EX-2.11 13 exh_211.htm EXHIBIT 2.11

Exhibit 2.11

São Francisco Mine

NI 43-101 Technical Report

RESOURCE AND RESERVE ESTIMATES ON

THE SÃO FRANCISCO MINE IN THE MUNICIPALITY OF VILA BELA DA SANTÍSSIMA

TRINDADE, STATE OF MATO GROSSO

BRAZIL

NI 43-101 Technical Report

JANUARY 31/2012

EFFECTIVE DATE: September 30, 2011

PREPARED FOR: AURA MINERALS INC.

J. Britt Reid – P. Eng. Executive Vice President and Chief Operating Officer, Aura Minerals Inc.

Bruce Butcher – P. Eng. Vice President, Technical Services, Aura Minerals Inc.

Chris Keech – P. Geo. CGK Consulting Services Inc.

São Francisco Mine

NI 43-101 Technical Report

Table of Contents

1.0 SUMMARY	1
1.1 Introduction and Property Description	1
1.2 Location and Accessibility	1
1.3 Geology and Mineralization	1
1.4 Exploration	3
1.5 Metallurgical Testing	5
1.6 Mineral Resources	6
1.7 Mineral Resources Summary	6
1.8 Mineral Reserve	6
1.9 Mineral Reserve Summary	7
1.10 Mining Methods	7
1.11 Recovery Methods	8
1.12 Project Infrastructure	8
1.13 Environmental Studies, Permitting and Social Impact	8
1.14 Capital and Operating Costs	9
1.15 Economic Analysis	10
1.16 Interpretations, Conclusions and Recommendation	10
2.0 INTRODUCTION	12
3.0 RELIANCE ON OTHER EXPERTS	14
4.0 PROPERTY DESCRIPTION AND LOCATION	15
4.1 Location	15
4.2 Review of Mineral Properties and Titles in Brazil	16
4.2.1 Mineral and Mining Rights Tenure	16
4.2.2 Surface Rights	17
4.3 Royalties	17
4.4 Environmental Permits, Studies, and Considerations	18
5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY	19
5.1 Physiography	19
5.2 Accessibility	19
5.3 Climate	19
5.4 Local Resources and Infrastructure	19
6.0 HISTORY	21
6.1 Previous and Current Owners	21
6.2 Exploration and Drilling	21
6.3 Previous Mineral Resource and Reserve Estimates	22
6.4 Historical Production	22
7.0 GEOLOGICAL SETTING AND MINERALIZATION	23
7.1 Regional Geology	23
7.2 Local Geology	26
7.3 Mineralization	27

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8.0 DEPOSIT TYPES	28
9.0 EXPLORATION	29
9.1 Sampling Method and Quality	30
9.2 Densities	32
10.0 DRILLING	33
11.0 SAMPLE PREPARATION, ANALYSES, AND SECURITY	36
11.1 Sample Preparation and Analysis	36
11.1.1 Drill Core Samples	36
11.1.2 Sample Security	38
11.2 Author's Opinion on Sample Preparation, Security, and Analytical Procedures	38
12.0 DATA VERIFICATION	39
12.1 Previous Exploration QA/QC Programs	39
12.2 Apoena Exploration QA/QC Program	39
12.2.1 Blanks	39
12.2.2 Standards	40
12.2.3 Twin Samples	43
12.2.4 Duplicate Pulp Assays	43
13.0 MINERAL PROCESSING AND METALLURGICAL TESTING	45
13.1 Pilot Scale Column Tests	45
13.2 Direct Column Leach Tests	45
13.3 Pilot Gravity Tests	46
13.4 KCA Column Leach Tests of Gravity Pilot Plant Ore	47
14.0 MINERAL RESOURCE ESTIMATES	48
14.1 Geological Interpretation	48
14.1.1 Long-term Wireframes	48
14.1.2 Short-term Wireframes	49
14.2 Exploratory Data Analysis	49
14.2.1 Exploration Drill Hole Database	49
14.2.2 Grade Control Database	50
14.2.3 Discussion and Interpretation	50
14.3 Geological Modeling	51
14.3.1 Block Model Parameters	51
14.3.2 Indicator Block Model	51
14.4 Composite Statistics and Capping	56
14.4.1 Reconciliation to Blast Hole Data	57
14.5 Grade Estimation	59
14.5.1 Variography	59
14.5.2 Estimation Parameters	60
14.6 Block Model Validation	61
14.6.1 Global Mean and Global Variance Validation	61
14.6.2 Trend Check	61
14.6.3 Change of Support Validation	63
14.6.4 Visual Validation	64

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14.7 Mineral Resource Classification	65
14.8 Mineral Resource Classification	67
15.0 MINERAL RESERVE ESTIMATES	68
15.1 Mineral Reserve Summary	69
16.0 MINING METHODS	70
16.1 Drilling and Blasting	70
16.2 Loading and Hauling	70
16.3 Pit Slope Stability	70
16.4 Road Maintenance	71
16.5 Dewatering	71
16.6 Maintenance	71
16.7 Power	71
16.8 Communications	71
16.9 Mine Plan and Schedule	71
16.10 Mining Fleet and MachinerY	77
17.0 RECOVERY METHODS	79
17.1 Process Plant Description	79
17.2 Plant Design, Equipment Characteristics, and Specifications	81
17.2.1 Crushing Circuit	81
17.2.2 Gravity Circuit	83
17.2.3 Heap Leach Pads	85
17.2.4 Carbon ADR Facility	87
18.0 PROJECT INFRASTRUCTURE	89
19.0 MARKET STUDIES AND CONTRACTS	90
20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL IMPACT	91
21.0 CAPITAL AND OPERATING COSTS	92
22.0 ECONOMIC ANALYSIS	93
23.0 ADJACENT PROPERTIES	94
24.0 OTHER RELEVANT DATA AND INFORMATION	95
25.0 INTERPRETATION AND CONCLUSIONS	97
26.0 RECOMMENDATIONS	98
27.0 REFERENCES	99
28.0 DATE AND SIGNATURE PAGE	100

LIST OF TABLES

Table 1-1 – September 30, 2011 Mineral Resource Estimate	6
Table 1-2 - Pit and Cost Parameters	7
Table 1-3 – September 30, 2011 Mineral Reserve Estimate	7
Table 1-4 –Capital Expenditures	9

