been the principal hydrothermal process, then the rock may have been silicified (and/or albitized). Less extensive, but important locally are hematitization, chloritization, the introduction of manganese and possibly sericitization. Disseminated pyrite may occur up to an estimated 15%.

Karstification and solution weathering has resulted in the development of dolines (sink holes) that may be seen at surface and expressed as large depressions at least 15 m in diameter.

Meteoric weathering of this hydrothermalite led to the formation of a phosphate resource in the saprolite overlying bedrock (Pau Seco target).


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Figure 7.2_1 Project Scale Surface Geology

Figure 7.2_2 Outcrop showing main hydrothermal unit with overlying Tuff


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Figure 7.2_3 Diamond Core - 6m of fresh Hydrothermalite contained within the saprolite unit

7.3 Mineralization

The phosphate mineralization occurs in a hydrothermally altered (or more correctly metasomatic) lapilli tuff, consisting mainly of carbonate and/or silica. MBAC refer to this as a hydrothemalite. There is a basalt unit identified in the footwall, but this has not been intercepted in drilling to date.

The highest grade mineralization occurs in the saprolite. This residual mineralization was formed by the downward and lateral movement of meteoric waters that have weathered the crystalline carbonate bedrock to material that is often gritty in texture at the base, becoming more argillaceous vertically. However, the gritty passages may occur in any interval where the weathering has been less intense. There are manganiferous units, and units consisting of buff-coloured clay, interpreted as having once been more carbonatic.

The saprolite overlies the bedrock forming essentially tabular bodies, the geometry of which is controlled by the bedrock topography.

The thickness of the saprolite varies considerably from less than a metre overlying bedrock to greater than 75m. Limited drilling suggests that bodies are thicker near the base of the volcanic massifs becoming shallower outwards (15 to 20 m) towards the central parts of the alluvial plains.

Limited X-ray diffraction studies of 15 core samples from eight of the first 12 drillholes showed the most common phosphate-bearing mineral species in the saprolite to be Hydroxyapatite Ca5(PO4)3(OH) followed by Crandallite CaAl2(OH)6(PO3(O0.5(OH)0.5))2. Hydroxyapatite also occurs in the bedrock. Wavellite Al2(PO4)2(OH)3(H2O)5 and Goiazite SrAl3(PO4)2(OH)5 are present in minor amounts.


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Figure 7_1 Apatite mineralization along fracture within diamond core

8 DEPOSIT TYPES

The Santana Geological model which is described in more detail in section 14 of this report, is a post-tectonic hydrothermal related phosphate deposit with supergene enrichment. The origin of phosphate mineralization is currently unknown.

Amazon Geoservices is unaware of any similar occurrences of phosphate in Brazil.

The basis of which the exploration program is being planned by MBAC is covered in section 9 of this report.


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9 EXPLORATION

MBAC commenced exploration in December 2010 and has included project scale mapping, 20 line kilometres of ground magnetic surveys, a ground penetrating radar test survey along with Auger, Reverse Circulation (RC) and Diamond Core (DC) drilling.

9.1 Ground Magnetic Survey

MBAC undertook a total of 20km of Ground Magnetics at 100m spacing on east-west lines. The survey is displaced to the southwest of the current mineralization.

The E-W direction of the lines is also incorrect at this latitude, due to the proximity of the magnetic equator. Given these two factors, the result of the magnetic survey was considered to be of low quality and did not assist in interpretation or target definition.

The figure below shows the results of the survey relative to the location of the modeled ore body at the Pau-Seco target.

Figure 9_1

Santana Ground Magnetics Survey - TMI

9.2 Ground Penetrating Radar

A 5km trial ground penetrating radar survey was undertaken by Groundradar Measured Resources (Canada) using the Ultra GPR developed in-house. The aim of the trial was to verify if the equipment was capable of defining the base of saprolite (which represents the higher grade material defined by drilling).

The survey was partially successful in defining this boundary. It correlated well with the drilling were the GPR signal managed to penetrate the saprolite, but there are several gaps where


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the signal was 100% absorbed before reaching the target. It is believed that this is due to the local compact nature of the clay. The trial survey covers only parts of two lines in the target area.

9.3 Auger Drilling

MBAC has completed a total of 131 Auger holes for 1040m. Vertical holes were completed on a 50m by 700m grid utilising a “trado” Auger unit which takes a 30 – 50cm sample which is then recovered by pulling out all the rods. Minor contamination from wall material falling to the base of the hole occurs using this method of drilling. Holes were drilled between 5 and 10m in depth.

The Auger drilling has defined a regional anomalous zone which was then utilised in planning follow-up RC and DC drilling.

RC and DC drilling has been covered in the following section.

9.4 Bulk Density Determinations

MBAC has taken a total of 452 bulk density determinations from weathered and fresh DC.

Bulk Density measurements were undertaken by MBAC technicians using the following procedure:

20cm full core is wrapped in plastic film on the drill rig

Sample is weighed wet and then dried in a small oven

Dry core sample is weighed on electronic scale to determine mass of dry core and then weighed immersed in water to determine the volume (Archimedes method)

Both wet and dry bulk densities are then determined

Amazon Geoservices revised the density data and removed outlier data prior to determining the average bulk density. The Dry bulk density used in the resource estimation is summarized in the following table:

Table 9_4

Bulk Density for Santana Phosphate Project

Material Density g/cm3 Oxide – 368 samples 1.62 Fresh – 59 samples 2.68


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10 DRILLING

Drilling has been completed at the Santana Phosphate Project, as summarized in Table 10_1 below.

Table 10_1

Summary Drilling Statistics for Santana Phosphate Project

As of 31 July 2011

Item Company Number of Drillholes

Metres Drilled

Auger Drilling MBAC 131 1,040 Reverse Circulation Drilling Servitec Sondagens 57 2,432 Diamond Core Drilling Geosonda Sondagens 12 883.64 Diamond Core Drilling Servitec Sondagens 60 2,398.12 Total 260 6,753.86

All drilling has been undertaken and/or supervised by MBAC technical personnel.

RC and DC drilling and data collection methods applied by MBAC have been reviewed by Amazon Geoservices during the site visit. Auger drilling was complete and as such has not been assessed in operation by Amazon Geoservices, and is not included in the resource estimate.

10.1 RC Drilling

Drilling was contracted to Servitec Sondagens. RC Drilling equipment used on the project included a Explorak R50 RC rig and a 5.5 inch face sampling hammer.

Holes were drilled vertical and have not been surveyed downhole. Amazon Geoservices does not consider that these holes would deviate enough to make a material difference.

Observations:

Samples taken and weighed on meter by meter basis

Cyclone is cleaned on a rod by rod basis

Samples split to around 3kg via a single tier splitter

Logging of alteration, lithology and weathering

Hole collar coordinates picked up utilising a hand held GPS (accuracy +/- 10m)

10.1.1 RC Drilling Results and Quality

During the site visit Amazon Geoservices noted the drilling of hole SAN-RC-0050. The hole was in dry material at 50m but the intervals 10 to 25m were saturated.

On further inspection of the bag farm it was noted that a large number of RC samples were saturated (Figure below). The exact amount is not known at this stage as the sample humidity has not been clearly identified in the database (This is now in progress as it has been recorded on paper) Amazon Geoservices estimates that around 20% of RC samples are wet.


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Figure 10.2.1_1 Wet RC Samples at Bag Farm

Figure 9_1

Wet RC Drilling

This is a material issue as wet RC samples in the saprolitic material can ultimately create a sample bias. Amazon Geoservices has seen examples of phosphate being washed out of the sample (such as Apatite mineralisation along fractures) which ultimately can result in underestimation of the phosphate grade.

A total of 8 twin holes (RC and DC holes within 5 meters) have been undertaken to allow a comparative analysis of the results to determine the precision of the RC versus the DC. Results are pending.

Amazon Geoservices considers the wet RC samples should be avoided in RC drilling as this causes both downhole smearing and also washing. Until the twin hole results are available Amazon places a low confidence on the wet RC results to date.


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10.2 DC Drilling

Drilling was initially conducted by Geosonda who utilised a Chinese rotary drill called Drill XY-4. Productivity was poor after less than 900m drilled, so MBAC changed contractors to Servitec Sondagens. Servitec utilised a Boart Longyear DB-525 and a Maquesonda FS-320. Both Geosonda and Servitec drilling is dominantly HQ sized core with minor NQ sized core utilised on holes greater than 100m in depth and HW utilised to collar some holes.

Holes were drilled vertical and have not been surveyed downhole. Amazon Geoservices does not consider that these holes would deviate enough to make a material difference. Core has not been oriented as all holes are vertical.

Observations:

Storage of all core in wooden core boxes at drill site and then transported to the base for logging and sampling

Run markers with metal tags indicating drilled depth and recovery

Measurement and recording of core recovery for each drilling run

Photography and detailed logging of core before splitting

Detailed logging of alteration, lithology, structures and sulphides

Hole collar picked up utilising a hand held GPS (accuracy +/- 10m)

10.2.1 DC Drilling Results and Quality

From observations of drilling in progress and 4 drillholes reviewed during the site visit, Amazon Geoservices noted that MBAC DC procedures are to high quality with > 90% recovery returned in both saprolite and fresh material.

Amazon Geoservices considers the DC drilling procedures to be of an acceptable industry standard.

10.3 Drilling Results

Significant drill results have not been individually reported as this is a mineral resource estimate and would involve an extensive table that is summarised in the mineral resource section of this report. Drilling was orientated to enable perpendicular intercepts of the main trend of mineralisation but once again the three dimensional modeling has ensured that this is accounted for.


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11 SAMPLE PREPARATION, ANALYSES AND SECURITY

11.1 MBAC Sampling Method

11.1.1 Diamond Core

MBAC geologists supervised all core sampling undertaken. Core samples were normally taken on 1m intervals with some taken between 0.75m and 1.25m intervals based on the geological logging.

Core is split in half using a blade in the weathered material and via a diamond saw in the fresh material. The ½ core is bagged and sent for preparation and the remaining ½ core is returned to the core box and a ply wood lid is nailed on and the box is stored for future reference.

Amazon Geoservices recommends that the practice of irregular sample intervals should not be continued. The mineralization in Santana is often subtle with no clearly defined visible geological controls and as such a regular 1m sample interval for all the drillcore is recommended.

11.1.2 Reverse Circulation

MBAC utilise a single tier riffle splitter and pass the dry sample approximately 3 times to get a 3kg sample. Amazon Geoservices have recommended that MBAC purchase or fabricate a 3 tier riffle splitter to reduce the work load and reduce potential error. The single tier splitter MBAC was utilising was of poor design.

Figure 11.1.2_1

Single Tier Splitter currently used

Figure 11.1.2_2 Typical 3 Tier Splitter Recommended

For wet RC samples MBAC are drying the full sample in the sun (on a flat galvanised tin surface) and then using a quartering technique to determine a 3kg sample. During the site visit Amazon Geoservices recommended that they utilise the riffle splitter for all samples with the exception of saturated samples (they were drying moist samples that passed through the riffle splitter). Amazon Geoservices also recommended utilising a spear technique to sample the wet samples, but more importantly Amazon Geoservices recommended that the wet samples be removed from contention by increasing the drill rigs air capacity (by either


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employing a bigger compressor or adding a booster unit). The wet samples will be reviewed in conjunction with the twin holes completed.

11.2 Sample Security

Core is currently stored at the farm house MBAC has rented. After logging, core samples are marked for splitting and sampling by MBAC geologists. Each core sample is placed in a plastic bag which in turn is placed in a nylon bag for transporting via truck to the ALS sample preparation laboratory located in Campos Belos. The laboratory is approximately 500km by road from Santana.

MBAC is currently changing the sample preparation area from the rented farm house to the new base at Vila Mandi city located around 3 hours drive from Santana. Core transport will need to be done carefully as there is a risk of mixing up core during transport if not properly packaged.

Amazon Geoservices considers the core sampling security to meet current industry best practice.

11.3 ALS Laboratory Sample Preparation and Analysis

Sample preparation of samples taken by MBAC was performed by ALS Sample Preparation Laboratory located in Campos Belos, Goias State, Brazil. Sample preparation procedures are:

Drying and weighing of whole sample

Crushing of sample to -2mm

Sample homogenization and splitting to a 1kg sub-sample

Pulverization to 95% passing -150 mesh

Splitting of pulverized material to 50 gram pulp

Sample pulps are air freighted to the ALS analytical laboratory in Lima, Peru for oxide analysis. Sample pulps are analysed for phosphate using a using a lithium tetraborate fusion followed by XRF analysis (0.01% Detection limit).

The ALS analytical procedures are ISO 9001 certified and are in accordance with ISO/IEC 17025.

11.4 Adequacy of Procedures

The sampling methods, chain of custody procedures, and analytical techniques are all considered appropriate and are compatible with accepted industry standards.


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12 DATA VERIFICATION

12.1 Geological Database

Amazon Geoservices validated the Santana MBAC access database using the Gemcom Surpac Software System Database Audit tool with no inconsistencies noted.

A comparison of hardcopy assay and geological logging versus the digital database was performed on a total of 10% of the MBAC drillholes. No errors were identified with the original log and the digital database.

12.2 QAQC

MBAC has set in place a QAQC programme that included the submission of blanks, field duplicates and standards.

MBAC undertakes quality control on approximately 10% of the total samples prepared. This includes three blanks, three duplicates and three certified standards for every 90 rock samples submitted. The first QAQC sample in each batch is a blank to check that the system was clean.

12.2.1 Certified Standards and Blanks

MBAC utilized a total of 3 standards (These standards have been prepared and certified by SGS- Geosol) and one blank material (locally sourced and not certified). The results have been analysed by Amazon Geoservices and are summarised in the table below.

Table 12.2.1_1

Standards Utilized by MBAC

Submitted Blanks and Standards

Standard Name

Expected Value (EV)

(%)

+/- 2SD (EV) (%)

Date range No of Analyses

Minimum (%)

Maximum (%)

Mean (%)

% Within +/- 2 SD of EV

MBAC Submitted Blanks

Blank 0.01 0.009 to 0.011

Feb 2011 to Aug 2011 98 0.005 0.44 0.018 41.8

SGS Certified Standards

PFA 12.99 12.69 to 13.28

Apr 2011 to Aug 2011 31 12.9 13.2 13.04 100

PFB 2.49 2.44 to 2.54

Apr 2011 to Jul 2011 30 2.5 2.65 2.54 73.30

PFM 11.09 10.84 to 11.33

Apr 2011 to Aug 2011 51 10.95 11.65 11.18 96.08


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The blank material was sourced by MBAC from un-mineralized granite material. There are results for blanks up to 0.44% P2O5 with the average of the results being 0.018%. The causes for the low level values returned from the blank material are potentially related to the actual un-certified granite material utilised containing low levels of phosphate.

Figure 12.2.1_1 Blank

0

0.05

0.1

0.15

0.2

0.25

0.3

0.35

0.4

0.45

0.5P 2

O5

(%)

Date

Mbac Blank

Results

2 DL

Zero


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Figure 12.2.1_2 Standard PFA

Figure 12.2.1_3 Standard PFM

12.312.412.512.612.712.812.9

1313.113.213.313.4

P 2O

5(%

)

Date

Mbac Standard - PFA

Results

EV

2 SD

10.4

10.6

10.8

11

11.2

11.4

11.6

11.8

P 2O

5(%

)

Date

Mbac Standard - PFM

Results

EV

2 SD


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Figure 12.2.1_4 Standard PFB

It can be seen for the results of Standard PFB above, that 27% of all standards analysed have failed the tolerance test.

12.2.2 Field Duplicates

The field duplicate data has been analysed by Amazon Geoservices using a number of graphical comparative analyse methods. The objective of this is to determine relative precision levels between various sets of assay pairs and the quantum of relative error. This directly reflects on the precision of the sampling technique utilised.

MBAC completed field duplicate by splitting the crush reject and resubmitting for analysis. Based in the analysis, Amazon Geoservices concludes that the precision of field duplicates is acceptable as shown in the figure below.

2.3

2.35

2.4

2.45

2.5

2.55

2.6

2.65

2.7

31-Mar-11 30-Apr-11 31-May-11 30-Jun-11

P 2O

5(%

)

Date

Mbac Standard - PFB

Results

EV

2 SD


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Figure 12.2.2_1 Field Duplicate Data (Crushed reject resubmitted)


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12.2.3 Data Quality Summary

The standards data has shown a high accuracy (within 2 standard deviations) as reported by the SGS Geosol laboratory for PFA and PFM of the 3 certified standards utilised.

The third standard, PFB, returned 26 samples outside the tolerance limit. This poor result with PFB is potentially related to the quality of the standard as the other two standards returned acceptable precision, although additional umpire analysis of this standard will be required to qualify this issue.

For a period of time MBAC did not proactively reviewed the QAQC data on a batch by batch basis, and the failed PFB standards were not queried with the laboratory immediately after results arrived. Problems detected with the standard PFB, which is the only low level standard utilised, can either be caused by the standards poor quality or by the laboratory accuracy and the latter results in a low confidence on the results around this grade.

Amazon Geoservices has recommended to MBAC to review these failed standards with the laboratory and also to undertake umpire assays at a separate laboratory.

The field duplicate data has returned acceptable precision. No issues have been identified with the current conventional sampling method.


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13 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 Samples Preparation and Methods

13.1.1 Outcrop Samples

The first study was completed by ALS laboratories on three outcrops samples. (Samples MO1, M1 and M3). A quantitative mineralogical composition was completed on these three samples.

At the end of 2010, an additional three outcrop samples we sent for testwork at MBAC’s Itafos mine. The samples were crushed to minus 2mm and homogenised with the following testwork undertaken:

Size distributions

Chemical analysis

Mineralogical identifications

Work index determinations

Flotation response of the sample RK 72833

Three Flotation tests were completed on sample RK 72833 at a batch scale with variation on the grinding size and keeping the following conditions constant:

Table 13.1.1_1

Exploratory flotation tests with 1.5 kg outcrop samples

Operation Time (min)

Impeller speed (rpm) Reagent

Cell volume (l)

Conditioning 1 3 1200 Corn starch (1.67%) - 30 ml 4.3

NaOH (5%) – until pH = 10 4.3

Conditioning 2 2 1200 Soy bean oil (5%) - 20 ml 4.3 flotation rougher/

cleaner 6/3 1000 4.3/2.3

Samples were not de-slimed.


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13.1.2 Drill Core Samples

The samples used for the main test work were collected from the first 22 bore holes samples. The drill core samples after excluding the ones with a P2O5 grade less than 3% were classified according the chemical composition as follows:

Table 13.1.2_1

Criteria used to classify the drill core samples. LR= low relationship CaO/ P2O5 <1.18 Types Meters of Core % of total P2O5 (%) CaO/ P2O5

High grade 61.22 17.23 22.1 a 35 1.18-1.4

apatite 22.3 6.5 > 35 1.3-1.4

Medium grade 63.59 18.08 7.8 to 22 1.18-1.4

Medium grade LR 23.2 3.11 10.5 to 22 <1.18

High grade LR 11.3 6.50 22 to 27 <1.18 Fresh rock low

grade 71.0 20.06 3 a 6.2 7.8 a 18.6

Low grade rich soil 50.6 14.97 3 to 15 <1.18

Low grade 19.19 4.80 3 to 7 1.18-1.44

Low grade LR 31.07 8.76 3 to 12 <1.18

TOTAL 353.4 100.00

The samples used for the main test work were collected from the first 22 bore holes samples. The drill core samples after excluding the ones with a P2O5 grade less than 3% were classified according the chemical composition as follows:

For each type of material it was collected approximately 30 kg from several drill core samples. Table 13.1.2_2 presents the expected grades of the samples as per the ALS grades used for mineral resource estimations.