iii

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Table 1-5 - 2011 Operation Cost Breakdown	9
Table 4-1 - DNPM Permits	16
Table 6-1 – Exploration Program	21
Table 6-2 – December 31, 2010 Mineral Resource Estimate	22
Table 6-3 – December 31, 2010 Mineral Reserve Estimate	22
Table 6-4 - Historic and Recent Production	22
Table 9-1 - Exploration Program	30
Table 10-1 – Apoena Minerals Drill Hole Collar Locations	34
Table 10-2 - Apoena Minerals Drilling Significant Intervals 2010-2011	34
Table 12-1 - Standards to Check 2009-2011 Drilling Results	40
Table 13-1 - ROM Large Column Test Results	45
Table 13-2 - Summary of KCA Bottle Roll Leach Tests	45
Table 13-3 - Summary of KCA Direct Column Leach Tests	46
Table 13-4 - Summary of the Gravity Plant Column Leach Tests	47
Table 14-1 - Summary Statistics of Exploration Database - Au (g/t)	49
Table 14-2 - Summary Exploration Lithology Code	49
Table 14-3 - Summary Statistics of Exploration Database - Au (g/t) - Coded by Exploration Wireframes	50
Table 14-4 - Summary Statistics of Grade Control & Blast Hole Database - Au (g/t)	50
Table 14-5 - São Francisco Block Model Parameters	51
Table 14-6 - São Francisco 0.15 g/t Au Indicator Variogram Parameters	53
Table 14-7 - Estimation Parameters for 0.15 g/t Au Indicator	54
Table 14-8 - São Francisco Estimation Domain Codes	55
Table 14-9 - São Francisco Estimation Domain Codes	55
Table 14-10 - Summary Statistics of São Francisco 5 m Composites - Au (g/t)	56
Table 14-11 - Capping Summary of São Francisco 5m Composites - Au (g/t)	58
Table 14-12 - Summary Statistics of São Francisco - Capped 5m Composites - Au (g/t)	58
Table 14-13 - São Francisco Semi-Variogram Parameters	59
Table 14-14 - São Francisco Estimation Parameters for Ordinary Kriging	60
Table 14-15 - Composite, NN, OK and ILC Statistics for Domain 51	61
Table 14-16 – September 30, 2011 Mineral Resource Estimate	67
Table 15-1 – Pit and Cost Parameters	68
Table 15-2 - September 30, 2011 Mineral Reserve Estimate	69
Table 16-1 - 4th Quarter 2011 Material Movement Forecast	71
Table 16-2 – 1st Quarter 2012 Material Movement Forecast	72
Table 17-3 - 2nd Quarter 2012 Material Movement Forecast	73
Table 16-4 - 3rd Quarter 2012 Material Movement Forecast	73
Table 16-5 - 4th Quarter 2012 Material Movement Forecast	74
Table 17-6 – 1st Quarter 2013 Material Movement Forecast	75
Table 17-7 - 2nd Quarter 2013 Material Movement Forecast	75
Table 16-8 - 3rd Quarter 2013 Material Movement Forecast	76
Table 16-9 - Mine Sequence Results – Total	77
Table 16-10 - Mine Equipment	77
Table 16-11 - 2010 Contract Haulage Cost	78

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NI 43-101 Technical Report


Table 17-1 - 2011 Monthly Plant Production for São Francisco Mine	81
Table 21-1 - Capital Expenditures	92
Table 21-2 - 2011 Operation Cost Breakdown	92

LISt of FIGURes

Figure 4-1 - Location Map from Rezende Eng. Minas S/S Ltda. November 2006	15
Figure 4-2 - Site Location from Rogerio Moreno, 2011	16
Figure 4-3 -Concession Map (from Scott Wilson RPA)	17
Figure 7-1 - Regional Geology from Rezende Eng. Minas S/S Ltda. November 2006	24
Figure 7-2 - Regional Stratigraphic Column from Rezende Eng. Minas S/S Ltda. November 2006	25
Figure 10-1 - 2008 Drill Hole Locations	33
Figure 10-2 - Cross Section 11,775 NW	35
Figure 12-1 - Blank Analysis	40
Figure 12-2 - Results from Standard 2Pd	41
Figure 12-3 - Results from Standard 18C	41
Figure 12-4 - Results from Standard 18Pb	42
Figure 12-5 - Results from Standard 6Pc	42
Figure 12-6 - Twin Sample Assays	43
Figure 12-7- Pulp Check Assays	44
Figure 14-1- São Francisco 735m Plan View – 0.15 g/t Au Indicator	52
Figure 14-2- São Francisco Cross Section 250S – 0.15 g/t Au Indicator	53
Figure 14-3 - São Francisco 0.15 g/t Au Indicator Variogram Model	54
Figure 14-4 - São Francisco Contact Profiles Across LG-HAZ/HG-HAZ Contact	55
Figure 14-5 - São Francisco Histogram and Lognormal Probability Plot of 5 m Composites	56
Figure 14-6 - Comparison of Blast Hole Estimate and Exploration OK Estimates	58
Figure 14-7 - São Francisco Pairwise Relative Semi-variogram and Model for Domain 50	59
Figure 14-8 - São Francisco Pairwise Relative Semi-variogram and Model for Domain 51	60
Figure 14-9 - Swath Plot for Domain 51 Comparing NN and OK Au (g/t) Grades in the East Direction	62
Figure 14-10 - Swath Plot for Domain 51 Comparing NN and OK Au (g/t) Grades in the North Direction	62
Figure 14-11 - Swath Plot for Domain 51 Comparing NN and OK Au (g/t) Grades in the Vertical Direction	63
Figure 14-12 - Grade-Tonnage Curves of the OK Estimates (LT), and the Predicted 10x10x10 m SMU	64
Figure 14-13 - Section 11,175 NW – OK Block Model Grades and 5 m Composites	65
Figure 14-14 - Section 11,775 NW – OK Block Model Grades and 5 m Composites	65
Figure 16-1 – 2011 4th Quarter Material Movement Forecast	72
Figure 16-2 2012 1 Quarter Material Movement Forecast	72
Figure 16-3 –2nd Quarter 2012 Material Movement Forecast	73
Figure 16-4 –3rd Quarter 2012 Material Movement Forecast	74
Figure 16-5 –4th Quarter 2012 Material Movement Forecast	74
Figure 16-6 –1st Quarter 2013 Material Movement Forecast	75

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NI 43-101 Technical Report

Figure 16-7 –2nd Quarter 2013 Material Movement Forecast	76
Figure 16-8 –3rd Quarter 2013 Material Movement Forecast	76
Figure 16-9 – Mine Sequence Results - Total	77
Figure 17-1 - General Flow Sheet São Francisco Mine	80
Figure 17-2 - Process Flow Diagram ("PFD") of the Crushing Circuit	82
Figure 17-3 - PFD of the Gravity Circuit	84
Figure 17-4 - PFD of Heap Leach Circuit	86
Figure 17-5 - PFD of the ADR facility	88
Figure 24-1 – Monthly Reconciliation between LT and ST Block Model Results	96
APPENDICES
Appendix 1 - Glossary of Mining and Related Terms and Abbreviations

São Francisco Mine

NI 43-101 Technical Report

1.0

SUMMARY

1.1

Introduction and Property Description

Aura Minerals Inc. (“Aura”) has prepared a technical report (the “Report”) compliant with National Instrument 43-101 (“NI 43-101”) on the updated Mineral Resource and Reserve estimates pertaining to its São Francisco Mine (or the “Mine”), located in Mato Grosso State in Brazil. The Mine’s surface and mineral rights are owned by Mineração Apoena S.A. (“Apoena”), a beneficially-owned indirect subsidiary of Aura. MCB Serviços e Mineração Ltda. (“MCB”) was retained by Aura in 2010 to prepare an independent technical report which was filed March 30^th 2011. MCB recommended that the resource model be updated, and the estimation parameters be modified to improve the reconciliation between the model and production. As a result of MCB’s recommendation, a new mine plan and this Report have been prepared by Aura to update the Mineral Resource and Reserve estimates pertaining to the Mine.

The São Francisco Mine is located in the western portion of Mato Grosso State in west central Brazil, close to the Bolivian frontier some 560 km west of the capital city of Cuiabá.

The Mine consists of four contiguous mining and exploration permits totalling 36,308.34 hectares granted between 1980 and 1982 by the Departamento Nacional da Produção Mineral (“DNPM”), Brazil’s department of national mineral production. The process numbers for the mine permits are 860938/1982 and 860937/1982. The permitting process for the Mine has been completed and all the relevant licenses for the operation of the São Francisco Mine are in place. There are no current environmental liabilities or non-compliance issues for the property.

The São Francisco Mine is an open-pit, heap leach gold mine that involved two separate gold recovery processes including crushing-gravity gold recovery-heap leach and run-of-mine heap leach up until the start of 2011. As of early 2011, the mine has focused on the more profitable crushing-gravity gold recovery-heap leach process rather than run-of-mine heap leach. The ore contains a significant component of gravity gold, which requires detailed sampling and attention to mine planning to ensure that the gravity gold is recovered prior to placement of ore on the leach pad.

1.2

Location and Accessibility

The São Francisco property is located approximately 560 km west of Cuiabá, the capital of Mato Grosso State. There are daily commercial jet flights to Cuiabá from São Paulo, Rio de Janeiro and other major Brazilian cities. There is also good road access from Cuiabá to the São Francisco Mine. Pontes e Lacerda, the main town close to the mine site, is accessible by a 435 km paved highway from Cuiabá. The open pit is located at approximately 14°50’S latitude and 59°37’W longitude.