Table 13.1.2_2

Expected Chemical Characteristics of the several types of material classified according criteria on table 3.1

Type Assays(%)

P2O5 Al2O3 CaO Fe2O3 SiO2 CaO/ P2O5 apatite 37.11 1.12 50.76 2.72 0.96 1.37

high grade 29.42 3.46 38.47 8.17 11.21 1.31

high grade LR 24.64 9.61 26.69 16.39 7.16 1.08

medium grade 14.44 5.51 18.89 13.97 38.12 1.31

medium grade LR 15.78 13.34 15.43 19.91 19.41 0.99

Low grade rich soil 7.41 16.73 1.86 30.13 23.28 0.23

Fresh rock low grade 3.92 0.30 49.26 1.41 1.35 12.94

Low grade 5.10 10.52 6.58 15.68 52.84 1.29

Low grade LR 5.26 12.72 2.92 18.27 47.31 0.57

Material from each one of the types presented on table 13.1.2_3, excluding the “apatite” that was considered final product and the “fresh rock” that will not be mined was crushed to minus 2mm, homogenized and distributed into 1.5 kg plastic bags for the metallurgical tests.

The following tests were done:

Size distributions

Mineralogical identifications on selected drill core samples


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Chemical characterizations of material types and concentrates

Flotation tests

All the testwork was done in bench scale in the mineral processing laboratory of Itafos in Campos Belos, Goiais state. The following equipment was used:

One set of tyler mesh screens

One laboratory rod mill

One desliming unit composed of one pump box, one pump and one hidrocyclone 25mm diameter. Slurry was fed to the cyclone at 15% solids at a pressure around 220 kPag

One laboratory flotation cell with different tanks sizes

One laboratory attrition cell

The flotation tests 1 until 18 were done with the medium grade material type, tests 19 to 22 with the low grade material type and tests 23 to 27 with the other types. Tests 28 to 32 were done with composite samples with the following proportion:

Table 13.1.2_3

Composite samples tested

Test Level Weight (%)

28- Composite High grade 70%

Low grade 30%

29-Composite LR

Low grade LR 30%

Medium grade LR 30%

High grade LR 20%

Rich soil 20%

30-Global

High grade 35%

Low grade 15%

Low grade LR 15%

Medium grade LR 15%

High grade LR 10%

Rich soil 10%

32 Global2

High grade 15%

Low grade 35%

Low grade LR 15%

Medium grade LR 15%

High grade LR 10%

Rich soil 10%

The samples were screened on the specified size; the coarse product was ground until reach the desired size, mixed with the fine product and then scrubbed with NaOH in an attrition cell during 15 minutes before being deslimed in the desliming unit. The cyclone underflow was floated in the laboratory cell. In some tests the rougher concentrates were reground and in some tests the material was not deslimed or scrubbed.

Reagents were diluted to 5% weight before being fed. Starch and soy bean oil (collector) were prepared with NaOH in a concentration of 15% weight. Conditioning time was 3 minutes for


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depressor and 2 minutes for collector. Table 13.1.2_4 summarizes the main conditions. Detailed conditions are presented on annex 1- Batch flotation tests conditions and results.

Table 13.1.2_4

Main conditions of the flotation tests. Attrition was done with soda, except for test 20

Test

FLOTATION CONDITIONS

GRINDING 80%<# pH SOY

BEAN g/t STARCH

g/t Na2SiO3

g/t Desliming Attrition Cleaners Regrind

1 65 10 340 350 no no 1 no

2 100 10 340 350 no no 1 no

3 150 10 340 350 no no 1 no

4 65 10 340 350 yes no 1 no

5 100 10 340 350 yes no 1 no

6 150 10 340 350 yes no 1 no

7 65 7.5 340 350 no no 1 no

8 65 9 340 350 no no 1 no

9 65 10 340 350 no no 1 no

10 65 11 340 350 no no 1 no

11 65 9 340 170 no no 1 no

12 65 9 340 170 no no 1 no

13 65 9 340 430 no no 1 no

14 65 9 340 451.0 no no 1 no

15 65 9 340 340.0 no no 1 no

16 65 10 340 430.0 yes yes 1 yes

17 150 10 340 430.0 yes yes 2 yes

18 270 10 340 430.0 yes yes 2 no

19 270 10 340 430.0 yes yes 2 yes

20 150 10 340 430.0 yes yes 2 yes

21 150 10 340 430.0 yes no 2 yes

22 150 10 340 430.0 yes yes 2 no

23 150 10 340 430.0 yes yes 2 yes

24 150 10 340 430.0 yes yes 1 yes

25 150 10 340 430.0 yes yes 2 yes

26 150 10 340 430.0 yes yes 2 yes

27 150 10 340 430.0 yes yes 1 yes

28 150 10 340 430.0 yes yes 1 yes

29 150 10 340 430.0 yes yes 2 yes

30 150 10 340 430.0 yes yes 2 yes

31 150 10 340 430.0 yes yes 1 yes

32 150 10 340 430.0 yes yes 2 yes


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13.2 Results and Discussion

13.2.1 Outcrop Samples

The mineralogical composition is presented in table 13.2.1_1:

Mineralogical characterization

Table 13.2.1_1

Mineralogical composition of three outcrops samples

Minerals Wt%

Sample MO1 Sample M1 Sample M3 Iron_oxide 61.09 0.08 2.38

Iron_Oxide_Ti 3.19 0.01 0

Rutile 0.01 0 0.01

Manganosite 0.03 1 1.04

Gohetite 3.71 0.03 1.63

Apatite 17.64 88.84 20.42

Gorceixite 0 0.34 0.32

Kulanite 0.35 0.48 0.84

Childrenite 0.94 2.88 0.03

Trolleite 0.08 0.1 0.05

Cheralite 0 0.18 0.02

Monazite 0.09 4.08 0.08

Illite 0 0.04 0.01

Glauconite 0 0.05 0

Quartz 0.03 0.32 70.6

Mix_Iron_Oxide-Apat 12.82 1.05 2.34

Mix_Kulanit_Mangano 0 0 0.27 0.02

Apatite_Qz 0 0.01 0.24 0.21

Total 100 100 100

Sample MO1 was mainly composed of iron oxides, sample M1 with apatite and sample M3 with quartz. The main phosphate mineral was identified as apatite, secondary phosphates identified as Gorceixite, Kulanite, Childrenite and Trolleite presented minor proportion. Some association of apatite-iron oxides and apatite-quartz were also identified on some microscopic photos. Sample M1 presented a high grade of Monazite, 4.08%.


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Another mineralogical identification was done on three outcrop samples send by the geology team to Itafós mine at the end of 2010. Results are shown in Table 13.2.1_2.

Table 13.2.1_2

Mineralogical identification of three outcrop samples

Sample Minerals RK 72831 Hidroxiapatite, Hematite, Goethite

RK 72832 Hidroxiapatite

RK 72833 Quartz, Hidroxiapatite, Goethite , Crandalite

RK 72833CONC CLEANER Hidroxiapatite, Quartz

Tables 13.2.1_3, 13.2.1_4, 13.2.1_5 and 13.2.1_5 present the results obtained with the three outcrop samples.

Chemical Characterization

Table 13.2.1_3

Chemical Characterization of the three outcrop samples

SAMPLE RK 72831 RK 72832 RK 72833 RK

72833CONC CLEANER

%

P2O5 24.4 36.9 6.2 28.8 %

Solubility citric acid 2% (1:100)- P2O5

5.3 5.8 2.7 4.0 %

CaO 30.7 53.2 8.0 38.1 %

MgO. 0.10 0.087 0.083 0.066 %

Na2O 0.15 0.13 0.061 0.219 %

K2O 0.060 0.10 0.138 0.036 %

MnO. 0.56 1.02 0.983 0.848 %

Cu. 0.059 0.050 0.041 0.038 %

Co. 0.010 0.004 0.007 0.003 %

Fe2O3 2.55 1.75 8.23 4.36 %

Al2O3 1.1 1.0 1.5 0.90 %

SiO2 1.2 0.80 68.9 23.0 %

F. 0.55 0.68 0.57 0.69 %

As. < 0.10 < 0.10 < 0.10 < 0.10 mg/kg

Cd. 20.0 15.0 6.2 10.0 mg/kg

Pb. 300.0 175.0 175.0 250.0 mg/kg

Cr. 45.0 25.0 100.0 20.0 mg/kg

Hg. < 0.10 < 0.10 < 0.10 < 0.10 mg/kg

Ni 117.5 420.0 82.5 60.0 mg/kg

Se. < 0.10 < 0.10 < 0.10 < 0.10 mg/kg

ORGANIC MATER 0.65 0.60 0.85 0.62 %

UMIDITY 100º C. 0.45 0.32 0.70 0.33 %

LOSS IGNITION 460º C. 1.3 1.1 1.8 1.4 %

LOSS IGNITION 1.050º C. 2.4 3.0 3.0 2.3 %


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Size and P2O5 distributions

Table 13.2.1_4

Weight and P2O5 distribution in selected sizes of the outcrop samples crushed to minus 2mm

Size (mm)

RK 72831 RK 72832 RK 72833

Weight (%)

P2O5 (%)

P2O5 Distr(%)

Weight (%)

P2O5 (%)

P2O5 Distr. (%)

Weight (%)

P2O5 (%)

P2O5 Distr(%)

+2 14.71 26.16 15.00 18.65 34.64 18.60 15.38 8.91 32.27

-2+0.21 24.90 24.7 23.00 33.11 36.62 34.90 26.06 4.17 29.63

-0.21+0.037 13.78 19.11 10.27 14.39 35.43 14.68 28.74 2.20 17.27

-0.037 46.61 28.45 51.73 33.85 32.66 31.82 29.82 2.55 20.73

TOTAL 100.00 25.64 100.00 100.00 34.74 100.00 100.00 3.67 100.00

Table 13.2.1_5

Flotation tests

Exploratory flotation results with sample RK 72833

Test nº Grinding size 85% passing

Rougher Tailings Cleaner Concentrate Flotation recovery

% weight %P2O5 % weight %P2O5 % weight

% P2O5

1 65# 81.25% 2.05% 11.22% 26.02% 11.22% 58.70%

2 100# 81.72% 2.16% 10.90% 27.65% 10.90% 58.29%

3 150# 86.56% 2.20% 9.26% 29.47% 9.26% 56.40%

Table 13.2.1_6

Bond Work Index

Bond Work index determinations in outcrop samples

Sample test screen

(mm)

% passing of the feed in test screen

F80

(mm) P80

(mm) Gbp

(g/rev) WI

(kWh/st) WI

(kWh/t)

RK 72831 0.149 50.5 1.00 0.096 5.549 4.9 5.4

RK 72832 0.149 43.8 1.26 0.112 3.040 8.5 9.4

RK 72833 0.149 58.8 1.00 0.090 8.143 3.4 3.7

Bond work index presented low values, from 3.7 to 9.4 kWh/t. A value of 10kWh/t was used in the preliminary assessment study to size the mill.


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13.2.2 Drill Core Samples

After the results of the first drilling campaign some drill core samples were chosen for mineralogical identification based on the chemical compositions. Table 13.2.2_1 and 13.2.2_2 below present the results.

Chemical Characterizations

Table 13.2.2_1

Chemical composition of selected drill core samples for mineralogical identifications

Drill hole depth (m) P2O5 Al2O3 CaO Fe2O3 MgO MnO2 SiO2 CaO/ P2O5

SAN-DD-0001 3 4 29.8 2.09 40 2.58 0 1.93 19 1.33

SAN-DD-0001 10 11 17.6 15.8 14 25.1 0.3 3.26 9.3 0.78

SAN-DD-0001 19 20 4.14 13.2 2 18.6 0.2 0.63 47 0.49

SAN-DD-0001 49 50 3.49 0.08 55 0.45 0.9 0.47 0.2 15.82

SAN-DD-0001 108 109 3.04 0.09 55 0.21 0.8 0.41 0.4 18.03

SAN-DD-0002 5 6 30.7 4.6 40 11.9 0.1 2.14 2.2 1.31

SAN-DD-0005 25 26 14.7 6.89 20 8.15 0.4 1 43 1.39

SAN-DD-0007 1 2 26.7 9.43 30 17.8 0.2 1.46 8.3 1.10

SAN-DD-0008 1 2 17.8 11.2 17 22.5 0.2 1.42 20 0.93

SAN-DD-0010 36 37 20.9 4.33 28 6.42 0.1 1.76 34 1.32

SAN-DD-0010 57 58 4.43 6.98 5.4 13.9 0.3 1.28 62 1.22

SAN-DD-0011 3 4 12.1 18.5 1.8 28.2 0.1 9.73 11 0.15

SAN-DD-0011 13 14 35.9 1.56 49 3.77 0 2.42 1.5 1.36

SAN-DD-0012 21 22 6.27 13.3 7.9 13.1 0.6 1.03 50 1.26

SAN-DD-0012 34 35 32.2 0.35 43 3.66 0 1.72 17 1.34

Table 13.2.2_2

Mineralogical Identifications

Mineralogical identification of selected drill core samples

Drill hole Depth (m) Minerals SAN-DD-0001 3 4 hidroxiapatita,quartzo,Crandalita

SAN-DD-0001 10 11 hidroxiapatita, Crandalita, Goetita,Muscovita, Hematita, Caulinita, Enstatita,

Anatasio, Goiazita

SAN-DD-0001 19 20 Quartzo,ortoclásio,hematita

SAN-DD-0001 49 50 Calcita,hidroxiapatita,dolomita

SAN-DD-0001 108 109 Calcita,hidroxiapatita,dolomita

SAN-DD-0002 5 6 hidroxiapatita, crandalita, hematita, goetita, caulinita

SAN-DD-0005 25 26 Quartzo, hematita, caulinita, muscovita,anatasio, hidroxiapatita

SAN-DD-0007 1 2 hidroxiapatita, quartzo, hematita,crandalita

SAN-DD-0008 1 2 hidroxiapatita, quartzo, hematita, goetita, wavelita, muscovita, anatásio.

SAN-DD-0010 36 37 hidroxiapatita, quartzo, hematita, caulinita

SAN-DD-0010 57 58 quartzo, hematita, hidroxiapatita, goetita, muscovita caulinita

SAN-DD-0011 3 4 hematita, quartzo, goiazita, magnetita, goetita, piroxmangita

SAN-DD-0011 13 14 hidroxiapatita, quartzo, goiazita

SAN-DD-0012 21 22 quartzo, hematita, hidroxiapatita,caulinita, muscovita

SAN-DD-0012 34 35 hidroxiapatita, quartzo, hematita

The main phosphate mineral identified was the hidroxiapatite followed by some secondary phosphates as crandalite, goiazite and wavelite. Iron minerals identified were goetite,


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hematite and magnetite. Silicates found the quartz, ortoclasio, muscovite, caulinite, and enstatite. Titanium and manganese oxides were identified as anatasio and piroxmanguite. Only at the bottom of the deposit were calcite and dolomite found.

Table 13.2.2_3

Size distributions

Size distributions of several types of materials crushed to minus 2mm

size (mm)

% weight accumulated on size

apatite High grade

Medium grade

Low grade

High grade LR

Medium grade LR

Rich soil

Low grade

LR +2 15.49 2.09 3.19 2.43 2.35 1.53 3.13 2.70

+0.21 43.93 13.74 15.75 19.90 16.18 12.72 18.13 19.89

+0.037 62.03 30.33 29.86 36.43 29.42 27.49 31.24 33.17

-0.037 39.77 69.67 70.14 63.57 70.58 72.51 68.76 66.83

The very rich material, classified as “apatite” presented around 40% of the weight finer than 37 microns but all other types were much finer and presented from 64% to 73% of the weight finer than 37 microns.

The type medium grade was chosen to study the main flotation variables. After this definition these conditions were used to estimate the recovery and the concentrate qualities that could be gotten with other types. Results are shown on Table 13.2.2_4.

Flotation tests

Tests 1 to 6 were done with variations on the grinding size, being tests 1 to 3 without desliming and tests 4 to 6 with desliming. All other variables were kept constant. In the first three tests the grinding size was varied from 85% < 65# to 85% < 150# without desliming the sample. The results did not show any improvement in grinding finer than 65#. Concentrates presented a grade around 29% P2O5 and recovery stay around 73%.

Effect of grinding size

Tests 3 to 6 the samples were deslimed after grinding each sample from 85% < 65# to 85%< 150#. Results show an increasing global P2O5 recovery, from 68.6% to 73.1%. It is Interesting to note that flotation recovery of the deslimed ore (85.4% test 6) is higher than the recovery without desliming (72.3% test 3), as shown on annex 1, although global recovery is similar. Desliming showed to be important to reduce the mass going to flotation and to eliminate iron and aluminium from flotation feed. Grinding size of 85% passing 150# showed to be the best grinding size.

Tests 7 to 10 pH was varied from 7.5 to 11. All other variables were kept constant. Results show a strong effect of the pH. P2O5 Recovery increased from 13.6% at pH 7.5, test 7 to 79.5% at pH 11, test 10.

Effect of the pH


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Tests 11 to 15 tested the corn starch against the sodium silicate in two concentrations. All other variables were kept constant. Better P2O5 grade and recovery, respectively 32.2% and 69.7 was received with sodium silicate at a concentration of 430 g/t.

Effect of depressor type and concentration

The concentrates received up to and including test 15 presented high grades of Fe2O3 and Al2O3 that could be deleterious for the acidulation in order to get a minimum of 18% solubility in citric acid (SSP specifications). Tests 16 to 18 the ground samples to 80% less than 150# were scrubbed with NaOH before desliming and the rougher concentrates were reground and then floated twice. It was possible to obtain concentrates with high P2O5 grades and lower Al2O3 and Fe2O3 grades. Based on these results, these conditions were used to evaluate other material types. Tests 19 to 22 were done with the low grade type material with variations on attrition and regrind steps showed a positive effect of these two variables to obtain low aluminum and iron grades in the concentrates.

Effect of attrition with soda and concentrate regrind


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Table 13.2.2_4

Summary of flotation tests results with drill core samples.

Types of Materials

P2O5 grade(%)

test #

CONCENTRATES

Assay(%) Weight

(%)

P2O5 recovery (%)

kgR2O3 /t

P2O5

kgFe2O3 /t P2O5 P2O5 CaO Fe2O3 Al2O3 SiO2

Medium grade 16.5-18

1 29.7 33.6 7.0 1.8 14.1 43.3 73.2 297.17 236.40

2 29.4 33.3 6.4 1.4 14.2 42.1 70.3 265.99 217.69

3 29.5 43.1 4.3 6.3 11.6 44.7 72.5 357.06 144.71

4 30.3 33.9 6.0 1.1 13.6 38.4 68.6 235.22 197.89

5 29.5 40.6 5.4 5.5 14.1 41.8 71.1 369.05 182.88

6 29.1 32.7 8.2 1.4 16.8 42.2 73.1 329.52 282.05

7 32.1 41.0 8.3 1.7 3.7 13.5 13.6 310.59 258.26

8 31.8 46.5 4.7 5.6 8.9 32.2 61.3 323.74 147.03

9 29.2 42.8 5.8 5.0 12.1 43.9 73.7 371.11 199.98

10 27.9 31.8 5.5 2.0 16.3 51.9 79.5 268.63 198.00

11 31.6 40.6 4.1 1.3 12.2 34.9 66.2 169.98 129.82

12 31.0 39.4 4.8 1.5 12.4 36.2 66.0 203.52 154.53

13 32.2 38.3 5.2 1.2 11.6 36.6 69.7 198.55 160.17

14 31.6 40.6 4.1 1.3 12.2 36.4 66.2 169.98 129.82

15 32.1 41.1 4.2 1.6 11.1 36.2 67.8 180.52 130.86

16 35.1 45.74 2.4 0.40 10.54 27.8 57.9 80.82 69.44

17 35.8 47.5 2.1 0.4 8.2 27.6 59.1 69.43 59.05

18 34.1 43.4 2.4 0.4 11.1 28.6 56.7 82.92 71.53

Low grade 5-6

19 35.4 45.0 2.4 0.6 6.4 4.4 31.8 84.58 67.89

20 34.0 42.0 2.8 0.7 10.1 5.7 40.1 101.94 82.26

21 33.7 42.2 2.8 0.7 9.6 6.1 43.3 105.11 83.14

22 33.8 45.2 3.0 1.4 9.4 6.4 44.6 131.32 88.73

Medium grade LR 16.2 23 33.6 35.3 7.6 4.4 2.6 22.3 46.2 357.61 225.92

High grade LR 24.3 24 33.7 37.5 3.6 3.5 10.6 30.1 44.1 211.55 106.64

Low grade LR 4.5 25 33.6 31.1 4.8 6.8 9.0 2.5 18.3 344.65 142.98

High grade 29.2 26 36.92 48.00 1.60 0.20 6.84 57.0 72.0 48.75 43.34

Rich soil 6.6 27 35.0 18.8 4.8 18.4 6.1 1.0 5.4 661.71 137.14

Composed 21.7 28 34.15 44.81 2.51 0.44 10.40 47.0 74.1 86.48 73.65

Composed LR 12.1 29 35.65 42.28 5.20 3.38 3.12 13.0 36.7 240.67 145.86

Global 17.1 30 35.91 45.22 2.80 1.15 6.29 28.0 57.1 110.00 77.97

Composed 2 - 12.2 31 34.36 45.80 2.57 0.54 8.93 24.3 68.4 90.52 74.88

Global 2 12.4 32 35.88 45.92 3.20 1.77 5.68 17.2 49.9 138.52 89.19


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13.3 Concentrate Characteristics

Table 13.3_1 shows the chemical composition of the concentrates obtained in the several tests. A typical concentrate is the one from test 18.

Table 13.3_1

Chemical composition of typical concentrate. Fluorine analysis by Celqa. Other elements by pro-solo

Assay(%)

P2O5 CaO Fe2O3 Al2O3 SiO2 F

34.1 43.4 2.4 0.4 11.1 1.7

13.4 Recovery Evaluation

Tests 19 to 32 evaluated the concentrates chemical characteristics and the P2O5 recovery of the several types of materials to estimate the deposit overall recovery to be used in this PEA. Flotation conditions were the same as the ones for tests 16 to 18.

Although it was obtained good concentrate grades for all material types, the iron and the aluminium of the concentrates of the material types classified as “medium grade LR”, “Low grade LR” and “rich soil” were very high and as so, these types of materials were considered as “non recoverable material”. It can be observed in table 4.3 that the amount of these materials are very low and represent only 11.45% of the total P2O5 of the deposit. Table 13.4_1 below presents an estimation of the overall deposit P2O5 metallurgical recovery based on the estimation of the proportion of each type of material and the metallurgical recovery of each one.

Table 13.4_1

Estimation of the overall deposit recovery

TYPES % of total Grade P2O5 (%)

P2O5 distribution

(%)

P2O5 recovery

(%)

P2O5

recovered (%)

apatite 6.50 37.11 13.69% 100.00 13.69

High grade 17.23 29.42 28.77% 72.02 20.72

Medium grade 18.08 24.64 25.28% 50.90 12.87

Medium grade LR 3.11 14.44 2.55% 0.00 0.00

Low grade fresh rock 6.50 3.92 1.45% 0.00 0.00

High grade LR 20.06 15.78 17.97% 44.10 7.92

Low grade rich soil 14.97 7.41 6.29% 0.00 0.00

Low grade 4.80 5.10 1.39% 43.30 0.60

Low grade LR 8.76 5.26 2.61% 0.00 0.00

TOTAL 100.00 17.62 100.00% 55.80 55.80

Based on this, a metallurgical recovery of 55% of the total P2O5 contained in the mineral resource estimation was found.


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13.5 Solubility Tests

Several acidulation tests with sulphuric acid were done at the Celqa laboratory in order to confirm the possibility to obtain Single Super Phosphate from the Santana material. Table 13.5_1 summarizes the results. It is possible to observe that the products met the SSP specifications, minimum of 18% solubility in citric acid (CNA) + water and 16% in water.

Table 13.5_1

SSP characteristics with outcrop sample and flotation concentrates after 15 days of the acid attack. Products of tests 3, 4 and 5 were dried in an oven at 65 °C.

test Sample (100g) H2SO4 (g) Water

(g)

P2O5 Total (%)

P2O5

CNA +water

(%)

P2O5 sol

water (%)

Free acidity

(H3PO4) (%)

F (%)

Moisture (%)

6 RK72832 70 35 21.5 20.2 19.1 5.6 0.83 -

7 RK72832 69 32 21.5 20.1 19.0 5.6 0.83 -

3 Conc. 16 62 29 21.7 20.4 19.6 5.0 0.83 0.0

4 Conc. 17 60 28 21.7 20.6 19.3 4.6 0.82 0.0

5 Conc. 18 54 25 21.0 20.0 18.6 4.4 0.84 0.0

A preliminary test for phosphoric acid characterization was done with a 2kg flotation concentrate obtained using the same conditions of test 16. Results are shown on Table 13.5_2.

Table 13.5_2.

Chemical characterization of a phosphoric acid produced with a flotation concentrate

Compounds (%) Flotation concentrate Phosphoric acid produced P2O5 36.60 54

P 16.27 24

CaO 48.00 1.03

Ca 34.29 0.74

SiO2 4.00 -

Al2O3 0.23 0.18

Fe2O3 1.10 0.37

SO3 <0.10 0.02

K2O 0.24 0.02

TiO2 <0.1 <0.01

MnO 0.55 0.02

SrO 0.15 <0.01

MgO <0.01 0.01

ZnO <0.01 0.01

ZrO2 <0.01 <0.01

Y2O3 <0.01 <0.01

BaO 0.22 <0.01

F 2.5 0.41

The concentrate was attacked with 98% sulphuric acid at a rate of 2.42m3/td. P2O5 under 75°C, the slurry was filtrated; the cake was washed with water at a rate of 0.65 parts of water


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to each part of slurry and the liquor was evaporated at 78°C. At this stage the liquor was not defluorinization.

The acid produced after elimination of the fluorine presents optimum quality to produce di-calcium phosphate.

13.6 Conclusion

Concentration of great part of the material from the deposit using anionic flotation was shown to be technically feasible. High P2O5 with low Fe2O3 and Al2O3 grades concentrates are easily obtainable by the process after attrition with soda, desliming and regrind of the first concentrate.

Material classified as “low grade rich soil” must be considered as waste since very low recovery was obtained. Concentrates generated from the material types “low and medium grades LR” presented high Fe2O3 and Al2O3 grades and were not considered as recoverable in this study. Mining studies must evaluate the possibility to waste this material. If this is possible, the average grade of the mineable material would increase with a consequent reduction of the capacity of the concentrator reducing the capital and operating costs of this unit. The metallurgical recovery of the deposit based on the first 22 bore holes was estimated as 55%.

The concentrator must have all the facilities and utilities to:

Crush the ore to minus 150mm.

Classify it in screens to 6mm.

Scrubber the minus 6mm fraction in attrition cells

De-slime the scrubbed fraction to eliminate the natural slimes, below 5 microns to improve flotation responses.

Grind the plus 6mm and the de-slimed fraction to 80% less than 150 microns.

Float in column cells the ground material.

Regrind the rougher concentrate

Float in column the reground rougher concentrate

Use a High Intensity Wet Magnetic Separation on the cleaner concentrate.

Dewater the concentrate in a thickener and a filter.


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14 MINERAL RESOURCE ESTIMATES

14.1 Introduction

Amazon Geoservices along with Leandro Silva (MBAC resource geologist) have estimated the Mineral Resource for the Santana Phosphate Project utilising drilling data as of 31st July 2011. All grade estimation was completed using inverse distance weighted to the power 2 (IDW2) for P2O5. This estimation approach was considered appropriate based on review of a number of factors, including the quantity and spacing of available data, the interpreted controls on mineralization, and the style of mineralization. The estimation was constrained within the saprolite domain. There is not sufficient data to interpret the fresh material.

The mineral resource estimate is based upon 53 diamond holes (2,386m) and 6 RC holes (238m). Auger holes were utilised only in the geological model and 3D wireframing.

14.2 Geological Modeling

At this stage no lithology model has been defined as the majority of drilling has been completed mainly within the saprolite of the hyrothermalite unit. A Saprolite/Fresh boundary has been constructed utilizing the geological logging and is an important grade control boundary (along with being utilised to determine the bulk densities) as the highest grade mineralization is constrained within the supergene enriched saprolite.

Amazon Geoservices and MBAC have interpreted a saprolite domain utilizing a 3% P2O5 lower limit to guide the interpretation. A total of 20 oblique vertical sections have been created snapping to drillholes. Two separated mineralized domains were defined reflecting the lack of drilling amidst them. Additional drilling will potentially join these two domains together. The interpretation and wireframe models have been developed using the Gemcom mine planning software package.


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Figure 14.2_1 Plan view of Block Model

The figure below shows a cross section of the main mineralized domain. The fresh rock has not been modelled due to lack of drill data but highlights a consistent large tonnage potential with grades between 1% to 3% P2O5.


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Figure 14.2_2 Supergene Saprolite Domain


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The block model has been assigned the following attributes to allow filtering and reporting.

Table 14.2_1

Domains Wireframes

Domain Block Attribute - Geology Saprolite Supergene domain 0 Saprolite Supergene domain – Above topo (Air) 1000

14.3 Block Model Development

A three-dimensional block model was defined for the Santana deposit, covering the interpreted supergene saprolite domain and including suitable additional waste material to allow later pit optimisation studies.

A parent block size of 12.5mE x 12.5mN x 3mRL has been used. No sub-blocking occurs in Gemcom. The attributes coded into the block models included mineralization, grade and weathering. A visual review of the wireframe solids and the block model indicates robust flagging of the block model. Bulk density has been coded to the block model based on the defined density values listed in section 9.4 with oxide material having a bulk density of 1.62 g/cm3. No fresh material has been estimated due to lack of drillhole data.

Table 14.3_1 below shows the summary of the block model created.

Table 14.3_1

Blo ck Mo d e l S um mary

Y X Z Minimum Coordinates 8965500 412500 155 Maximum Coordinates 8968000 414500 350 User Block Size 12.5 12.5 3 Rotation - - 320 Extension 2000 2500 195


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14.4 Statistical Analysis

The drillhole database was composited to a 1m down-hole composite interval, recording the geological model. The 1m composites were used for all statistical, geostatistical and grade estimation studies.

Statistical analysis of the composite saprolite supergene domain is presented in Table 14.4_1.

Table 14.4_1

Summary Statistics – 1m Composites inside Saprolite Supergene Domain

P2O5 (%)

Supergene Saprolite Domain

Count 514 Minimum 0.01 Maximum 38.7 Mean 13.93 Std. Dev. 10.76 CV 0.77

A top cut was applied to the composite P2O5 data based on the 99.28% percentile which effectively cut one sample back to 35% P2O5. This has minimal effect on the final mineral resource estimate.

14.5 Variography

Variography was attempted but lack of data (>200m drill spacing) resulted in poor variograms.

14.6 Grade Estimation

The interpolation method used was inverse distance weighted to the power 2 (IDW2) for P2O5, and the same parameters were also applied to other oxides reported. The selection of samples used in the interpolation is made by an ellipsoidal search without octants. The search ellipse was configured as shown in the table below. As selection criteria, samples used for the interpolation was limited on minimum 2 samples being maximum 1 per hole.

Table 14.6_1 IDW2 Grade Estimation Parameters Pass 1 X 250 Y 250 Z 60 Min Nr octants - Max per octant - Min N Comp. 2 Max N Comp 16 Nr of discretizations - Classification Inf Ellipse Axis Direction

Azimuth Dip 0 0

320 0 0 0


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The search ellipse was configured to match the main mineralization direction, which is sub- horizontal. The dimension of the search ellipse radii was 250x250x60m.

14.7 Model Validation

In order to check that the estimation has worked correctly, the model has been validated through a visual comparison through the generation of validation plots.

An example of the visual validation is shown below with a cross section of the block model compared against the drillhole results.

Figure 14.7_1 Block Model Visual Validation

14.8 Mineral Resource Reporting

The grade estimates have been classified as inferred mineral resource in accordance with NI 43-101 guidelines based on the confidence levels of the key criteria that were considered during the mineral resource estimation. Key criteria are tabulated below.

A summary of the estimated resources for the Santana deposit is provided in Tables 14.8_2. The mineral resource was classified to include all material within the saprolite supergene domain as this domain is flat lying and located entirely within the top 80m from surface.


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Table 14.8_1


Confidence Levels of Key Categorisation Criteria

Items Discussion Confidence

Drilling Techniques Diamond drilling is Industry standard approach. RC drilling has been wet but has minimal assays at the time of this estimate

High

Logging Standard nomenclature and apparent high quality. High Drill Sample Recovery Good recovery recorded. High

Sub-sampling Techniques & Sample Preparation

Sampling to industry standard for DC. RC wet samples will require twin holes to attain high confidence in the future

Moderate

Quality of Assay Data

Blank samples returned low level P2O5. The cause of this needs resolving. One standard has failed 2SD for 27% of assays. This needs resolving via additional Umpire sampling and analysis..

Moderate to low

Verification of Sampling and Assaying

No umpire samples taken to date. low

Location of Sampling Points Collar surveys need high accuracy surveying. (Currently hand held GPS) Topography requires higher accuracy as currently SRTM data.

Moderate to Low

Data Density and Distribution Approximately 200m spaced drilling which is somewhat sparse given the generally poor continuity of grade that is evident.

Low

Audits or Reviews Amazon Geoservices is unaware of external reviews. N/A Database Integrity No Material errors identified. High

Geological Interpretation Geological interpretation is excellent given the limited drill data to date. A high quality geological team is continually improving the 3D geological model

Moderate

Estimation and Modelling Techniques

IDW2 Moderate

Cutoff Grades Lower Cutoff Grade of 3% P2O5 defines the saprolite supergene domain well. High

Mining Factors or Assumptions

10mE by 10mN by 2.5mRL SMU. High


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The statement has been classified by Qualified Person Beau Nicholls (MAIG (CP)) in accordance with the Guidelines of National Instrument 43-101 and accompanying documents 43-101.F1 and 43-101.CP. It has an effective date of 31 July 2011. Mineral resources that are not mineral reserves do not have demonstrated economic viability. Amazon Geoservices and MBAC are not aware of any factors (environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors) that have materially affected the Mineral Resource Estimate. The Santana deposit is a greenfield site and therefore is not affected by any mining, metallurgical or infrastructure factors.

Table 14.8_2

Santana Deposit

Inferred Grade Tonnage Report Inverse Distance Weighted Power 2 (IDW2)

12.5E x 12.5mN x 3mRL

Lower Cutoff Grade (% P2O5)

Million Tonnes P2O5 %

CaO %

Fe2O3 %

Al2O3 %

SiO2 %

3 33.5 12.39 16.7 17.6 8.8 27.5 10 17.2 17.6 23.8 15.32 6.42 22.18 20 5.8 24.3 35.1 10.6 4.1 11.8

(1) The effective date of the Mineral Resource is July 31, 2011 (2) The Mineral Resource Estimate for the Santana deposit was constrained within lithological and grade based solids (Saprolite Supergene Domain) within the top 80m from surface. No optimisation studies have been applied to this simple flat lying mineralization. (3) Mineral Resources for the Santana deposit has been classified according to the “CIM Standards on Mineral Resources and Reserves: Definitions and Guidelines (December 2005) by Beau Nicholls (BSc (Geo) MAIG) an independent Qualified Person as defined by National Instrument 43-101.

In total, Amazon Geoservices has derived an Inferred Mineral Resource of 33.5. Mt with a grade of 12.39% P2O5, (3% P2O5 cut off applied)


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The following grade tonnage curve for Santana is shown below in Figure 14.8_1

Figure 14.8_1 Grade Tonnage Curve – Santana Phosphate Project


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15 MINERAL RESERVE ESTIMATES

No mineral reserves have been estimated for the Santana Phosphate Project.

16 MINING METHODS

16.1 Mineral Resource Block Model

NCL was provided with the 31 July 2011 mineral resource block model, developed by Amazon Geoservices as described in section 14.

The following corresponds to the list of variables contained in the received block model data:

Lithology code

Bulk density assigned by lithology type constant values (t/m3)

% P2O5

%CaO

CaO to P2O5 ratio

MBAC personnel provided the initial topography.

NCL did not audit the sampling data or the block model used for this project.

It is becoming common in the industry to develop mineral resource models (particularly where nonlinear estimation techniques are applied), which essentially take into account potential dilution within the blocks, or adopt selective mining unit (SMU) as part of the resource modelling process. NCL did not add additional dilution to the provided block model as it was considered to be already diluted. Furthermore, the relationship between the recommended mining fleet and the block model’s block size allows for selective mining when proper mining practice is followed.

The 31 July 2011 block model considers all material as inferred mineral resources. All optimization and mine planning work considers inferred mineral resources material as potential ore, depending on its rock type and economic attributes. No material can be converted to mineral reserves.

The reader is cautioned that the open pit mining study is a preliminary economic assessment that is preliminary in nature, includes inferred mineral resources that are considered too speculative geologically to have economic considerations applied and therefore be categorized as mineral reserves and there is no certainty that the preliminary assessment will be realized. No Mineral Reserves have been estimated. Mineral resources that are not mineral reserves do not demonstrate economic viability.


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16.2 Operating Parameters

A mine plan was developed for the Santana Phosphate Project to produce 300 k tonnes of phosphate concentrate at a grade of 34% P2O5 per year, with a peak total material movement rate of 4.2 M tonnes per year. This scenario results in the processing of approximately 1.5 M tonnes per year of ore with an average grade of 12.4% P2O5. The expected mine life for this project is 24 years. Given that the mineralization outcrops to surface, no major pre-stripping is scheduled. The mine is scheduled to work seven days per week or 365 days per year. Each day will consist of three 8-hour shifts. Four mining crews will cover the operation.

The study is based on operating the open pit mine with excavators of 3.5 cubic meter capacities for ore and excavators of 4.6 cubic meter capacities for waste, both with trucks of a capacity of 39 tonnes. This type of equipment is able to develop the require productivity to achieve a maximum annual total material movement of 4.2 M; and also to have good mining selectivity with the excavators as defined by the grade control activities.

It is becoming common in the industry to develop resource models (particularly where nonlinear estimation techniques are applied), which essentially take into account potential dilution within the blocks, or adopt selective mining unit (SMU) as part of the mineral resource modelling process. The Santana mineral resource model has been assessed to achieve this outcome and hence considered to be a ‘diluted’ model.

16.3 Parameters for Mine Design

16.3.1 Base Parameters

The table below summarizes the base case economic parameters used for the Lerch-Grossman economic shells analysis and mine design.

The mining cost estimate for the pit optimization process is based on NCL experience on similar projects developed in Brazil. The estimated average mining cost resulted in US$1.50 per tonne of mined material, as only a small proportion of material will require drilling and blasting (Table 16.3.1_1).

The product prices, processing costs and processing recoveries were provided to NCL by MBAC personnel, based on their own estimates and test work.


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Figure 16.3.1_1 Lerch-Grossman Optimization Parameters

Prices Super Simple Phosphate (SSP) 350 US$/tonne SSP (18% P2O5)

Recoveries by product

P2O5 55.0%

Concentrate Grades

P2O5 34.0%

Costs

Mining 1.50 US$/mined tonne

Plant 36.0 US$/tonne SSP

G&A 5.9 US$/tonne SSP

Others 9.2 US$/tonne SSP

Sulphuric Acid 55.3 US$/tonne SSP

Granulation 43.3 US$/tonne SSP

Total processing cost 149.7 US$/tonne SSP

Pit Slope angle 31° considers an allowance for ramping


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16.3.2 Lerch-Grossman Economic Shells Results

The following tables show the results of the optimization runs for the Santana Phosphate Project. Pit shells were generated for several revenue factors, from 0.44 to 1.0. The following figures show graphically those results. Inferred mineral resources are used, but they cannot be converted to reserves.