1.3

Geology and Mineralization

The São Francisco Mine is a shear hosted lode gold deposit. It is epigenetic, structurally controlled, and composed of narrow, 1 to 5 cm wide, quartz veins containing free gold. The veins and vein systems and stockworks both parallel and crosscut the bedding planes and appear to represent separate but closely related mineralizing events.

The regional geological setting for the São Francisco Mine and numerous other gold occurrences that comprise the Guaporé Gold Belt of central west Brazil and east central Bolivia is the Aguapeí Mobile and Mafic Arc Belt. This belt follows a major crustal-scale break or shear zone separating the Amazon Craton of Archean-age (3,800 to 2,500 million years ago) on the east from the Paragua Craton of Proterozoic-age (2,500 to 1,600 million years ago) on the west. The belt extends more than 600 km in a NNW-SSE direction and is characterized by a prominent mountain range made up of a 1,200 m thick sequence of Proterozoic-age sedimentary rocks known as the Aguapeí Group.

São Francisco Mine

NI 43-101 Technical Report

The Aguapeí Group, the host rocks for the gold mineralization, is a sequence of texturally and mineralogically supermature detrital sediments made up of braided river facies, Aeolian (wind-deposited) dunes, and shallow marine platform facies. Southward along the belt, the lower part of the Aguapeí Group contains interbedded volcanic units and mafic sills and dykes (that may be thrusted from the east). The Aguapeí Group overlies the central part of the Amazon Craton (Brazilian Precambrian Shield), locally known as the Xingu Complex. The Complex contains lower Proterozoic volcano-sedimentary belts elongated in a NW-SE direction and surrounded by Archean metamorphic rock masses. The volcano-sedimentary belts and the Xingu Complex have both been intruded by large bodies of granitic rock. The flat area surrounding the mountain range is believed to be mainly underlain by the Xingu Complex, but most of the area is covered by residual soils with few outcrops and poorly known geology. Prolonged and deep erosion of this continental mass during Proterozoic time was accompanied by the development of basins in which were deposited the 1,200 m thick Aguapeí Group of sediments. These sediments, now metamorphosed, have been mapped over a 300 km strike length in Brazil, and continue southward into Bolivia for at least another 200 km, and then pass again into Brazil.

Structurally, the Aguapeí Group rocks have been subjected to a NW-SE compression that folded the eastern edge of the belt into a series of broad to tight folds. These folds form NNW-SSE ridges that run parallel to the axis of the mobile belt. Faulting, fracturing and shearing have also developed, some running parallel to the axis, but with at least one well-developed NE-SW fracture system that crosscuts the regional trend. The mountain range is bounded on both sides by faults, with the fault on the east dipping away from the range at a shallow angle. This fault separates the Archean basement on the east from the Proterozoic metamorphosed sediments on the west. The internal part of the Aguapeí Mobile Belt away from the mountain front contains extensive plateaus of Aguapeí Group rocks that show little or no deformation.

The known bedrock gold deposits and occurrences in Mato Grosso State are separated into two districts: the São Vicente Borda district and the Pontes e Lacerda district (to the south of the São Francisco Mine).

The local rocks at the São Francisco Mine have been subjected to low-grade metamorphism. They consist of fine to coarse-grained meta-arenites (metamorphic sandstones), with locally reddish-coloured metapelites (metamorphic mudstones) and occasionally metaconglomerates (old pebble beds) of the Fortuna Formation, the basal unit of the Aguapeí Group. The metaconglomerates are composed of quartz pebbles and grains set in a siliceous sandy matrix. Primary sedimentary structures such as cross-bedding, graded bedding, and ripple marks are commonly observed.

The rock units are folded into a series of broad folds that can be traced over several kilometres. The folds trend NNW-SSE and plunge NW. They are faulted and sheared, generally parallel to the folding, and are crosscut by fractures that strike WSW-ENE.

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NI 43-101 Technical Report

Mineralization is enclosed by a hydrothermal alteration zone (“HAZ”). The HAZ occurs as a regular steeply dipping tabular zone; the depth of the HAZ has yet to be defined by drilling. Alteration associated with mineralization in the HAZ consists of silicification and occasional sericite and chlorite.

The gold occurs as free gold and frequently as coarse nuggets measuring several millimeters in diameter with the quartz, as laminations along the fracture planes, and within limonite boxwork after pyrite and arsenopyrite.

1.4

Exploration

Exploration work was conducted at the São Francisco Mine between 1985 and 1997 by Santa Elina Industria e Comercio S/A (“Santa Elina”) or companies working for or with Santa Elina. There were no exploration activities from 1998 to late 2003.

Exploration at the São Francisco Mine was resumed by Yamana Gold Inc. (“Yamana”) in November 2003 and a total of 49 core holes, a 150 m vertical shaft and 107 m of underground drifts were completed. As well, bulk samples were taken from nine surface trench sites (each about 100 m long x 4 m wide x 2.5 m deep across the strike of the deposit) and the bulk samples ranged in weight from 265 tonnes to 2,208 tonnes each.

The 49 core holes (44 were vertical and 5 were inclined) totalling 15,030 m were drilled across the entire mineralized area to collect material for metallurgical testing. In 2005, Yamana drilled 104 core holes totalling 10,231 m to upgrade Inferred Mineral Resource to Indicated Mineral Resource and create the basis for a Mineral Reserve update.

In addition, an extensive metallurgical testing program was conducted at the Kappes Cassiday and Associates (“KCA”) Reno facilities, where large pilot-sized column leach tests were conducted on the São Francisco Mine’s ore. The material was tested at the São Vicente facility and pilot-plant sized gravity concentration circuit tests were conducted on-site at the São Francisco Mine.

In 2006, Yamana drilled 20 core holes totalling 4,620 m to check the continuity of the gold mineralization along the open strike extensions of the deposit to the northwest and southeast. In 2007, a total of 7,684 m were drilled in 18 core holes by Yamana. Three of these holes were drilled to investigate the extension of the deposit at depth, and the others being infill holes. In 2008, Yamana drilled 9 core holes totalling 2,823 m to infill the depth extension of higher-grade mineralization at the southeast and northwest ends of the deposit.

To the end of December 2008, the drilling totalled 90,580 m in 507 drill holes. The majority of the Mineral Resource has been drilled on a 25 m x 25 m spacing and core recoveries average better than 95%. Note that a limited number of reverse circulation (‘RC”) holes were drilled by Santa Elina in 1997.

From 1985 to 1997 Santa Elina drilled HX (76.2 mm diameter) and NQ (47.6 mm diameter) holes in order to obtain samples of sufficient size. Both core sizes were sampled in 2 m length, but only half the HX core was sent for analysis, while the whole NQ core was submitted. This sampling methodology was maintained, but in 2007 Yamana changed the protocol and retained one-quarter of the NQ core as an archive. Diamond drilling by Yamana in 2008 used HQ rods, and the core (63.5 mm diameter) was sawn in half for assaying.

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NI 43-101 Technical Report

From 1990 to 1997, samples were taken at 2 m intervals from one-half of the diamond-sawn core. Samples were crushed on site to quarter-inch and pulverized to minus 2 mm in a hammer mill. A fraction of this was saved for ore characterization, the rest was panned to remove the so-called “gravity” gold (plus 150 mesh), which was then amalgamated and assayed at the São Vicente Mine laboratory. The heavies that remained after amalgamation were assumed to contain no gold, but were captured and saved. The entire pan tailings were collected, dried, and split to a 1.8 kg sample and shipped to the Nomos Análises Minerais Ltda (“Nomos”) laboratory in Rio de Janeiro for gold fire assay using a 50 g subsample and an atomic absorption (“AA”) finish. The value obtained from this assay was termed the fine or “chemical” gold assay. Therefore, each 2 m interval has two assays, gravity and chemical, and the sum of them being the total gold grade.