Table 16.3.2_1

Lerch-Grossman Optimization Results

Revenue Optimization Total Waste

Total Ore

Factor Price Material

Strip

Rec. P2O5

Pit

USD/Tonne SSP Ktonnes Ktonnes Ratio ktonnes % P2O5 Ktonnes

1 0.44 154 4,482 573 0.15 3,908 22.3 478

2 0.46 161 38,127 16,659 0.78 21,467 14.5 1,708

3 0.48 168 54,257 26,950 0.99 27,308 13.5 2,022

4 0.50 175 59,884 30,887 1.07 28,998 13.1 2,094

5 0.52 182 63,442 33,661 1.13 29,782 13.0 2,129

6 0.54 189 69,526 38,701 1.26 30,825 12.8 2,173

7 0.56 196 76,324 44,386 1.39 31,938 12.6 2,217

8 0.58 203 77,923 45,763 1.42 32,160 12.6 2,225

9 0.60 210 79,486 47,142 1.46 32,344 12.6 2,233

10 0.62 217 80,739 48,251 1.49 32,487 12.5 2,238

11 0.64 224 81,999 49,386 1.51 32,613 12.5 2,243

12 0.66 231 82,664 49,993 1.53 32,671 12.5 2,245

13 0.68 238 83,317 50,590 1.55 32,727 12.5 2,248

14 0.70 245 86,393 53,430 1.62 32,963 12.5 2,256

15 0.72 252 86,850 53,854 1.63 32,996 12.4 2,258

16 0.74 259 87,129 54,109 1.64 33,020 12.4 2,258

17 0.76 266 87,363 54,329 1.64 33,034 12.4 2,259

18 0.78 273 87,493 54,452 1.65 33,040 12.4 2,259

19 0.80 280 87,649 54,598 1.65 33,051 12.4 2,260

20 0.82 287 87,942 54,878 1.66 33,064 12.4 2,260

21 0.84 294 88,020 54,953 1.66 33,067 12.4 2,260

22 0.86 301 88,231 55,150 1.67 33,081 12.4 2,261

23 0.88 308 88,383 55,296 1.67 33,087 12.4 2,261

24 0.90 315 88,457 55,365 1.67 33,092 12.4 2,261

25 0.92 322 88,655 55,556 1.68 33,098 12.4 2,261

26 0.94 329 88,697 55,596 1.68 33,100 12.4 2,262

27 0.96 336 88,776 55,671 1.68 33,105 12.4 2,262

28 0.98 343 88,889 55,781 1.68 33,108 12.4 2,262

29 1.00 350 89,016 55,904 1.69 33,112 12.4 2,262


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Figure 16.3.2_1 Tonnes and Grades versus optimization price

Figure 16.3.2_2 Final pit shell – General Review

0.0

5.0

10.0

15.0

20.0

25.0

0

10

20

30

40

50

60

70

80

90

100

0.44

0.48

0.52

0.56

0.60

0.64

0.68

0.72

0.76

0.80

0.84

0.88

0.92

0.96

1.00

P2O

5 gr

ade

(%)

Mas

s (M

ton

nes)

Pit shell optimization price (USD/tonne SSP)

Total Material Total Ore %P2O5

N


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Physically analyzing the obtained shells, a single sector can be identified. The final pit shell is roughly tabular with a depth between 30 and 100m. The main axis has a NW orientation and a strike of approximately 1.9 km, the minor axis strike varies between 0.4 and 1.5 km.

It is important to note that the parameters are initial estimates, done at the beginning of the study, for the purpose of starting the design process.

16.4 Pit and Mine Design Phase

16.4.1 Final Pit Design

The final pit design was based on the economic shell generated at a revenue factor equal to 1.0 and a minimum cut-off grade of 3% P2O5, with constant 31.0° inter ramp angle. Table below shows the key open pit design parameters.

Table 16.4.1_1

Pit Design Parameters

Bench Height 12 m (four stacked 3m benches)

Berm Width 10 m

Batter Angle 50.0°

Inter-ramp Angle 31.0°

Ramp Width 15 m

Ramp Gradient 10%

Detailed Geotechnical information is not available, but these are very conservative parameters. MBAC’s Arraias Tocantins project located in Brazil has similar characteristics to the Santana Phosphate Project and uses more aggressive parameters. The adopted parameters are deemed adequate and conservative by NCL for the current stage of the Project.

The ramps width of 15m can accommodate up to 39 tonne trucks. NCL used a 10% road gradient, which is common in the industry for this type of trucks.

The current mine plan is designed with 3m benches stacked to 12m (i.e. four stacked benches).

For the required mining rate the appropriate loading equipment is medium size hydraulic excavators, ranging between 4 to 6 cubic yards capacities. The minimum operating widths to achieve the required productivities with this equipment is about 25m.

The Figure below shows the final pit, designed according to the obtained pit shell of the optimization process. The lower pit bottom is at 194 m RL. The total area disturbed by the pits is about 1.7 km2.


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Figure 16.4.1_1 Plan View Final Pit and Dump Designs

16.4.2 Tabulation of Pit Contained Resources

Table 16.4.2_1 summarizes the pit contained inferred mineral resources for the final design pit at several different P2O5 cut-off grades.

Table 16.4.2_1

Resources Contained in Final Pit at Various Cut-off Grades

Total Ore Waste TOTAL

Cut-Off (%P2O5)

Ktonnes (%) P2O5

Ktonnes Ktonnes Strip ratio

Total 33,045 12.4 61,274 94.319 1.9

Waste

61,274

>3% 33,045 12.4

>6% 26,307 14.4

>9% 19,287 16.9

>12% 13,712 19.5

>15% 9,929 21.8

N

East Dump

West Dump


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16.5 Mine Production Schedule

A mine production schedule was developed to show the ore tonnes, grades, total material and waste material by year throughout of the life of the mine. The distribution of ore and waste contained inside the final pit was used to develop the schedule, assuring that criteria such as continuous ore exposure, mining accessibility, and consistent material movements were met.

NCL used an in-house system developed to evaluate several potential production mine schedules. Required annual ore tonnes and user specified annual total material movements are provided to the algorithm, which then calculates the mine schedule. Several runs at various proposed total material movement schedules were done to determine a good production schedule strategy. It is important to note that this program is not a simulation package, but a tool for calculation of the mine schedule and haulage profiles for a given set of phases and constraints that must be set by the user.

The mine plan developed by NCL does not include any special provisions for dilution because the mineral resource block model is considered as already diluted.

Nevertheless, careful grade control must be carried out during mining to minimize misplaced ore due to the important effect of head grade on phosphate concentrate production. These efforts should include, among others, the following standard procedures:

Implement an intense and systematic program of sampling, mapping, laboratory analyses, and reporting.

Utilize specialized in-pit bench sampling drills for sampling well ahead of production.

Use of excavators and benches no higher than 3 m (as presently planned) to selectively mine ore zones.

Maintain proficient laboratory staff, equipment, and procedures to provide accurate and timely assay reporting.

Utilize trained geologists and technicians to work with excavator operators in identifying, marking, and selectively mining and dispatching ore and waste.

Maintain an adequate communication and control system between excavator and truck operators and the mine dispatcher to ensure that the materials are sent to the appropriate destination.

According to initial estimates for processing costs, metallurgical recovery and long term price for obtained products, the marginal cut-off which defines the separation between ore and waste is calculated as 3% P2O5.

Table 16.5_2 shows the mine production of ore for each mining year. The schedule is based on 1.5 M tonnes maximum per year to the processing plant and a production of 300 K tonnes of phosphate concentrate at 34% P2O5. The table also shows the total material movement from the mine by year, which peaks at 4.2 M tonnes per year during commercial production.

Show in Figure 16.4.1_1 are the East and West dumps where the waste mined will be stored. These dumps have enough capacity to store all the mine waste. Table 16.5_1 shows the waste dump design parameters.


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Table 16.5_1

Waste dump design parameters Lift height 5 m

Berm with 10 m

Batter angle 35°

Overall angle 16°

Waste density 1.3 t/m3

Given that the mineralization outcrops to surface, no substantial preproduction movements are expected.

The plant feed is also shown on Table 16.5_2. It shows that the plant feed is close to 1.5 M tonnes per year to produce constantly 300 K tonnes of concentrate. Given the favourable ore distribution there is no need for a long term stockpile.

Table 16.5_2

Mine Production Schedule

Period

Ore Waste

Total Strip Ratio

Plant Feed

Mass ktonnes

P2O5 (%)

Mass ktonnes

Mass ktonnes

Mass ktonnes

P2O5 (%)

Conc. Mass ktonnes

CaO (%)

CaO / P2O5 Ratio

Y01 1,500 12.4 2,608 4,108 1.7 1,500 12.4 300 10.7 0.86

Y02 1,489 12.5 2,628 4,117 1.8 1,489 12.5 300 14.3 1.15

Y03 1,484 12.5 2,611 4,096 1.8 1,484 12.5 300 11.0 0.88

Y04 1,481 12.5 2,562 4,043 1.7 1,481 12.5 300 13.5 1.08

Y05 1,472 12.6 2,565 4,037 1.7 1,472 12.6 300 16.3 1.29

Y06 1,510 12.3 2,603 4,113 1.7 1,510 12.3 300 17.9 1.46

Y07 1,480 12.5 2,626 4,105 1.8 1,480 12.5 300 17.4 1.39

Y08 1,474 12.6 2,636 4,110 1.8 1,474 12.6 300 24.7 1.96

Y09 1,500 12.4 2,614 4,114 1.7 1,500 12.4 300 21.2 1.71

Y10 1,496 12.4 2,639 4,135 1.8 1,496 12.4 300 21.2 1.71

Y11 1,480 12.5 2,577 4,057 1.7 1,480 12.5 300 17.3 1.38

Y12 1,504 12.3 2,589 4,093 1.7 1,504 12.3 300 17.5 1.42

Y13 1,438 12.9 2,619 4,058 1.8 1,438 12.9 300 17.6 1.37

Y14 1,451 12.8 2,647 4,098 1.8 1,451 12.8 300 17.5 1.37

Y15 1,423 13.0 2,701 4,124 1.9 1,423 13.0 300 18.7 1.43

Y16 1,450 12.8 2,655 4,104 1.8 1,450 12.8 300 15.2 1.19

Y17 1,426 13.0 2,696 4,122 1.9 1,426 13.0 300 15.8 1.21

Y18 1,443 12.9 2,680 4,123 1.9 1,443 12.9 300 17.1 1.33

Y19 1,446 12.8 2,647 4,093 1.8 1,446 12.8 300 16.7 1.30

Y20 1,467 12.6 2,562 4,029 1.7 1,467 12.6 300 16.9 1.33

Y21 1,512 12.3 2,623 4,136 1.7 1,512 12.3 300 18.6 1.52

Y22 875 9.8 3,165 4,040 3.6 875 9.8 139 12.9 1.32

Y23 1,061 8.5 2,743 3,804 2.6 1,061 8.5 145 9.8 1.16

Y24 183 9.8 278 461 1.5 183 9.8 29 9.9 1.01

Totals 33,045 12.4 61,274 94,319 1.9 33,045 12.4 6,612 16.6 1.34


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16.6 Mine Equipment

The study is based on operating the Santana mine with backhoe excavators each of 3.5 cubic meters capacity for ore and 4.6 cubic meters capacity for waste and 40 tonne conventional haul trucks. The current situation at mines in Brazil is that contractors are using this type of equipment at lower costs for material movement up several times the schedule mining rate.

Auxiliary equipment includes bulldozers (type CatD9T), motor graders (type Cat160M) and water truck (type 20,000 litres).

Mine equipment requirements were calculated based on the annual mine production schedule, the mine work schedule, and equipment annual production capacity estimates, and maintenance downtime.

Table 16.6_1 provides a summary of the peak number of units required for commercial production. This represents the equipment necessary to perform the following duties:

Construct haul and access roads to the initial mining areas as well as to the crusher, waste storage areas, and leach pads. Construct additional roads as needed to support mining activity

Perform the pre-production development required to expose ore for initial production

Develop new mining areas for ore extraction

Mine and transport ore to the crusher. Mine and transport waste material from the pit to the appropriate storage areas

Maintain all the mine work areas, in-pit haul roads, external haul roads, and maintain the waste storage areas

Load and transport topsoil to topsoil storage areas

Table 16.6_1

Peak Fleet Requirements for Commercial Production

Equipment Type: Required Additional

Fleet Fleet Atlascopco Explorac R50 1 0

Volvo EC460B (3.5 cu m) 1 1

Volvo EC700B (4.6 cu m) 1 0

Scania Truck (39mt) 5 0

CATD9T Track Dozer 1 1

CAT160M Grader 1 1

Scania Water Truck (20000 liter) 1 1

Additional equipment, beyond what is needed to fulfill production requirements, was deemed necessary to ensure reliable production. These are an extra EC460B excavator, CATD9T track dozer, CAT160M grader and a Scania Water Truck. It was considered that when these equipments are out of service due to planned or unplanned maintenance for anything more than short periods of time the effect on the mine is unacceptable.

This equipment was included in the capital estimation but no extra costs were included in the operational expenses.


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16.7 Mine Personnel

16.7.1 Salaried Staff

Mine salaried staff requirements over the project life are shown in Table 16. The staff consists of 26 persons through the life of mine. The salaried staff requirements are commensurate with the complexity of a mine the size of Santana.

Annual costs for the personnel, including fringe benefits, are also shown on Table 16.7.1_1.

The personnel costs used for this project were provided by MBAC personnel and correspond to actual Brazilian costs obtained from bench marking of other operations.


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Table 16.7.1_1

Salaried Staff Requirements

Annual

Cost ($US)

Y01 Y02 Y03 Y04 Y05 Y06 Y07 Y08 Y09 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17 Y18 Y19 Y20 Y21 Y22 Y23 Y24

MINE OPERATIONS:

Mining Engineer 65,000 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Supervisor - Operations 47,109 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 1

Mine Clerk 13,424 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Mine Operations Total 5 5 5 5 5 5 5 5 5 5 5 5 5 5 5 5 5 5 5 5 5 5 5 3

MINE MAINTENANCE

Maintenance Manager 188,059 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Mechanical Engineer 124,079 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Maintenance Supervisor 67,299 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 1

Planner Supervisor 67,299 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Maintenance Clerk 13,424 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 1 1

Mine Maint. Total 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 7 5

MINE ENGINEERING:

Senior Mine Engineer 124,079 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Chief Surveyor 40,319 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Surveyor Helper 13,424 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Dispatch Operator 48,759 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 1 1 1

Mine Engineering Total 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 4 4 4

MINE GEOLOGY:

Mine Geologist 56,046 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Junior Geologist 40,319 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Technician/Sampler 40,319 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 1

Mine Geology Total 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 3

TOTAL PERSONNEL 26 26 26 26 26 26 26 26 26 26 26 26 26 26 26 26 26 26 26 26 26 23 22 15

TOTAL COST US$

x1000 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,475 1,328 1,315 965

Annual Cost includes Fringes Benefits


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16.7.2 Hourly Labor

Mine total hourly personnel requirements are shown in Table 16.7.2_1. The maximum number of persons during commercial production is 69 in Year 21. Year 24 personnel requirements reduce significantly due to reduction of pit operations.

Table 16.7.2_1 also shows the annual cost for hourly personnel, including fringe benefits. The personnel costs used for this project correspond to actual Brazilian costs used by NCL during 2010 for similar projects. Exchange rate 1.60 R$/US$ was considered.

As shown on Table 16.7.2_1, the majority of persons in mine operations are equipment operators. The number of operators for major equipment was calculated based on equipment operating requirements.

Additional mine persons are assigned to perform the following tasks:

Pumpman. Set pumps and pipe as required and use small excavators for building sumps.

General labourer. The general labourer is an unskilled worker who assists with many of the mine support activities such as moving and setting pumps, road maintenance, general mine cleanup, etc.

Table 16.7.2_1 also shows the number of maintenance personnel required for each time period. It can be seen that the ratio of maintenance personnel to operations personnel is about 60% during the mine life. This is above industrial average, and is caused because of the low amount of equipment on site and the minimum number of personnel required servicing a piece of equipment.

An additional allowance in the manpower is required to cover vacations, sick leave, and absenteeism (VS&A). The 10% VS&A allowance is based on 30 vacation days plus 6 sick days out of 365 scheduled days per person per year.


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Table 16.7.2_1

Hourly Labour Requirement

Annual

Cost ($US)

Y01 Y02 Y03 Y04 Y05 Y06 Y07 Y08 Y09 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17 Y18 Y19 Y20 Y21 Y22 Y23 Y24

MINE OPERATIONS

Grade Control Drilling 35,949 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Loading 35,949 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 1

Hauling 35,949 15 16 16 18 19 19 17 19 20 18 18 19 20 19 19 18 18 19 20 19 21 19 20 3

Dozer 35,949 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 3

Grader 26,812 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 2

Grade Control 19,502 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Water truck 26,812 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 2

Laborer 19,502 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3

Pump man 19,502 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Total Operators 34 35 35 37 38 38 36 38 39 37 37 38 39 38 38 37 37 38 39 38 40 38 39 17

MINE MAINTENANCE

Mechanic 35,949 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 4 4

Mechanic's helper 19,502 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 2 2

Welder 35,949 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 2 2

Electrician 35,949 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 3 3

Fuel Man 19,502 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 3 3

Laborer 19,502 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 2 2

Maintenance Total

23 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23 16 16

Hourly Labor Cost US$x1000 1,604 1,640 1,640 1,712 1,748 1,748 1,676 1,748 1,784 1,712 1,712 1,748 1,784 1,748 1,748 1,712 1,712 1,748 1,784 1,748 1,819 1,748 1,581 845

VS&A Allowance 10% 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6 3

VS&A Cost US$x1000 184 185 185 186 186 186 185 186 187 186 186 186 187 186 186 186 186 186 187 186 187 186 188 87

Total Labor requirement 63 64 64 66 67 67 65 67 68 66 66 67 68 67 67 66 66 67 68 67 69 67 61 36

Total hourly labor cost US$x1000 1,788 1,824 1,824 1,897 1,934 1,934 1,861 1,934 1,970 1,897 1,897 1,934 1,970 1,934 1,934 1,897 1,897 1,934 1,970 1,934 2,006 1,934 1,769 931

Maint./Operations Ratio 68% 66% 66% 62% 61% 61% 64% 61% 59% 62% 62% 61% 59% 61% 61% 62% 62% 61% 59% 61% 58% 61% 41% 94%

Vacations, Sick leave and Absenteeism (VS&A) is based on 30 vacation days and 6 sick days/person out of 365 scheduled days/person per year. Annual Cost includes Fringes Benefits


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17 RECOVERY METHODS

17.1 Production Plants

17.1.1 Mine Site

The mineral processing plant was designed based on the metallurgical test work and considering the following design parameters:

Beneficiation Plant

Total ROM feed to the plant: 1.5 million tpy

Feed grade Ore: @ 12.5% P2O5

Production of concentrate: 300,000 tpy @ 34% P2O5 (dry basis)

Metallurgical recovery from flotation: 55%

Total plant availability: 89%

The process adopted consists of the following steps, as shown in the figure below:

Primary crushing of the ROM material

Classification of the ore into two size fractions

Grinding of the coarse product to 80% passing 100 microns

Flotation of the grinded product

Regrind of the rougher concentrate to 80% passing 44 microns

Flotation of the regrinded rougher concentrate

High intensity magnetic separation of the cleaner concentrate

Concentrate dewatering

All the tailings are combined and pumped to the tailings dam. See figure below.


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Figure 17.1.1_1 Process Flow Sheet

Due to the friable characteristics of the ROM material, the maximum lump size will be limited to 600 mm approximately; as a consequence, the crushing section will be simpler.

Process Summary

The ROM ore is discharged directly into a 160m3 feed bin, prior to screening on a 150 mm opening vibrating screen.

The oversize material is directed to the crusher; the undersize product, combined with the material from the crusher, is discharged on a single belt conveyor.

A second belt conveyor with a stacker will deliver the rock to two homogenization stockpiles of 30,000 m3 capacity, each.

This will provide an operation autonomy of about 12 days.

A reclaimer will be used to feed the ore from the stockpiles to the Beneficiation plant, through a belt conveyor equipped with a weight scale which records the throughput rate to the grinding mill.

The material is then discharged onto a three deck vibrating screen.


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The undersize material (less than 6.35 mm) is directed to an attrition scrubber fitted with two agitators; the pulp overflow is directed by gravity into a slurry hopper and then, pumped to a pre-classification hydro cyclone cluster.

The combined material from the three screen decks (over 6.35 mm), is directed a belt conveyor which will feed the SAG mill.

The SAG mill discharge material, passes through a screening trommel, with an oversize material (> 6.35 mm) recirculation circuit.

The trommel undersize material is discharged into the SAG mill hopper; where is combined with the underflow material from the pre classification hydro cyclones. Additional process water is added to the hopper; the material is finally pumped to the grinding circuit classification hydro cyclone cluster.

The underflow of this circuit is directed to the SAG mill inlet.

The overflow of the pre classification hydro cyclones is discharged via gravity into another slurry hopper and then pumped to a bank of slimes dewatering hydro cyclones.

The overflow of the fines dewatering hydro cyclones is them directed to a hopper and pumped to the tailings pond whilst the under flow is returned to SAG mill discharge hopper.

The overflow from the grinding circuit hydro cyclones is directed to a slurry hopper and then pumped to two conditioning tanks and in sequence, by gravity, to a slurry hopper.

This material is them pumped to a conventional flotation circuit employing two roughing flotation cells in series.

Flotation

The concentrate rock is then reground and floated again. And the tailings are directed to the tailings pond.

The cleaner flotation concentrate will be directed, to a slurry tank and will be pumped to the magnetic separation circuit.

Magnetic Separation

The magnet material from the magnetic separator reports via gravity to the tailings dam. The non-magnetic material then flows to a slurry tank via gravity and is then pumped to the main concentrate thickener.

The overflow from the main concentrate thickener flows by gravity to a tank and is combined with the natural fines thickener overflow and filtrate water. Filtrate water from the main

Thickening and Filtration


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concentrate filter system is then combined into a single tank and pumped to the process water system tank. The main thickener underflow is pumped to a conditioning tank fitted with a mechanical agitator and then pumped to a conventional press filter. A belt conveyor deliver the filter cake to the phosphate rock concentrate stockpile.

The output of the filtration section is a concentrate rock with a 15% average. moisture content. This material will be directed to a stockpile dimensioned for 10,000 t.

Rock Stock pile & Expedition

This material will be reclaimed by a FEL and sent to the Industrial site utilizing 27 tonne trucks. See flow chart in the figure below.

Figure 17.1.1_2

Beneficiation Flow Chart

The tailings area will be positioned at the north of the mine deposit taking advantage of a natural slop. See figure below.

Water Accumulation and Tailings disposal

The purpose of this reservoir is to receive the waste coming from the refining plant, while accumulating water available during the rainy season, thus compensating the eventual lack of water during the dry season.


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The dam project, will allow a convenient solids flow setting distribution, thus promoting the adequate water recirculation.

As reference, the location of the dam axis is S 9° 18’ 4,1” and W 51°47’54”.

Figure 17.1.1_3

Dam Project

The advantage of geographic tailings disposal in this area are:

Adequate distance from the Beneficiation plant (4,0 km)

Adequate reference level related to Beneficiation Unit

Contribution of small water streams

Easy access

Low volume of compacted massif

A preliminary evaluation indicates that three raisings are foreseen for the Dam project.

The crest of the initial Dam will be at 266m elevation, corresponding to a deposit volume of 12.0 million m3.

The final Dam crest elevation will be at 272m corresponding to a final volume of 29.0 million m3. This will allow about 24 years of useful life for the reservoir.


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The final volume of the compacted massif is estimated to be 220,000 m3.

The reservoir inundation area will be around 720 ha (final); the grassland vegetation is dominant but some natural flora will be suppressed; for this reason, a specific study will be needed.

The difference level between the refining plant and the reservoir inlet will be of about 6.0m which indicates the possible need of a buster to avoid waste deposit in the feeding pipe.

17.1.2 Industrial Site

Rock received from the mine site is stored in an open area with an estimated capacity of 15,000t.

Rock Concentrate Stockpile

This raw material will be transported to the Industrial site, to feed the Acidulation Unit.

The sulphuric acid plant will have capacity to produce 750 tpd. The technology adopted for the study is the MECS double absorption route.

Sulphuric Acid Plant

Process phases are as follows:

Sulphur melting and filter

Sulphur burning in reaction with air

SO2 conversion into SO3 in catalytic beds using Vanadium Pentoxid catalyst

SO3 gases scrubber and absorption in appropriate tower

Cooling and storage of acid product

The plant will be continuously monitored to assure both the emission of SOx and the sulphur mist meet the requirements of the applicable law.

In the sulphuric acid production process several chemical reactions occur: the sulphur combustion in a special Burner; SO2 conversion into SO3 (inside the converter), and SO3 absorption (in the Acid Towers). See flow chart below.

The oxidation and absorption steps in the manufacture of sulphuric acid from sulphur are all highly exothermic. The excess heat generated at each step of the process is recovered in the waste heat boiler, superheater, and economizers. The recovered heat is in the form of high pressure superheated steam which will be used to generate electric power in a turbo-generator. The process is designed to give a conversion of sulphur dioxide to sulphuric acid of

Process Summary


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99.7% in the acid plant as well as a high conversion of process heat to steam. The only normal plant effluent streams are the blowdown streams from the boiler and stack.

The principal steps in the process consist of burning sulphur (S) to form sulphur dioxide (SO2), combining the sulphur dioxide with oxygen (O2) to form sulphur trioxide (SO3), and combining the sulphur trioxide with water (H2O) to form a solution containing sulphuric acid (H2SO4).

Atmospheric air is drawn through an air filter and air silencer and is compressed in the main compressor which provides the motive force to move the gases through the downstream equipment. The gas leaving the main compressor goes to the dry tower where the water vapors in the air are removed by contacting a stream of 98% H2SO4 which flows down through the tower. From the dry tower the gas enters a horizontal spray-type sulphur furnace where the temperature of the SO2 gas from the sulphur furnace is higher than is required for inlet to the conversion system. Therefore, the gas is cooled in a waste heat boiler which recovers the surplus heat as high pressure saturated steam. The gas temperature from the boiler is controlled by a gas-side bypass. The boiler steam temperature is a function of the boiler steam pressure.

From the waste heat boiler, the gas flows to the first pass of the converter system where it is partially converted to sulphur trioxide gas in the presence of vanadium catalyst. The conversion reaction produces heat. The gases must be cooled to improve the yield of the sulphur dioxide oxidation in the next catalyst pass. Gases leaving the first converter pass flow to a superheater where they are cooled by heating the export steam. The temperature of the export steam and pass 2 inlet gas temperature is controlled in the proper range by bypassing a portion of the steam flow around the superheater. The cool gas stream flows from the superheater to the second converter pass where additional conversion of the suphur dioxide to sulphur trioxide takes place accompanied by the generation of additional heat. Hot gases leaving the second converter pass are cooled to improve the yield in the next catalyst pass by sending them through the hot interpass gas heat exchanger. Exit gas temperature is controlled by a bypass.

Cooled gases leaving the heat exchanger flow to the third converter pass where additional conversion of sulphur dioxide to sulphur trioxide takes place. Hot gases leaving the third converter pass are cooled by sending them through the cold interpass heat exchanger and the economizer 3B.

In the interpass tower the SO3 is removed from the gas stream by contacting the gas with circulating 98-99% sulphuric acid.

From the cold interpass heat exchanger, the gas stream flows to the hot interpass heat exchanger where it is further heated by hot gases leaving the second converter.

From the hot interpass heat exchanger, the gas stream flows to the fourth converter pass where final conversion of SO2 to SO3 is accomplished. The temperature to the fourth


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converter pass is controlled by bypassing a portion of gas around the cold and hot interpass heat exchangers.

The gas stream leaving the fourth pass enters the split water flow economizer 4A/4C where it is cooled by boiler feedwater. Water side bypasses are used to control the exit temperature from economizers 4A and 3B. The exit gas temperatures can be regulated to prevent economizer drip acid formation normally associated with variable hydrocarbon content in sulphur feed.

Gas leaving the economizer 4A/4C enters the final absorbing tower prior to exhausting to the atmosphere through a stack.

In the final absorbing tower, SO3 in the gas stream reacts with water in the 98-99% circulating acid. The temperature of the strong acid circulated over the final absorbing tower increases due to the heat of formation and the sensible heat of the gas stream entering the tower. The acid from the bottom of the final absorbing tower flows into the pump tank.

The 98% acid produced in the absorbing tower is pumped through a product acid cooler and then to a storage, tank.

The designed Sulphuric Acid plant will generate 33t/h of superheated steam @ 40 kg/cm2 which will feed the turbo generator.


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Figure 17.1.2_1 Sulphuric Acid Flow Sheet

The Sulphuric Acid plant will export superheated steam at 40 kg/cm2 pressure and 473ºC temperature which will feed the Co-generation unit which is designed to produce a total of 8.0 MWh of electric power @ 13.8 Kv.

Co-generation

The turbo-generator unit will generate electrical energy in a condensing turbine. A small portion of low pressure steam @ 7.0 Kg/cm2 will be extracted from the turbine casing. This low pressure steam at will be used for the sulphur melting and Granulation Unit. The balance of the steam flow will be condensed at the exit of the turbo-generator and returned as make-up to the deaerator.

Acidulation Plant

The Acidulation plant is designed to produce 500 Mtpy of SSP. This process will utilize the slurry route for the sulphur acid attack.

Process Summary

The phosphate rock from the stockpile will be reclaimed by a FEL to a belt conveyor to feed the re-pulping tank.

After being re-pulped, the slurry will be pumped to the intensive mixer where sulphuric acid (H2SO4) will be introduced. The reactional slurry product from the intensive mixer will direct to the curing belt (DEN).


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After 10 to 15 minutes residence time, the reactional slurry will solidify and it will pass through a lump breaker before being transferred to the cure warehouse by a belt conveyor.

During the chemical reaction in the DEN, silicon tetrafluoride (SiF4) will be released. The gases with fluorine content will be transferred to the absorption system where it will transform into fluosilicic acid (H2SiF6) and hydrated silica (SiO2.2H2O) by counter current contact with water.

The concentration of this suspension will be 18% to 20% H2SiF6 and it will be withdrawn from the gas treatment system to the surge tank.

The scrubbers will be fitted with sealing tanks and liquor circulation pumps. Reposition water will be fed onto the final stage tank and a chimney will release the gases to the atmosphere after being treated according to relevant environmental standards.

Subsequently, the suspension with 3% to 6% solid content will be pumped to the press filter to remove the hydrated silica. The silica cake will combine with the acidulate material from the DEN and both will be transferred to the cure warehouse. The filtered fluorosilicic acid will be collected in the reservoir tank and then to shipment and/or neutralization. See flow chart below.

Figure 17.1.2_2

Acidulation Plant Flow Sheet


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Granulation

The granulation plant will be able to produce up to 500.000 tpy of SSP.

Process Summary

The single superphosphate powder (SSP) from the cure warehouse will be transferred trough a belt conveyor, to the granulator drum where water and slurry from the gas treatment system will be introduced. The gases produced in the granulator will direct to the gas treatment system while the product will direct to the rotary dryer.

Heat from the combustion chamber which burns firewood will be used in the dryer. The ashes from the furnace will be disposed accordingly.

The gas generated when granulated SSP passes through the dryer will be treated in the gas treatment system consisting of a pneumatic cyclones and venturi scrubber before being released to the atmosphere.

When drying operation is completed, the granulated SSP will report to the hot screening. The oversize from the hot screening will direct to the cold screening after passing through the rotary cooler. Gases produced in the cooler will report to a bag filter to remove the particulate material, which will return to the process. Conveyors will withdraw the undersize from the hot screening and feed the recycle bin as circulating load.

The cold screening screens will be fitted with two decks. The oversize from the upper decks will feed a chain mill and the product will be delivered to the recycle bin. The undersize from the upper decks will combine with the oversize from the lower decks and will be stockpile in the granular product pile.

The undersize from the lower decks will combine with the hot screening undersize and chain mill product in the recycle bin. The material from the bin will be the circulating load to feed the granulator.

All dust collecting points located along the granulation plant will report to the bag filter. Then the clean air will be released to the environment after passing through an exhaust fan and chimney. Conveyors will transport the powder retained in the filtering process to the recycle bin. See flow chart below.


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Figure 17.1.2_3

Granulation Flow Sheet


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18 PROJECT INFRASTRUCTURE

18.1 Background

Santana Phosphate Project is located in the South-eastern region of Pará State, in the Northern Brazil. The DNPM concession area for mineral evaluation, is of about 86.0 km2 located in a farmland deforested area, showing a savannah aspect.

The Santana inferred mineral resource estimate is based on 53 diamond holes (2,386m) and 6 RC holes (238m) drilled at a spacing of approximately 200m by 200m. MBAC has an ongoing infill and extensional drillhole program underway.

An independent inferred mineral resource has been estimated comprising 33.5Mt with an average P2O5 content of 12.39% (using a preferred 3% P2O5 Cut-off).

The stripping ratio is estimated to be 1.8:1.

Bench scale metallurgical test work on rock samples collected from outcrops in the Project area, Indicated that a high grade (34% P2O5) phosphate concentrate with a good reactivity, can be generated utilizing a simple beneficiation process (MBAC).

The mineral deposit is strategically located, considering the growing fertilizer market in the project region of influence (Mato Grosso and Pará State).

Those facts encouraged the Company to develop a Project to produce low concentration fertilizers, mainly oriented to soya bean crops.

18.2 Project Scope

18.2.1 Production Units

The Santana Phosphate Project takes into account the favorable geographic position of the mine allied to the good rock characteristics, resulting in a lower production and logistics costs.

The Project aims the implementation of the following production Units:

Ore Beneficiation Plant - 300 kty

Sulphuric Acid Unit - 245 kty

Acidulation Unit (SSP) - 500 kty

Granulation Unit (GSSP) - 500 kty

Cogeneration Unit - 8.0 MWh

Utilities and Auxiliary Plants

Water & Tailings Reservoir


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Due to the non-existent social infrastructure in the proximity of the Mine area, in addition to the difficulty of access, the best approach for the whole Project will be to locate the production facilities in two sites as shown in the figure below.

18.2.2 Mine Site

It will contemplate the operation of the open pit Phosphate mine; the rock Concentration Unit; and the Water & Tailings deposit.


This site will be dedicated to the Chemical Units and should be located close to BR 158, at approximately 20 km from Santana do Araguaia.

Figure 18.2.3_1 Industrial Site

18.3 Project Location Characteristics

18.3.1 Mine Site

The Mine is located in the Southern of Pará State in the Brazil northern region at an elevation of 309m and geodesical coordinates: S 9° 40’.23” W 51° 47’.23”.

The area is located at 229 km west from Santana do Araguaia (near Araguaia river) and at about 25 km from the Xingu river, on the east side.

The topography is relatively flat with gentle slopes.


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The climate is tropical with a well defined dry season (May – October), followed by a wet summer.

Typical rainfall is of 1800 mm during the summer period and temperature may vary from 20 to 36°C.

The original forest vegetation was drastically reduced and substituted by grassland for cattle pasture; as a consequence, the hydrography of the region was seriously affected by the reduction of water streams.

Water for the Project will come from Capivara river located 15 km from the site.

Figure 18.3.1_1 Mine Site


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This site was positioned 200 km from the mine, in a flat area of 200 ha (existing farm), not far from Santana do Araguaia (20 km) and close to the BR 158. The position was defined, considering the logistic aspects related to raw materials input and product distribution to Mato Grosso and Pará state. See figure below.

Figure 18.3.2_1 Industrial Site


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18.4 Access

18.4.1 Road Access

Pará State road structure is very poor currently; this creates some logistic difficulties for this Project; however, this situation will improved in the coming years, with the implementation of PAC (Government Infrastructure Program) which contemplates the paving of BR 158 by the end of 2012.

This important highway will cross the south east of Mato Grosso up to the north, reaching the city of Redenção in Pará state. See figure below.

Figure 18.4.1_1 Location Plan

Currently, access to Mine site is very difficult even utilizing the shorter route. Starting from Santana do Araguaia (PA), you must travel 94 km on paved road (BR 158) up to Vila Mandi, and them the Mine site is approximately 141 km drive on an unpaved road, mainly used by farmers to transport cattle.

Mine Site

This is a country route, not adequate for heavy trucks due to its limited width, and 40 small, primitive wood bridges.

In the dry season, the distance from the Mine up to Vila Mandi, can be covered by a small truck in about 5 hours. However is difficult to estimate travelling time during the rainy season.

Considering the importance of this route for the Project, it is recommended that MBAC considers an investment of about R$ 12.0 million to update these road from the Mine Site up


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to Vila Mandi, to support the traffic of 27 tonne trucks to transport concentrate to the Industrial site.

In addition, a negotiation with farmers will be needed for their permission to improve and upgrade this road.

It must be emphasized however, that this situation can be radically improved in the future if the Government implements the road BR 235 scheduled for the five years plan.

Analyzing the road map from DNIT (Figure 18.4.1_1) we can notice that this new road BR 235 is planned to cross the Pará State from east (Caximbo) to West (Santa Maria das Barreiras) connecting the BR 153 (Xingu basin) to BR 158 (Araguaia basin).

The route of the concentrate from the Mine Site will be on the existing unpaved 141 km to Vila Mandi and then 97 km over paved road to the Industrial Site, for a total 238 km. A closed loop truck transport shall be utilized for this service.

Industrial Site

The access to and from the Industrial site will be straightforward, utilizing the BR158 which is an important road going from south to north Brazil connecting Mato Grosso State to Pará State, up to the city of Redenção, and other alternative routes linking to the Araguaia river water way.

This road still has some unpaved sections but work is underway and is expected to be completed by the end of 2013 (PAC program). It will be an important logistic alternative for the Project.

18.4.2 Waterway Access

The waterway flow utilizing the Araguaia–Tocantins river system is considered in the government plan (PAC 2). It will be part of an integrated intermodal platform (Marabá) to improve the logistic approach to the center region of Brazil, with a total extension of 2,794 km.

This route will be an important link between the Vila do Conde port (Belém) in the northern of Pará to the east-center of the country, allowing the flow of raw materials and products for export (grains), at a lower transportation cost.

Work is underway with an estimated completion date by the end of 2013. This will be an alternative route, which will reduce the inland freight cost for the imported sulphur by about 20%. (Figure 18.4.2_1).


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Figure 18.4.2_1 Water Access

18.5 Infrastructure Aspects

18.5.1 Mine Site

Livestock is the economic driver for this region (Pará State), having the highest cattle concentration in Brazil. The Project region is completely occupied by cattle farms with no social infrastructure.

The closest village (80 km) is Garimpinho which has only 200 inhabitants and absolutely no social infrastructure.

Vila Mandi (141 km) with approximately 1,000 inhabitants has a very poor infrastructure. It is the place where the farmers meet for cattle negotiation. Vila Mandi is likely too far from the mine and cannot be considered as a support for human resources and services.

Based on the above considerations, there is very little availability of human resource with basic technical knowledge. As well no professional schools were identified in this first survey. As a recommendation, a plan to create manpower for the Project should be considered, utilizing professionals to set up a “training program” for the operation of each specific plant.

Human Resources


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Mine site will utilize 160 workers including direct and indirect people.

Housing

This will require the building of accommodations, including a basic infrastructure to enable a normal living.

This site will be built in the outer limits of the plant.


The Industrial site will be located near Santana do Araguaia which will provide the adequate social support. However, the service infrastructure is poor specifically for technical services.

MBAC intends to implement training facilities in Santana do Araguaia in agreement with the local authorities to develop a Training Program for both sites.

18.6 Industrial Utilities

18.6.1 Electrical Power Networks

Due to the enormous distances required to be covered, the integration of high tension power transmission in Brazil is not complete specifically in the State of Amazonas and Pará. In addition, the high tension power distribution in Pará is not well balanced with a concentration in the north and eastern portions of the state.

As a consequence, 40 thermoelectric power stations were installed to provide power to small cities (under 15,000 inhabitants).

Within the Mine Area, only a low tension distribution line for farmer’s utilization exists.

The nearest available high tension point is located in Miracema, (Tocantins state) approximately 370 km from the Project site. CELPA (Centrais Elétricas do Pará) is the power distribution company, which receives energy from Tucurui, a hydroelectric power Company which generates 8,370 MWh and is located in the north of Pará State.

However, the distribution network (1,840 km) is mostly oriented to the northern part of Pará and Maranhão State.


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Figure 18.6.1_1 High Tension Electricity

As seen in Figure 18.6.1_1 the center and southeast of Pará state, as well as the north of Mato Grosso, are not presently included in the high tension distribution grid. This fact brings a new challenge to Santana Phosphate Project, for both sites.