The RC holes drilled by Santa Elina were also sampled at 2 m intervals. Sample collection was done using a cyclone installed at the drill rig. One of two splitting devices was set under the cyclone. A “Jones” splitter was used for dry samples, and a rotating hydraulic splitter for wet samples. Both splitters provided for 100% capture of the cuttings with division of the sample product handled automatically. A portion of this sample was dried and sent to the Nomos laboratory and a portion retained in a secure facility on site.

A quality control and quality assurance (“QA/QC”) program was carried out for Santa Elina by Bondar Clegg Laboratories Ltd (“Bondar Clegg”) in Vancouver and Geolab (M) Sdn. Bhd (“GeoLab”) in Brazil. The program consisted of the introduction of blanks and standards each 26 samples, duplicates each 10 samples and checks of 5% of the pulps at a second lab. Four of the standards were prepared by Bondar Clegg in a round robin process with material from Santa Elina’s Fazenda Nova project. Mr. Lovstrom (a sampling consultant) concluded at the end of the program in January 1998 that the assay data was acceptable for use in mineral resource estimation.

From 2003 to 2008, samples were collected, prepared and analyzed by Yamana using stringent protocols, as established by Francis Pitard, a sampling consultant from Broomfield, Colorado. The core was photographed in its entirety prior to splitting or sampling and then logged in detail. Samples were collected at 2 m intervals continuously for the full length of the core. The NQ (47.6 mm diameter) core drilled during the initial program was submitted in its entirety for sample analysis to maximize sample size, approximately 8 kg per each 2 m sample, and reduce the coarse gold nugget effect which can occur with small samples.

Later, drilling was done using the larger HQ (63.5 mm diameter) sized core. The core was sawn in half using a diamond rock saw and one-half of the cores submitted for sample analysis and the other half retained for future reference and stored in a secure facility. A 2 m interval sample of HQ half-core yields a sample weighing approximately the same as a 2 m interval sample of NQ whole-core (i.e., 8 kg). The 2 m core samples were carefully labelled using replicate tear-off control tags and sealed in secure plastic bags for shipment to the independent analytical laboratories. Blank samples and standard samples were inserted for QA/QC purposes for every 20 samples submitted for analysis. All sampling and bagging was personally carried out or closely supervised by a company geologist. Samples were then transported by a commercial trucking contractor directly to the Laboratory Lakefield-Geosol “Lakefield” or Société Générale de Surveillance do Brasil Ltda “SGS”, both independent certified analytical laboratories located at Belo Horizonte, Brazil.

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NI 43-101 Technical Report

Samples were assayed using either fire assay or screen fire assay methods depending on the degree of hydrothermal alteration. Fire assays were completed on samples from areas with low to medium alteration and screen fire assays were carried out on samples having strong hydrothermal alteration. The samples were dried at 110° C and crushed with a jaw crusher so that 95% passed a 10 mesh screen. A 1 kg subsample was then taken using a riffle splitter and pulverized using a ring mill so that more than 95% passed a 150 mesh screen. Standard fire assay methods using a 50 g pulp were then used to determine the gold content.

For the screen fire assays, each sample weighing approximately 8 kg is crushed and pulverized to 150 mesh and then sifted through a 100 mesh plastic sieve. The coarse fraction generally weighs between 30 and 50 g. In this case, the coarse fraction is analyzed in total and the fine fraction is assayed in triplicate after homogenization. Then standard fire assay methods are used to determine total gold content.

Details of the sample security procedures for the pre-2003 sample preparation, analysis, and transportation are no longer available; however the remaining drill core from the São Francisco Mine is stored at a secure facility at Apoena’s São Vicente Mine, as are the sample pulps and rejects.

For samples from 2003 and onward, the sampling and bagging was carried out or closely supervised by a Yamana geologist. Samples were then transported by a commercial trucking contractor directly to the independent certified analytical laboratories located in or near Belo Horizonte in Minas Gerais State, Brazil. Upon receipt of the samples, the laboratories were instructed to report any apparent tampering or inconsistencies in sample numbering to the company; however, no discrepancies or tampering were encountered.

Apoena conducted a program of exploration through 2009-2011 aiming to check the continuity of the mineralization near the mine. Seventeen drill holes were drilled totalling 3,070 m. Apoena continued to drill and further 43 drill holes were drilled totalling 5,144 m that were unavailable for the September 30, 2011 Mineral Resource and Mineral Reserve estimates.

1.5

Metallurgical Testing

Pilot plant scale column leach testing was performed by Santa Elina Desenvolvimento Mineral S.A “SEDM” personnel at the São Vicente Mine facilities. The resulting gold extractions ranged from 43% to 80%.

Bottle roll leach tests were performed by KCA in 2004 and the results indicate all ore types are amenable to cyanide leaching with moderate to high recoveries for the ore types tested.

Pilot gravity plant testing was performed in 2004. The main conclusions from the pilot plant tests are that about 30% of total gold content is recoverable in a gravity plant. The gold remaining in the minus 6mm material and the plus 6mm material requires heap leaching methods for recovery.

Samples of the minus 6 mm gravity tailings and the plus 6 mm screened material from the three special gravity pilot plant tests were sent to KCA for column leach testing. From the column leach test results KCA projected the ultimate gold recoveries to range from 43% to 84% based on an extrapolation of the data for columns that not yet reached equilibrium conditions.

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1.6

Mineral Resources

The Mineral Resource for the São Francisco Mine has been estimated by Aura using ordinary kriging and a geological model that defines the high-grade hydrothermal altered zone (“HGZ”), the low-grade hydrothermal alteration zone (“LGZ”) and the saprolite zone that consists of weathered mineralized material near the surface. The block model used blocks measuring 10 x 10 x 10 m. The drill hole data was composited to 5 m intervals and outlier gold grades were not capped at 15 Au g/t. The estimation search strategy used three passes and the search ellipsoid was oriented to align with the variograms. A minimum of two and a maximum of 15 composites were required to make a block grade estimate.

The block model was then updated using the September 30, 2011 topography to account for previously mined material and has been limited by an optimized pit using a $1,700 per ounce gold price to estimate Mineral Resources.

1.7

Mineral Resources Summary

The September 30, 2011 Mineral Resource is estimated by Aura Minerals to be 10.923 million tonnes of Indicated Mineral Resource at an average grade of approximately 0.95 g/t Au and Inferred Resource of 0.083 million tonnes at an average grade of 0.47 g/t Au, using a long term $1,700 per ounce gold price, and a 0.23 g/t Au cutoff (Table 1-1). Note that the extent of the September 30, 2011 Mineral Resource has been limited by an optimized pit using a $1,700 per ounce gold price. Note that the Mineral Resources are inclusive of Mineral Reserves. Also note that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

Table 1-1 – September 30, 2011 Mineral Resource Estimate

Resource Category	Tonnes (‘000)	Au (g/t)	Oz (‘000)
Measured	0	0	0
Indicated	10,923	0.95	334
Measured + Indicated	10,923	0.95	334
Inferred	83	0.47	1

* At 0.23 Au (g/t) cutoff – Note numbers may not add due to rounding.

1.8

Mineral Reserve

The estimate of Mineral Reserves is based on a life of mine plan (“LOM Plan”) and plant production schedule developed by Aura. The Economic criteria using the Lerch-Grossman algorithm for pit limit evaluations, including process recoveries and operating costs are provided in Table 1-2. Industry standard pit optimization followed by detailed design work, set the limits of the final pit to account for practical mining.