Due to its geographical location the operation will be directly affected by the situation mentioned above.

Mine Site

The Mine site is about 370 km from the closest high tension facility located in the city of Miracema.

The electric power consumption in this facility is estimated to be 5.0 MWh.

A possible solution will be to negotiate with CELPA the launching of a new high tension. power line (138 Kv) from Santana do Araguaia up to the Mine site.

CELPA belongs to GRUPO REDE which also controls CELTIN in Tocantins State, which is working with MBAC on the Itafos-Arraias Project.


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Initial discussions between MBAC and GRUPO REDE were held in order to develop a project which will contemplate to set up a new 138 KV line coming from Miracema (TO) up to Santana do Araguaia (215 km) and then, continuing up to the Mine Site (184 km).

The implementation of this project will have an enormous positive social impact, enabling the development of the south east region of Pará State (Figure 18.6.1_2).

Figure 18.6.1_2 Electrical Power

The investment needed to implement 184 km of power line from Santana do Araguaia to the Mine site should be assumed by MBAC and will be compensated in ten years by a reduction of project power bills. This investment has included in the CAPEX estimate.

The solution mentioned above will also be adequate for this facility due to its convenient location very close to the new power line.

Industrial Site

The total estimated electric consumption for this facility is 12.5 MWh being a part of this cogenerated internally, though the Sulphuric Acid Operation (8,0 MW).


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18.6.2 Water Supply

The total estimated water needed for this site will be 1,100 m3/h. A major part of this, will be recirculated from the water and tailings reservoir. However due to the high evaporation index, it will be necessary a make-up of 360 m3/h of clear water.

Mine Site

The Project will take this new water from Capivara river wich is at approximately 15 km from the Beneficiation Plant.

Due to the difference level of 103m between the Plant and the river, a compensation water tank was considered.

This investment related to fresh water impoundment was included in CAPEX estimates.

A drilling survey to evaluate the potential for ground water will be undertaken in the near future

With regards to process water, the chemical plants will demand approximately 260m3/h which will be provided by an intake station located in a river, close to the site and 9 km aqueduct to the site.

Industrial Site

In this case, the difference of elevation between the two points in less than 15m.

A pond for raw water is included in the Project.

In addition, the company will undertaken an underground water survey will take place.

18.7 Logistic Aspects

18.7.1 Raw Materials and Products

The poor access from the Industrial Mine, to the Industrial Site creates a challenge that may be solved in the coming years if the construction of a new federal road (BR 235) is completed as planned linking the east to the west side of the Pará State. This road will have an extension of 522 km and will go from the Araguaia River, an important water way up to BR 080, near to Xingu River passing at about 40 Km from the Mine Site.

Concentrate Rock Feed to Industrial Site

An investment of R$ 12 million was included in the Capex to support this action.

It was also taken in account as an addition in the OPEX due to the related increment in the rock concentrate transportation cost.


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This material is imported and utilized to produce Sulphuric Acid. The annual need for this Project will be of 82.2 ktpy.

Sulphur

The present logistic plan is to receive this input via Itaqui port in Maranhão State, taking advantage of the synergy between the two plants (Itafos and Santana), thus reducing the sea freight cost.

Sulphur will travel 850 km from Itaqui to the Santana Industrial Site by Carajás railway up to Parauapebas (Pará State), followed by truck, through roads PA 275 and PA 150 and finally road BR 158 to the Industrial facility, covering a total of 1,267 km at an estimated inland freight of US$ 100.00/t.

Another alternative would be to use the port of Vila do Conde (Belem-PA) and the Araguaia Tocantins system waterway up to Santa Maria das Barreiras 980 km then by road BR 158 (84 km) up to the Chemical Plants.

The total inland freight cost, utilizing this alternative should be certainly lower. (Figure 18.7.1_1).

Figure 18.7.1_1 Freight Route


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Wood chips or “brickets” to be utilized in the drying process of the Granulation Plant are available in the Pará State.

Fuel & Biomass

The products from Santana will be mainly delivered to customers located in southern Pará and the northeastern region of Mato Grosso which will account for approximately 90% of the total production.

Final Products

The flow of finished products will be easy, utilizing mainly the road BR 158 which will be totally paved up to the end of 2012.

This will enable MBAC to provide products up to 500 km radius, reaching the central region of Mato Grosso in a highly competitive basis.


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18.8 Project Implementation Schedule

Figure 18.8_1

Project Implementation Schedule


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19 MARKET STUDIES AND CONTRACTS

19.1 Brazilian Market Overview

19.1.1 Agribusiness

Brazil is one of the most important players in the international trade of agricultural commodities, with a leading position for the production and export of soybean, corn, sugar, coffee and orange juice, other than chicken and beef.

For this reason, it is being considered the most important source of grains to support the enormous world food demand increase, foreseen for the next decades. It must be emphasized also that Brazil has one of the world’s largest agricultural area, sufficient to support the growing global demand for food and biofluel.

The forecast soybean export for 2011 will follow the tendency of 5.2% yearly increase, and is estimated to be 32.0 million tonnes. This volume represents approximately 47% of the country production for this year.

Table 19.1.1_1

Soybean export evolution

Year 2005 2006 2007 2008 2009 2009 2010

Volume (000 t)

22,400 24,900 23,700 24,500 28,100 29,000 32,000

The importance of agribusiness in Brazilian economy has consolidated the farming activities as the main driver for country development.

A new and important agriculture frontier is represented by the gradual substitution of the cattle pasture by soybean plantations on the northern east side of Mato Grosso State.

In this scenario, Mato Grosso will consolidate its importance in the development of Brazil agribusiness.

19.1.2 Fertilizer

Brazil is not independent of fertilizers imports; currently, almost 67% of nutrients are imported. Besides, the continuous growth of agriculture (~5.0% py) will tend to increase the need of fertilizers imports, even considering the new fertilizer plants to be implemented in the coming years.

According to the National Association for the Promotion of Fertilizers (“ANDA”), in 2010, Brazil produced 9.3 million tonnes of fertilizers and has imported 15.0 million tonnes.


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Figure 19.1.2_1 Brazilian P2O5 Market Scenario

Also according to the ANDA, the consumption of fertilizers is distributed fairly evenly in the three regions of Brazil. The North Center of the country consumes 33% of the supply of NPK, the South 29% and the Southeast 28%. The Northeast region accounts for 9% of the consumption of fertilizers.

Despite the even distribution between the areas North Center, Southeast and South, the consumption of NPK is differentiated.

The state of Mato Grosso is responsible for 16% of national consumption of NPK and for 48% of the consumption in the North Center area of the country. Goiás consumes 9% of the national supply and 28% of this area. Mato Grosso do Sul absorbs 5% of the total and 15% of the supply in the region.

The total consumption by Southeast area, São Paulo absorbs 52% and Minas Gerais 43%. At the national total the same states participate with 15% and 12%, respectively.

In the South of the country, the State of Paraná accounts for 50% of the regional consumption, Rio Grande do Sul by 40% and 10% in Santa Catarina. In the Brazilian demand they account for 15%, 12% and 3%, respectively.

In the Northeast, the State of Bahia stands with consumption of 68% of the region. In the national demand, the consumption of this state is 6% of the total.


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In the last 23 years, Brazil has recorded a yearly increase 5.2% in fertilizer production; the evaluation of domestic consumption of nutrients is demonstrated in the figure below.

Figure 19.1.2_2

Domestic Nutrients Consumption


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19.2 Region of Project Influence

Santana Phosphate Project is strategically located in relation to the target area for the Project taking in account the development of the new agricultural frontier. Project Target region is: (i) the northern east of Mato Grosso, (ii) south Pará; and (ii) west limits of Tocantins.

Figure 19.2_1

Project Area Immediate Market

This market will represent in 2015, as a conservative approach, an additional of 150 ktpy of P2O5, being at least 110 ktpy as single superphosphate or 550 ktpy of product.

In addition, we are also considering the potential market of the east-central of MT. Where, there will be a clear competitive advantage of Santana CIF product due to its lower freight costs, if compared with product coming from Goias or even from Santos/Paranaguá.

The distance covered by products coming from the South east region is about 2,000km and for those coming from Goias is 1,500km, with an estimate freight cost of US$ 90.00/t.

As an average, we estimate a CIF freight cost advantage of US$ 40.00/t for the Santana products.

Considering soybean production the existing logistic structure will not support the estimated volumes forecasted for the next 5 years; besides, consider that in the period of January to July 2011, the fertilizers imports through Itaqui port reached 345 kt, being 118% higher if compared with the same period of 2010.


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In the near future a convenient alternative for the products directed to export, will be the utilization of Araguaia-Tocantins system then enabling to direct this flow to Itaqui, by rail, taking advantage of the new TERGRAN site planned to absorb 5.0 million tons of grains; the first phase of this important logistic investment will be operational by the end of 2013.

19.3 Competitors

Currently, there is no alternative to P2O5 required at the target area, resulting the requirement of import fertilizer from other domestic regions or abroad.

Figure 19.3_1 Competitors

The mainly supplement of the P2O5 consumed at the target region came from (i) Port of Itaqui (imported); (ii) Port of Santos (imported); (iii) Port of Paranaguá (Imported); and (iv) southern Minas Gerais and Goias states (Domestic), where there are some competitors.

The principal product route is via Port of Itaqui, because it is the closest port of the region and domestic producers are more competitive in other domestic regions, for example, south eastern Brazilian region.


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19.4 Project Santana Competitive Advantage

Santana Phosphate Project has a significant competitive advantage resulting from strategic location.

The distance covered by products coming from the South east region of Brazil (Port of Paranaguá and Port of Santos) is about 2,000km; for those coming from Goias and Minas Gerais states is 1,000km; and 1,500 km from Itaqui Port.

Those large distances results in a large freight costs comparing to Santana Phosphate Project product, giving the project a large competitive advantage.

Figure below shows Santana Phosphate Project location and the location of their main competitors.

Figure 19.4_1

Competitors Distances to Santana Phosphate Project Area


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20 ENVIROMENTAL STUDIES, PERMITTINGS AND SOCIAL OR COMMUNITY IMPACT

20.1 Environmental Requirements

The environmental requirements for a project implementation at the Para state consists of acquiring three series of Licenses: Preliminary License, Installation License and Operating License.

An Environmental Impact Study (EIS) with a detailed Operations Plan is required by the Environmental State Agency from the Pará state for the Preliminary License issuance. The Operations Plan consists of the basic mine plan, infrastructure, and operational aspects of the Project. The Environmental Study consists of a comprehensive study on the Plan of Operations for historical concerns, environmental and cultural issues and impacts. An Impact Report (RIMA) is also required and it consists of a document that summarizes the content of the Environmental Study and it should be written in a colloquial language in order to be easily understood by the population, especially those affected by the project.

A Public Hearing, in accordance with Federal CONAMA 009 / 1987, is also mandatory for the Preliminary License issuance. The Hearing has the objective to show the direct and indirect influences of the project to all the people living in the area. At the Public Hearing there will be opportunity for the company to answer all questions from the population, and collect comments and suggestions about the project. The company pursuing the project has the responsibility to facilitate the location and transportation to the proposed site.

The Installation License is mandatory before any construction and mine operations can begin, and it can be obtained shortly after the issuance of the Preliminary License (LP), given that any condition or adjustment requested upon LP issuance has been performed.

The last step on the licensing process would be the issuance of the Operating License, and it can only be obtained on completion of the construction, at the plant commissioning, when all the project aspects listed on the LP and LI can be checked by the Environmental Agency.

At the current project stage, MBAC has hired Geomma consulting to perform an initial Environmental assessment at the mine area, where the largest impacts are expected. It is not possible at this stage to evaluate or quantify the impacts, but the summary of the findings is shown in the next paragraphs. MBAC shall also start the preparation of a complete Environmental Impact Study shortly.


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20.2 Protected Areas

Protected areas are those areas considered restricted, where there are limitations to the full exploitation of the mineral industry. There are two of them around the project: Indigenous Lands and Badjonkore e Menkragnoti, which are located at a distance of about 11.6 km and 16 km northwest and west of the ADA respectively. At this distance there will be no impact for the project and preliminary there cannot be forecasted any related issues that should be considered in the project relating to the impacts of the project upon these areas.

20.3 Physical Environment

20.3.1 Climate and Weather

The municipality of Sao Felix do Xingu belongs to the southeastern region of Pará and the micro-region of Sao Felix do Xingu. The relative humidity values recorded average of 70% during the year.

The region is characterized by having a hot sub-tropical climate having a mean annual temperature around 30.5 º C. The rainfall is as high as 1370.3 mm / year, somewhat irregular during the year.

20.3.2 Soils

The area is home to two major types of soils: Acrisol + dystrophic Oxisol and Udorthent dystrophic and eutrophic.

The clay soil + dystrophic red-yellow sandy clay loam soil, which represents 70% of the project area, has a deep soil, porous and strophic, where there may be conditions for root growth in depth, otherwise it will take place featuring dystrophic or limitation of a chemical in depth, especially if the phosphorus content is low. The Oxisol soil has limited porosity and low nutritional potential, and it is probable a problem for compression. This type of soil is related to the occurrence of the acidic granitic rocks of the Xingu Complex.

20.3.3 Hydrography

The mesh covers the region's watershed tributaries of the Xingu River basin, which covers the grid coordinates Latitude 08 ° 30'0 "to 09 ° 35'0" S Longitude 52 ° to 30'0 "to 52 ° 0'0" W, and west of the Igarapé do Jacaré and east river da Paz and Capivara, within the municipality of São Felix do Xingu.

The project area is located in the basin between the rivers Rio da Paz and Capivara, both tributaries of the right bank of the Xingu River,Both riversare born in the high mountain ranges that divide the municipalities of San Felix, Cumaru do Norte. The river Capivara springs about 130 km in a northwesterly direction until it reaches the Xingu River, where the project is located, on its right bank tributaries. The Rio da Paz has its trajectory until it flows into the Xingu River near 120 km in the northwest, and the applicant's areas near the springs.


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The Xingu River in Mato Grosso is born in the junction of the rivers Seven and Culuene September, which, in turn, have their origin in the Serra do Roncador. Its main tributaries are the rivers Culuene, Curisevo, or Tamitatoala Batovi, Ronuro, Manissauá-Miçu, Iriri Acaraí and Jaraucu, the left margin, and Seven of September, or Suiá-Miçu Suiazão, Commander Fontoura, Fresh and Bacajá, the bank right.

The Xingu River basin covers the states of Mato Grosso and Para, being bounded to the south and east by the basin of the river Tocantins / Araguaia andto the west by southwest of the Tapajós River and the Paraguay River basin. It is situated between the parallels 15° and 1° south latitude and meridians 50° and 56° west longitude.

His contribution basin has a total area of approximately 509,000 km2 and has an elongated shape, with maximum width of 550 m, running in the south-north direction before flowing into the Amazon River, representing 7.8% of the basin, which covers about 6.1 million square kilometers, ranking among the largest in the Brazilian Amazon.

20.3.4 Hydrology

There are few remaining perennial streams within the area surrounding the project, including those presenting only a small volume of water. The study shows the devastation of the vegetation for use as pasture land, a fact that is now common in the Amazon region and causing changes in local water resources. Among human activities, those that most stand out are cattle ranching and timber extraction for commercialization.

The region is part of the municipality of Sao Felix do Xingu, which was strongly influenced by human actions, with damage to the vegetation types by deforestation, to carry out activities as: grazing, mining and logging, among others with consequences to the flora and fauna that resulted in the gradual loss of biodiversity.

One has to consider that the actual colonization process in southeast Pará imposed a known use on natural ecosystems, well beyond the carrying capacity of these systems, causing a collapse of these structural and functional, leading to succession types, highly impoverished in refers to biological and genetic resources.

The modified environments dominate the local scene, including large extensions of land, as the lower areas of flooded forests (riparian), and its altered systems, characterized by strong presence of more generalist species of greatest plasticity and environmental requirements.


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Figure 20.3.4_1 Farms in the project area

20.3.5 Flora

According to the text of Resolution CONAMA No. 10, 01.10.93 secondary or succession vegetation is the result of natural processes of succession, after the canceling of all or part of the primary vegetation by anthropogenic or natural causes..

Currently, it is known that the floristic composition of secondary forests, known as capoeira, is usually compounded predominantly of pioneer species, little shade tolerant and adapted to low fertility conditions or low demand in relation to nutrients. The plants are extremely aggressive when it comes to physical occupation of space, allowing the formation of a less hostile environment and suitable for larger species that require better shade to settle and reproduce.

The vertical structure of vegetation is considered high, with an average height of 8.31m, with values reaching up to 10 or 30m. The landscape vegetation type, the discontinuous canopy is uniform, with trees of different sizes, ranging from small to medium. There was sort of economic use of timber as Eschweilera ovata (kill kill), Mezilaurus Itauba (itaúba) and Socrates exorrhiza (paxiúba) as remaining species of primary education. It is also found individuals of the genus Cecropia (trumpet trees), which denote human action, in which the withdrawal of the original vegetation has given way to ecological succession to species of fast growth and good adaptability to environments considerably degraded.

Regardless of its stage, virtually all forest fragments present some degree of degradation, mainly at the border, with an imbalance of populations of lianas and invasion of grasses such as Panicum maximum, Brachiaria sp. and Mellini minutiflora. This goes so far as to prevent the evolution of forest succession and may decrease the supply of food and shelter for wildlife. Another aggravating factor is that in these situations, the fire can spread more easily, as the mentioned grasses are highly flammable. The occurrence of fire and hunting are the major threats to these forests, where biodiversity is being lost each year gradually, as shown in Figure below.


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Figure 20.3.5_1 Area Comprised of small individuals

The area looks much degraded in some places, mostly due to edge effects and fire, but it was classified as late stage because a portion of the area is still preserved and also because, legally, it does not lose this feature. In most parts of the fragment the tallest trees remain, reaching 25-30 feet high, highlighting the Jequitibá pink (Cariniana legally), Peroba-rosa (Aspidosperma polyneuron) and Jatoba (Hymenaeae courbaril). Forming the forest canopy, between 15 and 18 meters, the highlights are white Jequitibá (Cariniana estrellensis) and Cedro-rosa (Cedrela fissilis).

Figure 20.3.5_2

Vegetation Composed by pioneer species


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Figure 20.3.5_3 Palm species identified in the study area

20.3.6 Pasture Area

Livestock is the main use for land in southeast Pará, therefore, the area is predominantly of pasture use, abandoned or fallow.

Figure 20.3.6_1 Pasture Area

In pasture areas, the main forage grass identified is the grass Brachiaria sp., and, eventually, isolated individual trees are found. In some abandoned or fallow areas, the formations of individual trees of small dense tangles congregate where there is a predominance of a few tree species. Both environments have a significant non-timber income. There is the occurrence also of species of palm trees scattered throughout the pasture with or without the association of tree species.


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Figure 20.3.6_2 Cleaning Areas of pastures

Periodically, "dirty" pastures, invaded by weeds, (locally known as "Juquirá") and species of secondary succession, are "cleaned." To clean them ranchers use fire, whose control and extension are not always well managed.

Figure 20.3.6_2

Streams in Permanent Preservation Areas

20.3.7 Fauna

Among terrestrial vertebrates, birds, mammals and reptiles are the groups that contribute to a more efficient characterization of environmental conditions in a particular area, because, besides being quite diverse in their habits and ecological requirements, they are generally active throughout the year and can be recorded by direct methods (contact) or indirect (tracks, feces, shelters) with relative safety.

The inventory carried out has obtained 21 species of mammals, 26 birds and 12 reptiles, 59 species, in total number of surveys.


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The group of birds is the richest in species and represents 44% of the fauna identified, followed by mammals that correspond to 36% and 20% reptiles.

20.3.8 Mammals

In the surveyed area, 23 related mammal species, represented in 12 families and seven orders, were found. Among the mammals there is great diversity of habits and therefore the standards body that require the application of various methods for the determination of ecological parameters (VOSS, Emmons, 1996).

20.3.9 Avifauna

In the study area were diagnosed 26 bird species, classified in 10 orders and 17 families. These taxa correspond to the populations that remain in different environments fragmented forest birds as well as some more requirements for environmental conditions.