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Table 1-2 - Pit and Cost Parameters

Pit and Cost Parameters
Bench height (m)	10
Road width(m)	15
Overall Pit Slope (°)	52
Bench face angle (°)	80
Minimum pit bottom(m)	10
Berm width (m) (Alternating	10 / 4.3
Ramp slope (%)	10
Gold Price(US$/oz)	1500
Gold Recovery (%)	80
Mining cost (US$/t)	3.34
CGO Plant Cost (US$/t)	7.32
G&A cost (US$/t)	2.76

The September 30, 2011 Mineral Reserve is estimated by Aura to be 10,890 million tonnes of Probable Mineral Reserve at an average grade of 0.91 g/t Au. The Mineral Reserve has been estimated using a 0.25 g/t Au cutoff. The authors note that the Mineral Resource and Mineral Reserve are estimated in accordance with the CIM definitions and are considered NI 43-101 compliant.

1.9

Mineral Reserve Summary

The Mineral Reserve estimate is presented in Table 1-3.

Table 1-3 – September 30, 2011 Mineral Reserve Estimate

Ore ^*	Tonne (t)'000	Au (g/t)	oz'000
Proven	-	-	-
Probable	10,890	0.91	318
Proven + Probable	10,890	0.91	318
Waste	36,512
W/O	3.35

*Reserves at a 0.25 Au (g/t) cutoff– Note numbers may not add due to rounding.

1.10

Mining Methods

The São Francisco Mine is an open-pit operation that is designed to send a steady feed of crushed gravity leach ore (“CGO”), on a daily basis, to the gravity circuit while lower grade material that was the run-of-mine dump leach ore (“DLO”) is now being stockpiled and will be run through the plant at the end of the mine life. Low grade ore and sub marginal material has replaced DLO starting in January 2011. CGO has a 0.40 g/t cutoff, low grade ore (LGR) is 0.25 g/t and sub marginal grade (SMG) has a 0.15 g/t cutoff. The LGR and SMG material will be stockpiled and fed through as CGO at the end of the mine life if economic.

Mining at the São Francisco Mine is carried out by contractors using a combination of hydraulic excavators, front end loaders, and 35 t haulage trucks with typical pit support equipment to sustain the operation. The pit is accessed via a standard access road or ramp 15 m wide at a minus 10% grade which provides access to the benches. Benches are 10 m high, with alternating 4.3 and 10 m wide safety berms.

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Production drilling is carried out with 5.5 inch diameter holes and patterns for waste and ore of 6 x 4 m and 3.5 x 4.5 m, respectively. Standard bench height is 10 m, with one meter of sub-grade drilling for floor control. The bottom seven meters of the blast holes are loaded with emulsion type explosives, and a three meters stemmed collar is left at the top of the hole.

1.11

Recovery Methods

The São Francisco Mine’s mineral processing system consists of a gravity concentration circuit combined with a multi-lift, heap leach operation. During 2011, the two categories of lower grade material were stockpiled separately. Sub-marginal grade is material between 0.15 g/t Au to 0.25 g/t Au, low grade is material between 0.25 g/t Au to 0.40 g/t Au. CGO or high grade is material above 0.40 g/t Au. The CGO is processed by crushing and gravity concentration, followed by heap leaching of the tailings product from the circuit.

The crushing circuit consists of three stages of crushing and has been designed to generate a product of 30% passing less than 3.5 mm to feed the gravity circuit. The remaining 70% is directed to the leach pads with the gravity tails. Low grade ore and sub marginal grade material ore is being stockpiled and will be processed at the end of the mine life, if economic, by treating as CGO.

At the heap leach area, CGO is stacked using trucks and track dozers. The leach pad is a multi-lift, single-use type pad. The basic components of the leach pad include a compacted earth foundation (a compacted soil sub-base), a 1.5 mm thick geo-membrane liner, a cushion layer of material on top of the geo-membrane, and a series of perforated plastic drainpipes, which are placed on top of the geo-membrane, under the cushion layer during stacking, to collect leach solutions. Following stacking of the CGO, the ore is irrigated with leach solution and the resulting gold-bearing solutions are collected in a pregnant solution pond prior to further processing for recovery of gold.

A carbon Adsorption/Desorption/Recovery (“ADR”) facility is used for gold recovery. The ADR facility includes a train of five, cascade-type, open-top up-flow carbon adsorption columns, a carbon desorption system, a carbon acid wash circuit, a precious metals recovery circuit that utilizes electro-winning, and a complete smelting system for gold production. Gravity plant concentrates and the loaded cathode material from the carbon recovery plant are combined and smelted to produce a gold doré product.

1.12

Project Infrastructure

The existing Mine camp facilities include lodges, administrative offices, medical clinic, restaurant, and a gravel air strip.

National electrical service is available from Pontes e Lacerda approximately 50 km from the São Francisco Mine. Sufficient water to supply the mining operations is available at the Mine from streams and groundwater.

1.13

Environmental Studies, Permitting and Social Impact

The São Francisco Mine is designed to prevent short-term environmental damage and to minimize long-term environmental effects. The environmental and reclamation programs were developed to comply with Brazilian regulatory requirements and to restore the property to productive use after production has ceased.

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The closure and reclamation plan proposes that the open pit remains as a permanent feature, but it will have berms installed to prevent public access. At the end of the operation, the pit will be transformed into a lake depending on the outcome of the appropriate studies. The waste rock stockpiles will be constructed in lifts at their natural angle of repose, with setbacks between lifts and the overall waste dump slopes will be approximately 2.2 to 1.0 (horizontal to vertical) and will remain at this slope at closure. The horizontal surfaces will be re-vegetated, and natural re-vegetation will proceed on the slopes.

The ore processing plants and support facilities will be removed and the heap leach pads will be neutralized, re-contoured, and the land re-vegetated. The concrete pads and foundations will be removed and the roads and other operating surfaces will be reclaimed and re-vegetated. Reclamation activities will be initiated as soon as portions of the Mine are no longer required. The current Mine budget includes approximately $11.3 million for reclamation and closure.

The São Francisco Mine operation has a valid environmental license. There are no current environmental liabilities or non-compliance issues for the property.

1.14

Capital and Operating Costs

The sustaining capital expenditures (“CAPEX”) are estimated by Aura to be US$ 15.9 million, as shown in Table 1-4. The 2011 unit operating cost breakdown for the São Francisco Mine is shown in Table 1-5.

Table 1-4 –Capital Expenditures

Area (US$ 000)	2012	2013	2014	2015	Total
Mine Equipment , dewatering	600	600	150		1,380
Plant Sustaining Projects	800	800	200		1,840
ADR Sustaining capital	600	600	150		1,380
Reclamation and Closure			8,365	2,910	11,275
Total Capital	2,000	2,000	8,865	2,910	15,875

Table 1-5 - 2011 Operation Cost Breakdown

Area	Unit Costs
Mine (US$/t moved)
Drilling	0.39
Blasting	0.35
Haulage/Loading/Auxiliary (Contractor)	2.09
Engineering and Operations Management	0.11
Total Mine (US$/t moved)	2.94
Process ($/t processed)
Crushing and Stacking	2.58
Gravity	0.81
Leaching / Cyanidation	4.11
ADR/Electrolysis/Smelting	1.65
Total ($t processed)	9.15
G&A ($/t processed)	2.77

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1.15

Economic Analysis

A post-tax Cash Flow Model has been developed for the LOM production schedule and capital and operating cost estimates. For all periods the after tax cash flow is positive, supporting the reserve designation. A sensitivity analysis has been completed that examined gold price, and operating costs. None of the sensitivities show a negative net present value.

1.16

Interpretations, Conclusions and Recommendation

Aura has prepared a technical report compliant with NI 43-101 on the Mineral Resources and Mineral Reserves pertaining to its São Francisco Mine, located in Mato Grosso State in Brazil. The São Francisco Mine is an open-pit, heap leach gold mine that involves two separate gold recovery processes including crushing-gravity gold recovery-heap leach and run-of-mine heap leach up until the start of 2011. As of early 2011, the mine focused on the more profitable crushing-gravity gold recovery-heap leach process than run-of-mine heap leach. The ore contains a significant component of gravity gold, which requires detailed sampling and attention to mine planning to ensure that the gravity gold is recovered prior to placement of ore on the leach pad.

The resource model was updated from the March 2011 estimate since the reconciliation between the model and production was poor. The current model developed by Aura shows better confidence for predictability. Subsequently operating costs and design parameters were updated and a new mine plan was developed. The results show that a downgrade of reserves and resources were necessary over previous estimates, but the new mine plan is a better indicator of future production.

The economic results show that the deposit can be mined at a profit. Significant risks to the project’s economics, along with mitigation measures include:

Mechanical breakdown of plant equipment – unforeseen failure is always a risk, but through maintenance programs, and the availability of critical spares, this risk can be managed.

Continued resource model reconciliation issues – not achieving predicted grade and tonnes poses a risk. Monthly updates are ongoing to analyze and react to any potential problems. As the pit deepens, the grade of ore mined has been forecasted to improve.

Availability of consumables – Mine management must continually be aware of any shortfalls of consumables such as cyanide and explosives and quickly source other suppliers.

Is it recommended that as mining progresses, continued reconciliation needs to be reviewed and if parameters change, an update of the mine plan should be developed.

It is also recommended that the operating costs should be reviewed on a regular basis to ensure operating cutoffs are still valid.

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It is further recommended that the south east area of the open pit be reviewed to optimize the amount of recovered gold ounces. The pit was reduced in size slightly to account for a waste dump and dewatering lines.

Lastly, it is recommended that the LT model be updated to incorporate the additional drilling (43 drill holes totalling 5,144 m: holes SF-530 to SF-572) by Aura that were not available for the September 30, 2011 Mineral Resource estimate.

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2.0

INTRODUCTION

On November 10, 2011, Aura issued a press release announcing the updated Mineral Resource and Reserve Estimates for the São Francisco Mine, which had been prepared internally by Aura and its consultants, under the direction and supervision of Ivan C. Machado, M.Sc., P.E., P.Eng, Principal of TechnoMine Services, LLC, an independent Qualified Person for the purpose of NI 43-101.

On December 20, 2011, Aura was informed of Mr. Machado’s death. Mr. Machado is the only member of TechnoMine Services, LLC able to act as a Qualified Person under NI 43-101.

Given Mr. Machado’s death, the Mineral Resource and Reserve estimates for the Sao Francisco Mine was then reviewed and confirmed by J. Britt Reid, P.Eng., Executive Vice President and Chief Operating Officer of Aura, Bruce Butcher, P.Eng., Vice President, Technical Services of Aura and, Chris Keech, P.Geo., former Manager, Geostatistics of Aura (currently Principal Geologist of CGK Consulting Services Inc.) prior to preparing this Report on behalf of Aura. Messrs. Reid, Butcher and Keech are Qualified Persons for the purposes of NI 43-101.

Messrs. Reid, Butcher and Keech have visited the Mine on many occasions, but the most recent visit has been from, November 3 to 4, 2011, February 21 to 22, 2011, and October 19 to November 5, 2010, respectively.

All currency amounts are stated in US dollars, Canadian dollars or Brazilian real, as specified, with costs and commodity prices typically expressed in US dollars. Quantities are generally stated in Système International d’Unités (SI) metric units, the standard Canadian and international practice, including metric tons (tonnes, t) and kilograms (kg) for weight, kilometres (km) or metres (m) for distance, hectares (ha) for area, grams (g) and grams per metric tonne (g/t) for gold (g/t Au). Wherever applicable, imperial units have been converted to metric units for reporting consistency. Precious metal grades may be expressed in parts per million (ppm) or parts per billion (ppb) and their quantities may also be reported in troy ounces (ounces, oz), a common practice in the mining industry. Attached hereto as Appendix 1 is a glossary of common terms and abbreviations that may be used in this Report.

This Report follows the format and guidelines of Form 43-101F1, Technical Report for National Instrument 43-101, Standards of Disclosure for Mineral Projects, and its Companion Policy NI 43-101CP, as amended by the Canadian Securities Administrators and which came into force on June 30, 2011. The June 30, 2011 format and guidelines of Form 43-101F1 and its Companion Policy NI 43-101CP replace the former format, guidelines and companion policy which was dated December 23, 2005.

This Report includes technical information which requires subsequent calculations or estimates to derive sub-totals, totals and weighted averages. Such calculations or estimations inherently involve a degree of rounding and consequently introduce a margin of error. Where these occur, Aura does not consider them to be material. Further, the Report summarizes the professional opinion of the authors and includes conclusions and estimates that have been based on professional judgment and reasonable care.

Said conclusions and estimates are consistent with the level of detail of this study and based on the information available at the time this Report was completed. All conclusions and estimates presented are based on the assumptions and conditions outlined in this Report.

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This Report is to be issued and read in its entirety. Written or verbal excerpts from this Report may not be used without the express written consent of the authors or officers of Aura.

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3.0

RELIANCE ON OTHER EXPERTS

The Report was prepared by Aura. The information, conclusions, opinions, and estimates contained herein are based on:

Data, reports and other Information made available at the time of preparation of the Report; and

Assumptions, conditions, and qualifications as set forth in the Report.

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4.0

PROPERTY DESCRIPTION AND LOCATION

4.1

Location

The São Francisco property is located close to the Bolivian border in the western portion of the State of Mato Grosso in west central Brazil (Figure 4-1). It is approximately 560 km west of Cuiabá, the state’s capital. Coordinates for the Mine are approximately 14°50’S latitude and 59°37’W longitude.

Figure 4-1 - Location Map from Rezende Eng. Minas S/S Ltda. November 2006

Figure 4-2 shows the location of the mine workings, on-site accommodations, recreation facilities, and cafeteria for the workforce.

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Figure 4-2 - Site Location from Rogerio Moreno, 2011

4.2

Review of Mineral Properties and Titles in Brazil

4.2.1

Mineral and Mining Rights Tenure

The São Francisco property consists of two contiguous mining permits and two adjacent exploration claim applications covering some 36,308.34 ha. The two mining permits held by Apoena were granted in 1982 by the DNPM and the process numbers for the permits are 860938/1982 and 860937/1982. All these areas are shown in Figure 4-3.

The São Francisco Mine operation has a valid environmental license. There are no current environmental liabilities or non-compliance issues for the property.

Table 4-1 - DNPM Permits

DNPM Process Number	Area	DNPM Status
860937/1982	7,239.18	Mining Concession
860938/1982	9,129.16	Mining Concession
861146/1993	10,000.00	Exploration Concession
861609/1994	9,940.00	Exploration Concession

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Figure 4-3 -Concession Map (from Scott Wilson RPA)

4.2.2

Surface Rights

All areas encompassing the Mine infrastructure and mineral resources were acquired by Aura.

4.3

Royalties

Apoena is paying the Rural Property Tax (“ITR”). There is a 1% royalty payment to the government on metal sales.

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4.4

Environmental Permits, Studies, and Considerations

Since the beginning of the operation, some environmental impact and mitigation activities have been developed for the Mine such as re-vegetating historically mined areas and water quality monitoring.

Current environmental impact mitigation for the Mine includes: dust suppression; speed control of vehicles; segregating and stockpiling of topsoil for future reclamation activities; depositing waste rock by end dumping around the pit limits and close to the plant site; storage of jig tailings in reservoirs (B5 to B7 reservoirs); continuous monitoring of surface and groundwater; introduction of a solid waste management system; and a sewage and water treatment system.

For the heap leach operation a synthetic membrane has been installed at the base of the pad to collect the pregnant solution and to avoid groundwater contamination and a system installed along the ponds to avoid wildlife mortality.