In this light, one can consider the birds as an indicator of the presence of wildlife in the area due to the ease of observation of this segment and its better adaptation to changing environments and human occupation.

Figure 20.3.9_1 Avifauna


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20.3.10 Herpetofauna (Reptiles and Amphibians)

Considering the methodology applied in the study, based on records obtained from field observations and interviews with residents, it were diagnosed 14 species, including two species of frogs grouped in two families of reptiles and 12 species distributed in three orders and eight families.

Representatives of the herpetofauna play an important role in ecosystems due to their position in food webs (second-order consumers), controlling populations of vertebrates and invertebrates, especially land, and constitute the food source for numerous species of wildlife.

As a result of deforestation, species of reptiles and frogs of open areas have expanded geographically its limits, at the expense of more demanding species. The forest reptiles are more difficult to visualize due to their habits, while the snakes are found more randomly, and their records based primarily on interviews with the community.

Amphibians were not specific targets of this diagnosis, but its occurrence may be recorded in the area across species.

20.3.11 Biota

Vegetation with a worrying stage of conservation was identified in the area, containing invasive species and presenting low potential for the timber sector, thus, found in some sampling sites, a characteristic of secondary vegetation, from anthropogenic factors.

The field survey data indicate the occurrence of 61 species of terrestrial vertebrates in 26 birds, 21 mammals, 12 reptiles and 2 amphibians, although these preliminary figures represent a slowdown in wealth compared to other areas already studied in the region .

It can be said that the project area is home to a wildlife species characteristic of disturbed habitats, while still allowing the occurrence of some species more demanding as the demand for environmental and ecological resources.

The main causes of the reduced number of species should be related to fragmentation of the site, with destruction of the original forest and replacement with areas of scrub and anthropized environments, with a decreasing trend in the presence of wildlife species and individuals.

It is difficult to precisely measure the level of threat facing the emerging knowledge of the distribution, habitat use and biology, requiring further studies, not only because of the variety of environments found, but mainly to the degree of environmental characterization.

However, one can conclude that the studied area was dramatically changed with great loss of biodiversity, mainly due to widespread deforestation, burning, and livestock.


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20.3.12 Anthropic - Social Indicators

According to the Secretary of State for Planning, Budget and Finance (Sepof) in 2000, the number of people living in the municipality of Sao Felix do Xingu was 34,621, with 19,087 men and 15,534 women. In 2007, that number rose significantly, totaling 59,238 inhabitants (31,993 men and 26,828 women). In surveys conducted in 2010, the population of Sao Felix do Xingu rose to 1.08 inhabitants / km², which means a total of 90,908 inhabitants.

This study does not indicate what possible changes that the implementation of the project could lead to population growth of Sao Felix do Xingu, as the municipality is part of one of the areas destined to the mine project and other mineral projects and may thus be constituted in an attractive hub for migrating population, sprawl, violence, disease, among others.

20.3.13 HDI

The Human Development Index (HDI) and Index Firjan Municipal Development (IFDM) were raised to support public policies in order to implement improvements in the quality of life of local people.

The HDI was created by the United Nations (UN) and it aims to analyze data on life expectancy, education and income.

According to the UNDP, the figures for 2007 and published in 2009, Brazil ranked 75th in the HDI ranking. In the balance held in 182 countries, Brazil had a score of 0.813, remaining in the group of developing countries considered to be high, ie, HDI greater than 0.800.

To make up the HDI, it is customary to use economic and social data ranging from zero (no human development) to 1 (total human development). The closer to 1, more developed the country, region, or municipality.

Figure 20.3.13_1 Health in São Félix do Xingu


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20.4 Infrastructure

20.4.1 Electricity

The federal government, through the Luz no Campo Program has enabled the population living in rural areas to have access to electricity, thus increasing the consumption of the municipality.

The distribution of electricity is the responsibility of the concessionaire Celpa Network, however, it fails to meet the needs of the entire population, because in several parts of cities the illegal capture of energy can be observed.

20.4.2 Sanitation

Considering the supply of water, sewerage and garbage collection, the municipality of Sao Felix do Xingu has the worst regional results.

The water supply system does not have a network of efficient distribution of treated water, causing the locals to do the uptake of water in an irregular manner, usually through wells totally precarious and without any quality.

Some homes have artesian well or the abstraction is done via specialized dealer. Another method of water collection, usually held in areas furthest from the urban core, is through the opening of wells such as Amazon.

The system of garbage collection in the municipality studied, serves only urban center of the core, not meeting the demand of the population that grows more and more.

20.4.3 Trade

The commercial center of Sao Felix do Xingu presents very diverse businesses.

A trade more targeted to meet the demand for agricultural products helps to characterize the region of Sao Felix as a potential rancher area.

The city has clothing stores, appliance and electronics, and gas stations, network banking, pharmacies, etc. (Figures below)


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Figure 20.4.3_1 Commercial Streets in São Félix do Xingu

Figure 20.4.3_2 Comercial Shops in São Félix do Xingu

Figure 20.4.3_3 Streets in São Félix do Xingu


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Figure 20.4.3_4 Roads in São Félix do Xingu

20.4.4 Public Safety

According to map the violence of Brazilian cities, Sao Felix do Xingu is in the Para region with the highest rates of homicide, where most deaths are caused by conflicts related to land tenure.

As for the agrarian structure, the study indicates that the municipality of Sao Felix do Xingu are peasant units, farms and estates business, unregulated land tenure, aggravating problems with squatters and squatting.

Among the most serious crimes committed in the county, according to the military police of the municipality, are: those linked to land conflicts, armed robbery, theft of animals, vehicles and motorcycles.

20.4.5 Anthropic considerations

The enterprise and its structure will be deployed in an area devoid of communities and villages.

Surveys of primary and secondary data demonstrated the serious shortage of places in the city and areas of influence of the project. Sao Felix do Xingu indicates serious weaknesses in the services and utilities.

Within this perspective, despite the emergence of new economic activities often cause damage to the environment, on the other hand are responsible for stimulating the economy and consequently the improvement of life of local people by increasing employment and income.


6005 Australia

The council will appoint a strong tendency for agriculture, with the economy, especially employment and income by turning in large part to this sector.

In health and education, the limited number of establishments and professionals is one of the most serious problems faced by local people. The infrastructure is also quite fragile, with respect to the availability of sanitation, housing, electricity, trade, etc.

The locations shown total lack of resources in all sectors surveyed, i.e., they are totally dependent on the services of neighboring municipalities.

Therefore, the lifting of economic and demographic dynamics, as well as the use and occupation of a municipality or a community is critical to achieve positive results on the negative impacts that may result in a new activity.


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21 CAPITAL AND OPERATING COSTS

21.1 Capital Costs

21.1.1 Mine Capital Costs

The estimated mine capital cost includes the following items:

Major mine equipment

Mine support equipment

Shop tools

Initial spare parts

Engineering and geology equipment

This estimate does not include the following mine physical structures:

Fuel and lubricant storage facilities

Explosive storage facilities

The mine shop, offices, and warehouse

It is anticipated that the vendors will provide storage for fuel, lubricants, and explosives as part of their contract of work and that the prices for these items are included in the delivered price. The mine shop and warehouse are included in the infrastructure capital cost prepared by ATS Projetos for this study.

Table 21.1.1_1 summarizes the mine capital costs by category for the initial period plus all sustaining capital needed for the Santana Phosphate Project. Initial mine equipment capital is US$ 13.1 million. Sustaining capital, which is required for mine fleet increase and replacement, amounts to US$ 23.3 million. Total capital during the life of the project is US$ 36.4 million.

Table 21.1.1_2 shows the equipment unit prices used for this study. The following is noted:

Base equipment prices are shown in constant 2nd quarter 2010 US. It is assumed that payment for the equipment is made at the time of delivery

The costs for major equipment are based on quotes obtained by NCL for this project during the second quarter of 2010

The capital costs shown include delivery to the site and assembly

Exchange rate of 1.60 Brazilian Real per every US dollar was considered for mine equipment, when applicable


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Contingency is not included in the mine capital cost. It is possible that final negotiated sales prices, with fleet discounts, will be somewhat lower than the budget quotes used for this study

Table 21.1.1_1

Summary of Mine Capital Costs (US$’000)

Category Initial

Capital Sustaining

Capital Total

Capital

MINE EQUIPMENT:

Major Equipment 9,553 18,790 28,343

Support Equipment 1,333 3,799 5,133

Engineering/Safety Equipment 1,850 - 1,850

Shop Tools 96 5 100

Spare Parts 278 720 998

Subtotal Equipment 13,109 23,314 36,423

TOTAL 13,109 23,314 36,423

Table 21.2.1_2 Equipment Unit Costs (US$’000)

Mine Major Equipment: Lifetime Capital Cost Spare partshr KUS$ KUS$

Atlascopco Explorac R50 30,000 1,382 28 Volvo EC460B (3.5 cu m) 30,000 661 13 Volvo L220 F (4.4 cu. m) 25,000 752 15 Volvo EC700B (4.6 cu m) 30,000 1,079 22 Scânia Truck (39mt) 30,000 456 24 CATD9T Track Dozer 35,000 1,704 34 CAT160M Grader 35,000 681 14 Scania Water Truck (20000 l i ter) 24,000 278 6

0 0Mine Support Equipment: Capital Cost Spare parts

KUS$ KUS$Flatbed Truck (7.3 mt) 102 2 Mechanics Truck (4x4) 114 2 Scania Fuel Truck (12000 l i ter) 90 2 Backhoe Loader (1 cu m) 142 3 Pickup Truck (4x4) 31 1 Light Plants 27 1 Mine Radios 1 - Water Pipe - (per 1000m) 128 - Mine Pumps 48 -

EQUIPMENT UNIT COST ('000)


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21.1.2 Mine & Industrial Sites Capital Costs

Project Santana Capital Costs have been estimated to August 2011 rates base, which was based on Project Arraias Capital Cost. Total capital costs for the Project are estimated to be US$385 million, including US$73 million in contingencies. Table below show all Capital Cost breakdown.

The accuracy assumed for the CAPEX is ±15%.

B$/US$ exchange rate considered was 1.60.

Table 21.4.2_1

Capital Cost Breakdown


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21.2 Operating Costs (OPEX)

21.2.1 Mine Operating Costs

Mine operating costs were developed from the recommended equipment requirements presented in section 0 and the personnel requirements presented in section 0. The mine operating costs include all the parts, supplies, and labor costs associated with mine supervision, operation, and maintenance. The table below summarizes the total mine operating costs. Total cost, unit cost per total tonne of material, and unit cost per ore tonne are shown.

Table 21.2.5_1

Summary of Mine Operating Costs (US$’000)

Category Total Cost Cost Per Cost Per (US$'000) Total Tonne Ore Tonne

Commercial production 179,173 1.90 5.42

Given that the ore outcrops to surface, no material movement is required before the start of commercial production.

Total mine operating cost during commercial production is US$ 179.2 million. This amounts to US$ 1.90 per total tonne of material and US$ 5.42 per tonne of ore during this period.

Operating hours were calculated from estimated productivities for the selected mine equipment. Hourly costs were also estimated for the mine equipment, summarized in Table 21.2.1_2.

Table 21.2.1_1 Hourly Cost per Equipment

The following factors are considered for the operating cost calculations:

Local unit costs for consumable items such as diesel fuel and blasting agents

Total Wear Items Others ReparisUS$/hr lt/hr US$/hr lt/hr US$/hr hrs/set US$/h US$/h US$/h US$/h

Atlascopco Explorac R50 257.57 68.75 60.74 10.99 - 133.71 2.16 49.98 Volvo EC460B (3.5 cu m) 64.78 35.13 31.03 3.59 6.29 - 7.93 - 19.52 Volvo L220 F (4.4 cu. m) 83.68 25.00 22.09 1.75 3.06 51,540 14.32 1.52 - 42.69 Volvo EC700B (4.6 cu m) 113.60 53.75 47.49 4.94 8.64 - - 12.94 - 44.53 Scânia Truck (39mt) 37.64 18.75 16.57 0.11 0.19 17,496 4.37 - - 16.51 CATD9T Track Dozer 160.64 62.50 55.22 1.73 3.02 - 19.56 26.67 56.16 CAT160M Grader 68.64 32.13 28.38 0.74 1.30 30,000 7.50 4.15 - 27.32 Scania Water Truck (20000 l i ter) 27.92 15.00 13.25 2.00 3.50 6,020 5.02 - - 6.15

Flatbed Truck (7.3 mt) 75.85 62.50 55.22 0.79 1.38 - 3.75 3.00 - 12.50 Mechanics Truck (4x4) 19.05 18.75 16.57 0.37 0.65 - 0.59 - 0.63 0.63 Scania Fuel Truck (12000 l i ter) 19.05 18.75 16.57 0.37 0.65 - 0.59 - 0.63 0.63 Backhoe Loader (1 cu m) 25.09 10.00 8.84 1.00 1.75 - 2.50 - - 12.00 Pickup Truck (4x4) 7.12 6.25 5.52 0.20 0.35 - 0.50 - - 0.75 Light Plants 6.32 6.25 5.52 0.10 0.18 - - - - 0.63

HOURLY COST PER EQUIPMENT

Equipment Type: Diesel Lube-Oil-Filters- Tyres


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Local hourly labor rates and fringe benefits were used

Local costs for tires and spare parts were used

Constant exchange rate of 1.60 Brazilian Real per US dollar

The general activities that are included in the operating cost estimate are as follows:

Construct the initial out-of-pit mine access roads from the pit area to the ore crusher and waste storage areas

Removal and storage of the topsoil within the pit area

Preproduction development required to expose ore for initial production

Mine and transport ore to the crusher area. Mine and transport waste material from the pit to the waste storage areas

Maintain all the mine work areas, in-pit haul roads, and external haul roads. Also maintain the waste storage areas


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Table 21.2.1_3

Summary of Mine Operating Costs - Total Dollars

Mining Year

Total Material (ktonne)

Grade Control Loading Hauling Auxiliary Hourly

Labor General

Mine General Maint. G&A TOTAL

Cost/ Tonne

of Total Material

Cost/ Ore

Tonne

Y01 4,108 69 980 876 1,119 1,748 749 51 1,948 7,540 1.835 5.026 Y02 4,117 68 984 952 1,119 1,784 1,078 51 1,659 7,695 1.869 5.167 Y03 4,096 68 978 950 1,119 1,784 971 51 1,659 7,580 1.851 5.107 Y04 4,043 68 964 1,034 1,119 1,855 1,083 51 1,660 7,834 1.938 5.290 Y05 4,037 68 963 1,088 1,119 1,891 1,094 50 1,661 7,934 1.965 5.389 Y06 4,113 69 980 1,086 1,119 1,891 734 51 1,661 7,592 1.846 5.028 Y07 4,105 68 982 993 1,119 1,819 844 51 1,660 7,536 1.836 5.093 Y08 4,110 68 984 1,088 1,119 1,891 1,122 51 1,661 7,984 1.943 5.418 Y09 4,114 69 982 1,143 1,119 1,927 841 51 1,661 7,793 1.894 5.196 Y10 4,135 69 988 1,051 1,119 1,855 904 52 1,660 7,698 1.862 5.146 Y11 4,057 68 968 1,042 1,119 1,855 1,144 51 1,660 7,908 1.949 5.342 Y12 4,093 69 975 1,107 1,119 1,891 699 51 1,661 7,573 1.850 5.034 Y13 4,058 66 973 1,154 1,119 1,927 1,028 51 1,661 7,980 1.967 5.548 Y14 4,098 67 983 1,090 1,119 1,891 985 51 1,661 7,847 1.915 5.408 Y15 4,124 65 994 1,117 1,119 1,891 896 52 1,661 7,795 1.890 5.477 Y16 4,104 67 985 1,029 1,119 1,855 653 51 1,660 7,419 1.808 5.118 Y17 4,122 65 993 1,060 1,119 1,855 653 52 1,660 7,457 1.809 5.231 Y18 4,123 66 991 1,086 1,119 1,891 653 52 1,661 7,519 1.824 5.211 Y19 4,093 66 983 1,167 1,119 1,927 653 51 1,661 7,627 1.863 5.275 Y20 4,029 67 961 1,125 1,119 1,891 652 50 1,661 7,527 1.868 5.130 Y21 4,136 69 986 1,231 1,119 1,963 653 52 1,662 7,735 1.870 5.115 Y22 4,040 40 1,005 1,077 1,119 1,891 652 51 1,514 7,350 1.819 8.403 Y23 3,804 49 950 1,183 1,119 1,725 651 48 1,503 7,227 1.900 6.812 Y24 461 8 108 186 559 952 153 6 1,052 3,024 6.557 16.503 Total 94,319 1,517 22,640 24,917 26,292 43,856 19,545 1,179 39,228 179,173 1.900 5.422

Cost/ Tonne

of Total Mat'l

0.016 0.240 0.264 0.279 0.465 0.207 0.013 0.416 1.900

Percent

1% 13% 14% 15% 24% 11% 1% 22% 100%

Commercial Production - Total Cost, Unit Cost Per Total Tonne, and Unit Cost Per Ore Tonne 179,173 1.900 5.422


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Table 21.2.1_4

Summary of Mine Operating Costs – US$ per tonne

Mining Year

Total Material (ktonne)

Grade Control Loading Hauling Auxiliary Hourly

Labor General

Mine General Maint. G&A TOTAL

Total Cost $US 1000

Y01 4,108 0.017 0.239 0.213 0.272 0.425 0.182 0.013 0.474 1.835 7,540

Y02 4,117 0.017 0.239 0.231 0.272 0.433 0.262 0.013 0.403 1.869 7,695

Y03 4,096 0.017 0.239 0.232 0.273 0.435 0.237 0.013 0.405 1.851 7,580

Y04 4,043 0.017 0.238 0.256 0.277 0.459 0.268 0.013 0.411 1.938 7,834

Y05 4,037 0.017 0.239 0.269 0.277 0.468 0.271 0.013 0.411 1.965 7,934

Y06 4,113 0.017 0.238 0.264 0.272 0.460 0.179 0.013 0.404 1.846 7,592

Y07 4,105 0.017 0.239 0.242 0.273 0.443 0.206 0.013 0.404 1.836 7,536

Y08 4,110 0.016 0.239 0.265 0.272 0.460 0.273 0.013 0.404 1.943 7,984

Y09 4,114 0.017 0.239 0.278 0.272 0.468 0.204 0.013 0.404 1.894 7,793

Y10 4,135 0.017 0.239 0.254 0.271 0.449 0.219 0.013 0.402 1.862 7,698

Y11 4,057 0.017 0.239 0.257 0.276 0.457 0.282 0.013 0.409 1.949 7,908

Y12 4,093 0.017 0.238 0.271 0.273 0.462 0.171 0.013 0.406 1.850 7,573

Y13 4,058 0.016 0.240 0.284 0.276 0.475 0.253 0.013 0.409 1.966 7,980

Y14 4,098 0.016 0.240 0.266 0.273 0.462 0.240 0.013 0.405 1.915 7,847

Y15 4,124 0.016 0.241 0.271 0.271 0.459 0.217 0.013 0.403 1.890 7,795

Y16 4,104 0.016 0.240 0.251 0.273 0.452 0.159 0.013 0.405 1.808 7,419

Y17 4,122 0.016 0.241 0.257 0.271 0.450 0.158 0.013 0.403 1.809 7,457

Y18 4,123 0.016 0.240 0.263 0.271 0.459 0.158 0.013 0.403 1.824 7,519

Y19 4,093 0.016 0.240 0.285 0.273 0.471 0.159 0.013 0.406 1.863 7,627

Y20 4,029 0.017 0.239 0.279 0.278 0.469 0.162 0.013 0.412 1.868 7,527

Y21 4,136 0.017 0.238 0.298 0.271 0.475 0.158 0.013 0.402 1.870 7,735

Y22 4,040 0.010 0.249 0.267 0.277 0.468 0.161 0.013 0.375 1.819 7,350

Y23 3,804 0.013 0.250 0.311 0.294 0.453 0.171 0.013 0.395 1.900 7,227

Y24 461 0.018 0.234 0.403 1.213 2.065 0.331 0.013 2.280 6.560 3,024

Total 94,319 0.016 0.240 0.264 0.279 0.465 0.207 0.013 0.416 1.900 179,173

Percent

1% 10% 11% 12% 20% 9% 1% 18% 100.0%

21.2.2 Mine & Industrial Sites Operating Costs

Operational expenditure costs (“OPEX”) for Santana Phosphate Project have been organized into five categories:

Labor

Electric power

Reagents and consumables

Maintenance

Miscellaneous


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Annual costs have been estimated as of the project operation commencement except for maintenance since spare parts and first fill for the first year were included in the CAPEX.

The accuracy assumed for the OPEX is ±10%. A contingency of 10% was also included in the estimates.

The following summarizes the methods used for Santana Phosphate Project OPEX.

Estimating methodology

Labor relates to personnel including position, manpower quantity, salaries and social charges for the plant as defined by MBAC. The direct labor, indirect labor and administrative personnel were considered as a single group.

Labor

Direct labor includes beneficiation, fertilizer, granulation, acidulation and sulphuric acid plants and utilities. Indirect labor includes laboratory, plant maintenance, asset security, supporting services, internal transportation and logistics personnel. Finally, the administrative group relates to all administrative area personnel, management position level and higher.

Electrical power consumption was calculated in kWh/a across the process plant by applying utilization and efficiency factors to the installed power for each motor and multiplying by the number of hours in a year. Power costs were calculated by multiplying the calculated power consumption of each motor by the electricity rate. A rate of R$ 72.92/kWh was applied due the combination of co-generation and contracted power from the concessionaire. It was assumed a cost of R$ 271.09/kWh and a load reduction of 62.255 MWh/year from the demand study due to the power generation presumed from the sulphuric acid production.

Electricity

Consumption of reagents and consumables are based on consumption rates applied to the mass balance outflow. Prices for reagents have been based on in-house data information.

Reagents and Consumables

Table 21.2.2_1 shows the reagents consumption and supplier.


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Table 21.2.2_1

Reagent Consumption and Vendor

Reagent Unit Consumption

ku/year

Unit Price

USD/unit Vendor

Collector kg 470 0.58

Miracema Nuodex Indústria Química: LIACID 1815 (Soy bean

oil) IPI exemption, 7% ICMS included, CIF.

Alcaline sodium silicate kg 364 1.39 Riberquímica Produtos Químicos.

Supplier to bear freight costs. Caustic soda100% kg 2,377 0.39 Braskem S.A. Final price with taxes

included. FOB plant.

Flocculant kg 13 2.89 BASF: Magnafloc 1011. ICMS,

PIS/CONFINS and IPI excluded. Material at São Paulo.

Hydrated lime kg 8 0.13 Sasil Comercial e Industrial de

Petroquímicos Ltda. Freight CIF and ICMS included.

Aditives (covering) kg 1,620 2.00

Consumables for this project are listed in Table 21.2.2_2 Annual consumption of mill grinding media and linings were estimated from rates based in ATS Projetos previous experience in similar projects. Firewood consumption includes firewood for the granulate superphosphate drying and sulphuric acid steam. Water consumption based on the water balance and plant water cost relates costs to threat the water to a level suitable for plant consumption.

Table 21.2.2_2

Consumables Summary

Consumable Unit Consumption

ku/year Unit Price USD/unit

Rate source

Grinding media/linings kg 303.70 1.92 ATS database Firewood m st 99.59 33.75 ATS database Industrial water m3 4527.00 0.12 ATS database

Raw materials are summarized in Table 21.2.2_3.

Table 21.2.2_3

Raw Material Summary

Raw Material Unit Consumption ku/year

Unit Price 2011 USD/unit (1)

Rate source

Sulphur t 82.20 350.00 Based on an independent authority in fertilizer industry

Note: (1) The Sulphur cost through 2020 was based on Sulphur prices FOB Tampa furnished by a leading independent authority in fertilizer industry. For the period after 2020, the price was escalated at a rate of 2% per annum


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Maintenance costs were budgeted from plants costs based on quotations or experience derived from other similar projects. Maintenance costs were defined by applying a CAPEX factor upon the total costs. Maintenance includes spare parts and lubricants

Maintenance

Table 21.2.2_4

Shows the CAPEX factor for each project area. Description CAPEX Factor (%)

Infra-structure 1 Beneficiation 2

Shipment 1 Acidulation 3 Granulation 3

Sulphuric Acid Plant 3 Utility System 3

Support Installations 1 Energy, Telecommunication, Automation 1

Miscellaneous costs are those related to specialized outsourced work and other activities such as community relationship, insurance, building maintenance, etc. Total of miscellaneous annual costs estimates is US$ 4.9 million.

Miscellaneous

21.2.3 Extra Operating Costs

Besides the costs of mining and industrialization, an estimate of the concentrate freight costs has been prepared. The concentrate will be transported from the Mine Site to the Industrial Site.

The total annual freight cost of the concentrate amounts to R$ 3.9M, which refers to an annually transport of 353k tonnes of an intermediary product (phosphate concentrate with 85% of solids).

In addition to freight costs, it was considered the payment to CFEM (Brazilian Federal Contribution of Mineral Exploration) that represents 2% of the total cost of the Concentrate and a payment of royalties of 1% also of the total cost of the Concentrate.

21.2.4 Total SSP Annual Cost

Based on all operating costs assumptions, Project Santana SSP annual cost at site is US$127/tonne. Table below shows SSP cost breakdown. A 10% contingency was included in these estimates.

R$/US$ exchange rate considered was 1.60.


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Figure 21.2.4_1 SSP Annual Costs

22 ECONOMIC ANALYSIS

Please note that this preliminary economic assessment is preliminary in nature, includes inferred mineral resources that are considered too speculative to have the economic considerations applied to turn that would enable them to be categorized as mineral reserves, and there is no certainty that this preliminary economic assessment will be realized. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

22.1 SSP Long term price

The SSP price through 2020 was based on TSP prices furnished by a leading independent authority in fertilizer industry (FOB Morocco) and adjusted for additional logistics costs for freight to the targeted market for the Project. For the period after 2020, the price was escalated at a rate of 2% per annum.

The table below shows the real prices from 2015 to 2020 that were contemplated to the project economic analysis. Prices considered currency of 2011.

Table 22.1_1

Real SSP Prices (US$/tonne) 2015 2016 2017 2018 2019 2020

$ 350.9 $ 352.8 $ 354.6 $ 358.3 $ 363.8 $ 352.6 22.2 Valuation model

22.2.1 Assumptions

The PEA indicates that the Project is expected to generate robust returns. The assumptions for the economic analysis are as follows:

The estimated capital cost is US$ 385 M including US$ 73 M of contingencies and US$ 23 M of working capital.

Capital Cost


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Capital Expenditure (US$ 362 M) will be spent through the years 2013, 2014 and 2015 in a ratio of 10%, 25%, and 65% respectively. The remaining US$ 23 M of working capital will be spent in 2015.

The total sustaining distribution is shown below

Sustaining Capital

Sustaining Capital – US$ 1 M each year and additional of US$ 2 M every 3 years

Mine Equipment: US$ 23.3 M starting in 2016 and used during the whole mine life

Closure Cost: US$ 20 M, being US$ 10 M in 2037 and US$ 10 M in 2038

Operating Costs include labor, electricity, consumables, fuel and lubricant, equipment maintenance, management and administrative expenditures, and all others costs contemplated in the section 21.2. The model also contemplated selling costs of R$ 4.0 M (currency of 2011).

Operating Costs

All Investments, except land, were depreciated with no residual value at a rate of 10% in ten years, (Straight Line Method).

Depreciation

Working Capital was estimated using average number of days of receivables, payables and stocks, as shown below:

Working Capital

Receivables : 90 days

Payables : 30 days

Stocks : 70 days

22.2.2 Economics

Project life is 24 years after ramp-up, from 2015 to 2038.

Project Life

The Economic Evaluation of the Project was done using the Discounted Cash Flow Approach.

Economic Evaluation

Sales were calculated net of taxes.

Sales


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Taxes on Profits calculated according to Brazilian Tax Law.

Taxes

Income Tax: 15% plus 10% on the excess of R$ 240,000.00

Social Contribution: 9%

A reduction of 75% in the Income Tax was considered according to fiscal incentives (SUDAM area) during the period between 2016 and 2025.

Based on project assumptions preliminary financial model indicates robust project economics with a Net Present Value of US$ 423 million as at Q3 2013 (the estimated start of construction). This figure does not include US$20 million required for exploration and engineering development during 2011 and 2012.

Economic Indicators

The following economic indicators were used in the financial analysis:

Weighted Average Cost of Capital (WACC): 10%

Internal Rate of Return (IRR): 26.27%

Net Present Value @ WACC: US$ 423,340,292

Payback Period: 4.53 years

22.2.3 Sensitivity Analysis

Risk Analysis was done using the sensitivity of relevant inputs creating scenarios.

The Sensitivity Analysis showed low risk. All scenarios in the sensitivity analysis were feasible as shown in the table below (in US$ million):

Table 22.2.3_1

NPV at Q3 2013 Sensitivity Analysis by SSP and Sulphur prices

USD Million SSP Price (%VAR)

Sulphur Price CIF SAN (%VAR)

-20% -10% 0% 10% 20%

-20% 222.8 336.9 450.1 563.4 676.7 -10% 209.2 323.5 436.7 550.0 663.3 0% 195.5 310.1 423.3 536.6 649.9 10% 181.8 296.7 409.9 523.2 636.5 20% 168.1 283.3 396.5 509.8 623.1


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Table 22.2.3_2

NPV at Q3 2013 Sensitivity Analysis by SSP Price and Total CAPEX


Total CAPEX (%VAR) -20% -10% 0% 10% 20%

-20% 257.5 372.1 485.3 598.6 711.9 -10% 226.5 341.1 454.3 567.6 680.9 0% 195.5 310.1 423.3 536.6 649.9 10% 164.5 279.1 392.3 505.6 618.9 20% 133.5 248.1 361.3 474.6 587.9

Table 22.2.3_3

NPV at Q3 2013 Sensitivity Analysis by SSP Price and Total OPEX


Total OPEX (%VAR) -20% -10% 0% 10% 20%

-20% 291.3 404.5 517.8 631.1 744.4 -10% 243.5 357.3 470.6 583.9 697.1 0% 195.5 310.1 423.3 536.6 649.9 10% 147.4 262.6 376.1 489.4 602.7 20% 99.4 214.6 328.9 442.1 555.4

Table 22.2.3_4

NPV at Q3 2013 Sensitivity Analysis by SSP Price and WACC


WACC (%)

-20% -10% 0% 10% 20%

8% 290.6 431.2 570.3 709.5 848.7 9% 239.4 366.1 491.4 616.7 742.1 10% 195.5 310.1 423.3 536.6 649.9 11% 157.6 261.7 364.4 467.2 569.9 12% 125.0 219.7 313.3 406.8 500.3


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23 ADJACENT PROPERTIES

There are no adjacent or nearby phosphate permits to those of MBAC. The Santana Phosphate Project is related to the east extension of the regional 450km long northwest-southeast Cuiú-Cuiú - Tocantinzinho lineament which also hosts several important gold deposits including the Palito mine, Tocantinzinho deposit and Cuiú-Cuiú, Bom Jardim and Batalha gold prospects.

24 OTHER RELEVANT DATA AND INFORMATION

Amazon/NCL is not aware of other relevant data pertaining to the Santana Phosphate Project.

25 INTERPRETATION AND CONCLUSIONS

The style of phosphate mineralization is unique to phosphate occurrences in Brazil and will potentially change the current mind set of explorers for this commodity in this region.

MBAC has undertaken a systematic exploration program in the last year that has been successful in defining significant resources of phosphate in close proximity to one of the largest agricultural centres in Brazil.

The drilling has defined an inferred mineral resource which will require additional exploration work to increase the current resource confidence. Overall, Amazon Geoservices concludes that there are no fatal flaws in the current resource data.

NCL has undertaken a study of the mine production schedule and plant feed schedule, based on the pit contained resources inventory, for a processing rate of 1.5 million tonnes per year and a life of mine of 24 years, for an average annual production rate of 300,000 tonnes of concentrate with an average grade of 34% P2O5. Based on project assumptions preliminary financial model indicates robust project economics with a NPV of US$ 423 million.

The pertinent observations and interpretations which have been developed in producing this report are detailed in the sections above.

It is not anticipated that there are any reasonable foreseeable risks or uncertainties on the potential viability of this project.


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26 RECOMMENDATIONS

26.1 Exploration and Resources

Drilling and studies completed to date have defined an inferred mineral resource at Santana. The data collected is considered to be of moderate quality and suitable for resource estimation.

Further scope exists to improve the geological and mineral resource estimation confidence in the regions currently defined as an inferred mineral resource. Amazon Geoservices makes the following specific recommendations:

To review the RC and DC twin holes to determine the grade effect of the wet RC drilling that has been observed by Amazon (RC data was minimal in the current mineral resource estimate). If wet samples are biasing the RC results then RC drilling should be suspended until dry samples can be guaranteed by the drilling contractor

To improve the survey accuracy of both drillhole collars and topography

Increase the drill density to 50m by 50m to allow indicated and potentially measured resources to be defined

To follow up on the failed certified standards and blank material and to actively review this QAQC data on a batch by batch basis

To test via deeper drilling the lower grade phosphate and CaO potential

26.2 Mining

Using an updated mineral resource estimation, which should include a significant proportion of higher confidence Indicated category resources, and utilizing the results of the Pre-Feasibility study, a mineral reserve estimation will be reported. With these results, more detailed mine planning and production schedules should be generated for the open pit.

The following activities necessary to complete a pre-feasibility study need be carried out: infill resource definition drilling, geotechnical and metallurgical studies, and EIA-SIA work, and the mining permitting.


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26.3 Mineral resource and Evaluation Budget

MBAC has also provided Amazon/NCL with an ongoing exploration and evaluation budget, summarised in Table 26.3_1 below.

Table 26.3_1


Proposed Resource and Evaluation Expenditure

Activity Total (US$) DC and RC drilling $ 3,000,000 Assaying and Characterization $ 200,000 Geology $ 50,000 Pre-Feasibility Work $ 500,000 Travel and accommodation $ 30,000 Field supervision and support $ 150,000 Administration $ 70,000 Sub-total $4,050,000

The proposed expenditure of US$ 4,050,000 in Year 1 is considered to be consistent with the potential of the Santana Phosphate Project and is adequate to cover the costs of the proposed programs.


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27 REFERENCES

AusIMM. 1995. Code and Guidelines for Assessment and Valuation of Mineral Assets and Mineral Securities for Independent Expert Reports (The Valmin Code) Issued April 1998. AusIMM.

AusIMM. 1998. Code and Guidelines for Assessment and Valuation of Mineral Assets and Mineral Securities for Independent Expert Reports (The Valmin Code), issued April 1998. The Australasian Institute of Mining and Metallurgy.

2011 – MBAC Internal technical Report on Santana Phosphate Project – L Silva, MH Waring, R Lepine, M. Tedesco, P. Xavier, D. Lopes

ALS laboratory group- Mineral Division- Informe de Análisis Mineralogico MLA. N0 IF10170088. February, 7, 2011.

LCT-USP- Resultado de Identificação de fases por difratometria de raios-X- LCT 228-1152.HPF a 228-1170.HPF.

Estudo exploratório de flotação. Projeto Santana do Araguaia/PA. RT-PROC-SAN001. Campos Belos, Abril 2010.

Ensaios Work Index de Bond. RPT HDA/MBAC01/11 Rev.0-31/08/2011

Resultados detalhados de flotação com amostras de furos de sonda. Santana. Planilha Excel – 02/09/2011

GLS Consultores Associados Ltda. Primeiro Relatório de Andamento. Testes de laboratório para ácido fosfórico. DOC n 0 018-2011 MBAC REV 00.


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28 DATE AND SIGNATURE PAGE

The “qualified persons” (within the meaning of NI43-101) for the purposes of this report are Beau Nicholls, who is an employee of Amazon Geoservices and Carlos Guzman, who is an employee of NCL. The effective date of this report is 22 September 2011.

(signed by)

Beau Nicholls B.Sc Geol. MAIG Principal Consulting Geologist Amazon Geoservices Ltda. Signed on the 22 September 2011

(signed by)

Carlos Guzmán RM Chilean Mining Commission Principal /Project Director NCL Brasil Ltda. Signed on the 22 September 2011


6005 Australia

29 CERTIFICATES OF QUALIFIED PERSONS

Amazon Geoservices Ltda.

I, Beau Nicholls, do hereby certify that:

Certificate of Qualified Person

1. I have been working since 2010 as the Principal Consulting Geologist with the firm Amazon Geoservices Ltda. of 105/1203 Muzambinho Avenue, Anchieta, Belo Horizonte,-MG, Brazil, CEP 30310-280. My residential address is 463 Alameda Das Quaresmeiras, Rio Acima - MG, Brazil, CEP - 34300-000

2. I am a practising geologist with 17 years of Mining and Exploration geological experience. I have worked in Australia, Eastern Europe, West Africa and the Americas. I am a member of the Australian Institute of Geoscientists (“MAIG”)

3. I am a graduate of Western Australian School of Mines – Kalgoorlie and hold a Bachelor of Science Degree in Mineral Exploration and Mining Geology (1994)

4. I have practiced my profession continuously since 1995

5. I am a “qualified person” as that term is defined in National Instrument 43-101 Standards of Disclosure for Mineral Projects (the “Instrument”)

6. I have visited the Santana Phosphate Project between the 25th and 27th July 2011

7. I am responsible for sections 4 to 12, 14 and jointly responsible of sections 1 to 3 and 23 to 27 of the technical report dated effective 22 September 2011 and titled “Preliminary Economic Assessment- Santana Phosphate Project, Para State, Brazil”(the “Report”)

8. I am independent of MBAC Fertilizer Corp pursuant to section 1.5 of the Instrument

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

10. I do not have nor do I expect to receive a direct or indirect interest in the Santana Phosphate Project of MBAC Fertilizer Corp and I do not beneficially own, directly or indirectly, any securities of MBAC Fertilizer Corp or any associate or affiliate of such company

11. I have not had any prior involvement with the Santana Phosphate Project of MBAC Fertilizer Corp

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

Dated at Belo Horizonte, Brazil, on 22 September, 2011

(signed by)

Beau Nicholls BSc(Geology) MAIG Principal Consulting Geologist


6005 Australia

NCL Brasil Ltda.

I, Carlos Guzmán, do hereby certify that:

Certificate of Qualified Person

1. I have been working since 2001 as the Principal Mining Engineer and Project Director with the firm NCL Brasil Ltda. of Alameda da Serra 500/315, Nova Lima,-MG, Brazil, CEP 34000-000

2. I am a practising mining engineer and registered member with the Chilean Mining Commission

3. I am a graduate of the Universidad de Chile and hold a Mining Engineer title (1995)

4. I have practiced my profession continuously since 1995

5. I am a “qualified person” as that term is defined in National Instrument 43-101 Standards of Disclosure for Mineral Projects (the “Instrument”)

6. I have visited the Santana Phosphate Project between the 25th and 27th July 2011

7. I am responsible for sections 13, 15 to 22 and participation in sections 1 to 3, and 23 to 27 of the technical report dated effective 22 September 2011 and titled “Preliminary Economic Assessment - Santana Phosphate Project, Para State, Brazil” (the “Report”)

8. I am independent of MBAC Fertilizer Corp pursuant to section 1.5 of the Instrument

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

10. I do not have nor do I expect to receive a direct or indirect interest in the Santana Phosphate Project of MBAC Fertilizer Corp and I do not beneficially own, directly or indirectly, any securities of MBAC Fertilizer Corp or any associate or affiliate of such company

11. I have not had any prior involvement with the Santana Phosphate Project of MBAC Fertilizer Corp

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

Dated at Belo Horizonte, Brazil, on 22 September, 2011

(signed by)

Carlos Guzmán RM, Chilean Mining Commission Principal Mining Engineer / Project Director

Santana Pea (Sedar Copy)

Documents

Transcript of Santana Pea (Sedar Copy)