For the São Francisco Mine closure plan, environmental reclamation is the most significant activity including areas such as: the open pit, waste dump, heap leach pads, dumps leach pads, tailing ponds and others. The total area to be recovered is 195.35 ha.

The São Francisco Mine operation has a valid environmental license. There are no current environmental liabilities or non-compliance issues for the property.

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5.0

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY

5.1

Physiography

The physiography of the São Francisco Mine is characterized by a mountain range, part of the Aguapeí Mobile Belt and Mafic Arc, which follows the Brazil-Bolivia border. The range, in the vicinity of the Mine, forms a prominent ridge some 800 m in elevation that strikes approximately N30°W and is approximately 20 km wide. The ridge stands out from the plains with a gentle slope on the western side towards Bolivia with the east side forming a prominent vertical, to near vertical, cliff-like escarpment that extends for more than 200 km within the Aguapeí Mobile Belt. The São Francisco deposit occurs on gentle slopes of the São Vicente Ridge, a subdued hill surrounded by a broad, flat plain drained by a few major creeks.

Vegetation consists of mixed forest, savannah, and open fields. The tailings storage and waste disposal sites at the São Francisco Mine are on the flat plain adjacent to São Vicente Ridge and offer ample room to meet the needs of the mining operations as well as ample room for the heap leach pads. Sufficient surface and groundwater is available to supply the mining and processing operations.

5.2

Accessibility

The São Francisco Mine is accessible from Cuiabá by road (BR-070/174) or by chartered flights to an airstrip at the Mine. There are several daily commercial flights linking Cuiabá to major Brazilian cities. The town of Pontes e Lacerda with a population of some 50,000 inhabitants is the closest full service community where modern education, medical, shopping, and banking facilities are available. The São Francisco Mine is also accessible via a 58 km gravel road from the São Vicente Mine. This road has been upgraded to be serviceable on a year-round basis.

5.3

Climate

The climate at the Mine site is tropical to semi-tropical, typical of the central part of Brazil. The summers are hot and rainy with daily maximum temperatures in the range of 30° C to 35° C. The winters are cooler and dryer with daily maximum temperatures in the range of 16° C to 20° C. The annual average temperature is 24° C, with the maximum and minimum absolute temperatures of 40° C and 0° C, respectively. The summer rainy period starts in December and continues through to March. The dry season is from May to September. The total annual precipitation ranges from 1,300 mm to 2,400 mm, with an annual average of 1,700 mm. The Mine operates year round in this climate.

5.4

Local Resources and Infrastructure

The São Francisco Mine is located within the municipality of Vila Bela da Santíssima Trindade. A total of four communities are close-by the Mine. They are Vila Bela da Santíssima Trindade, Nova Lacerda, Conquista d’Oeste and Pontes e Lacerda. The town of Pontes e Lacerda with a population of some 50,000 inhabitants is the closest full service community where modern education, medical, shopping and banking facilities are available.

There is a readily available local labour supply from nearby towns. More specialized mining technical personnel are brought in to site from regional Brazil, and work on a rotation basis.

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The existing Mine camp facilities include lodges, administrative offices, medical clinic, restaurant, and a gravel air strip.

Drinking water is brought in as bottled water from the region. Shower water is pumped from a local freshwater dam.

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6.0

HISTORY

Gold was first discovered in the area in the 1700s and mining at São Francisco was initiated by the Portuguese. The area became the first significant gold mining district in Brazil, and the nearby settlement, Vila Bela, was at that time named the capital of Mato Grosso State. The district reportedly produced and shipped some 60 to 70 tonnes of gold to Portugal between 1720 and 1830. A portion of the São Francisco concession area has been set aside as a national archaeological site to the south west of the pit, where mining activities are not permitted.

6.1

Previous and Current Owners

In the mid-1970s garimpeiros began mining activities in the area. In 1977, Santa Elina began acquiring mineral concessions and in 1983, commenced a dredging/placer operation at São Vicente. From 1995 to 1997, Santa Elina used a flotation and gravity concentration operation at São Vicente and São Francisco mines. A total of some 186,810 ounces of gold were reportedly produced by these two operations. The São Vicente operation was shut down in 1997 due to high operating costs.

In 2002, Santa Elina re-examined prefeasibility studies completed in 1996 and 1997 based on reduced costs due to currency devaluation. In 2003, Yamana obtained control of the São Francisco Mine as part of a reverse takeover transaction. Watts, Griffis, and McQuat Limited (“WGM”) conducted a review of the property and produced a preliminary feasibility study. In 2004, a shaft and three trenches were excavated to obtain bulk samples for metallurgical testing, including large column tests and pilot-plant sized gravity concentration tests.

From 2006 to September 30, 2011, the São Francisco Mine has processed some 38.5 million tonnes grading an average of 0.55 g/t Au and has recovered more than 438,000 oz of gold. A 100% interest in the São Francisco Mine was acquired by Apoena in May 2010.

6.2

Exploration and Drilling

Historical drilling programs are outlined in Table 6-1 below:

Table 6-1 – Exploration Program

Year	Company	Work Done	Metres	Objective
1990	Santa Elina	Diamond Drilling	1,102.4	Exploration
1993 - 1995	Santa Elina	Diamond Drilling	30,239.6	Infill Drilling
1996	Santa Elina/ Echo Bay	Diamond Drilling	6,649.3	Infill, Exploration, Metallurgical Drilling
1997	Santa Elina/ Echo Bay	Diamond/RC Drilling	12,199.1	Infill Drilling
1997 - 1998	Santa Elina/ Echo Bay	Metallurgical Testing		Heap Leach Viability
2003 - 2004	Yamana	Diamond Drilling	15,029.9	Targeted Infill Drilling
		Bulk Sampling/Pilot Plant		Establish Grade and Continuity
		Pre-feasibility Study		Establish Mineral Reserves
2005	Yamana	Diamond Drilling	10,232.4	Infill/Upgrade Drilling
2006	Yamana	Diamond Drilling	4,620.3	Define Zone Extensions
2007	Yamana	Diamond Drilling	7,684.1	Infill/Deep Exploration
2008	Yamana	Diamond Drilling	2,823.3	Infill/Deep Exploration
2010/11	Apoena	Diamond Drilling	3,070.77	Define Zone Extensions
Total	Drilling		93,666.46

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6.3

Previous Mineral Resource and Reserve Estimates

Aura reported a December 31, 2010 Mineral Resource and Mineral Reserve estimates for São Francisco deposits in their NI 43-101 Report dated March 30, 2011.

The December 31, 2010 Mineral Resource estimates were classified as Measured, Indicated and Inferred, were based on a cutoff grade of 0.15 g/t Au, was restricted by a pit using a US$1,200 per ounce gold price (Table 6-2).

Table 6-2 – December 31, 2010 Mineral Resource Estimate

Class	Tonnes (t)	Au (g/t)	Ounces ‘000
Measured	-	-	-
Indicated	21,631	1.01	706
Measured + Indicated	21,631	1.01	706
Inferred	62	3.05	6

*cutoff grade 0.15 g/t Au – Note numbers may not add up due to rounding.

The December 31, 2010 Mineral Reserve estimates were classified as Proven and Probable and were based on a cutoff grade of 0.30 g/t. The Mineral Reserve estimates have been estimated based on a US$1,025 per ounce gold price (Table 6-2).

Table 6-3 – December 31, 2010 Mineral Reserve Estimate

Ore*	Tonne (t)'000	Au (g/t)	Ounces ‘000
Proven	-	-	-
Probable	13,731	1.29	571
Proven + Probable	13,731	1.29	571

*cutoff grade 0.30 g/t Au – Note numbers may not add up due to rounding

6.4

Historical Production

Table 6-4 presents a summary of the gold production from the São Francisco Mine.

Table 6-4 - Historic and Recent Production

Year	Ore Leached (tonnes)	Grade (g/t Au)	Gold Recovered (Oz)
2006	10,359,263	0.58	81,381
2007	7,469,099	0.54	104,763
2008	7,430,903	0.53	75,937
2009	6,525,000	0.58	82,665
2010	4,555,368	0.47	61,770
2011 to Sept 30,2011	2,168,303	0.61	31,542

Note: numbers may not add up due to rounding.

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7.0

GEOLOGICAL SETTING AND MINERALIZATION

7.1

Regional Geology

The regional geological setting for the São Francisco Mine, and numerous other gold occurrences that comprise the Guapore Gold Belt of central west Brazil and east central Bolivia, is the Aguapeí Mobile and Mafic Arc Belt. This belt follows a major crustal-scale break or shear zone separating the Amazon Craton of Archean-age (3,800 to 2,500 million years ago) on the east from the Paragua Craton of Proterozoic-age (2,500 to 1,600 million years ago) on the west. The belt extends more than 600 km in a NNW-SSE direction (Figure 7-1) and is characterized by a prominent mountain range made up of a 1,200 m thick sequence of Proterozoic-age sedimentary rocks known as the Aguapeí Group (Figure 7-2).

The Aguapeí Group, the host rocks for the gold mineralization, is a sequence of texturally and mineralogically supermature sediments made up of braided river facies, Aeolian (wind-deposited) dunes facies and shallow marine platform facies. Southward along the belt, the lower part of the Aguapeí Group contains interbedded volcanic units and mafic sills and dykes (that may be thrust from the east).

The Aguapeí Group unconformably overlies the central part of the Amazon Craton (Brazilian Precambrian Shield). This part of the shield, known as the Xingu Complex, contains lower Proterozoic volcano-sedimentary belts elongated in a NW-SE direction and are surrounded by Archean gneissic to migmatitic masses. Both the volcano-sedimentary belts and the Xingu Complex have been intruded by granitic batholiths. Prolonged erosion of the older rocks was accompanied by the development of basins into which the Aguapeí sediments were deposited. The flat area surrounding the mountain range is believed to be mainly underlain by the Xingu Complex, however, most of the area is covered by alluvial-lateritic soils with few outcrops and the geology is not well known. Prolonged and deep erosion of this continental mass occurred during Proterozoic time, and was accompanied by the development of intracratonic basins in which the 1,200 m thick Aguapeí Group sediments were deposited. The sediments are comprised of predominantly meta-arenites with lesser amounts of metapelites and even less common lenses of metaconglomerate. The Aguapeí Group has been mapped over a 300 km strike length in Brazil and continues southwards into Bolivia for more than 200 km (where it is known as the Sunsas Group) before passing again into Brazil.

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Figure 7-1 - Regional Geology from Rezende Eng. Minas S/S Ltda. November 2006

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Figure 7-2 - Regional Stratigraphic Column from Rezende Eng. Minas S/S Ltda. November 2006

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The mountain range is bounded on both sides by faults, with the fault on the east dipping away from the range at a shallow angle. This fault separates the Archean basement on the east from the Proterozoic metamorphosed sediments on the west and is thought to be a thrust fault in the vicinity of São Vicente to Pontes e Lacerda. A well-developed northeast-southwest fracture system also cross cuts the regional trend.

The most deformed parts of the Aguapeí Group form ridges such as the São Vicente, de Borda, Patrimônio, Cágado, Caldeirão, Pau-a-Pique, Aguapeí and Caramujo. The internal part of the Brazilian Aguapeí volcano-sedimentary belts contains extensive plateaus of Aguapeí Group metasediments that show little or no deformation, for example, the ridges of Ricardo Franco/Santa Bárbara in Brazil.

The last orogenic stage was accompanied by mafic igneous and granitoid phases including the generation of pegmatites.

The known bedrock gold deposits and occurrences in Mato Grosso state are separated into two districts – the São Vicente/Borda district and the Pontes e Lacerda district (to the south of São Francisco).

7.2

Local Geology

The São Francisco deposit occurs in the Fortuna Formation (the basal unit of the Aguapeí Group) which is composed of fine to coarse grained meta-arenites, with locally reddish metapelites and occasionally oligomictic metaconglomerates. The metaconglomerates are composed of quartz pebbles and quartz grains set in a siliceous-arenaceous matrix. Primary sedimentary structures such as cross-bedding, graded bedding and ripple marks are commonly observed.

Compressional forces from the NE and SW resulted in folding, faulting, shearing, and fracturing of the rocks and produced a series of broad anticlines and synclines, which are traceable over several kilometres.

The hydrothermal alteration zone (“HAZ”) that contains the mineralization is a relatively regular, steeply dipping tabular body which has not been defined by drilling at depth. Observed alteration in the HAZ consists of silicification with occasional sericite and chlorite alteration.

The São Francisco geological model defines three main zones: 1) a saprolite zone; 2) a higher grade zone that encloses the core of the deposit starting below the saprolite and continuing down dip; and 3) the overall HAZ where erratic and mostly lower grade mineralization exists (although there are significant areas with higher grades). The richest mineralization, occurring within the HAZ, is fairly continuous along the NW strike, plunges at a low angle towards the NW, and dips about 65° to the NE.

Gold bearing zones of alteration and mineralization can be visually identified in the drill core and the pit. The zones are characterized by a combination of quartz veins and pyrite (either fresh or oxidized/leached). An increase in the concentration of the quartz veins and pyrite is directly is associated with an increase in the concentration of visible gold (as nuggets) and the consequent increase in gold grade. The orange-to-red colour acquired by the host rock in the areas of larger sulphide concentration is another distinctive feature of the higher grade zones.

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7.3

Mineralization

Gold mineralization at the Mine occurs in epigenetic, quartz-filled shear zones generally along the foliation that is oriented parallel to the axis of the anticlinal-synclinal folds and in later, flat to shallow dipping quartz veins and in-filled shear zones that cut the primary bedding of the host rocks. Gold occurs as free gold, frequently as coarse nuggets measuring several millimetres in diameter with the quartz, as laminations along the fracture planes, and within limonite boxwork after pyrite and arsenopyrite. Pyrite and arsenopyrite commonly occur at depth and form the gangue minerals in the unweathered sulphide horizons. A significant “nugget” effect is present due to the coarse visible gold.

The highest elevation of the mineralization is 800 m (above mean sea level) and the lowest is 450 m (above mean sea level). The deposit is approximately 1,800 m long, 150 m wide and 350 m deep and is still open at depth.

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8.0

DEPOSIT TYPES

The São Francisco Mine is a shear hosted lode gold deposit. It is epigenetic, structurally controlled, and composed of narrow, 1 to 5 centimetres wide, and quartz veins containing free gold. The veins and vein systems and stockworks both parallel and crosscut the bedding planes and appear to represent separate but closely related mineralizing events.

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9.0

EXPLORATION

Exploration work was conducted at the São Francisco Mine between 1985 and 1997 by Santa Elina or companies working for or with Santa Elina. There were no exploration activities from 1998 to late 2003. From late 2003 to 2008 Yamana carried out additional diamond drilling, surface trenching, underground bulk sampling, and completed a pre-feasibility study. A summary of the Yamana activities is described below.

In November 2003, Yamana drilled 49 core holes (44 were vertical and 5 were inclined) totalling 15,030 m along the strike of the mineralization to collect material for metallurgical testing. Yamana also conducted an underground exploration program that consisted of a 150 m vertical shaft and 107 m of underground drifts. As well, bulk samples were taken from nine surface trench sites (each about 100 m long x 4 m wide x 2.5 m deep across the strike of the deposit) and the bulk samples ranged in weight from 265 tonnes to 2,208 tonnes.

In 2005, Yamana drilled 104 core holes totalling, 10,231 m to upgrade Inferred Mineral Resources to Indicated Mineral Resources and create the basis for a Mineral Reserve estimate.

In 2006, Yamana drilled 20 core holes totalling 4,620 m to check the continuity of the gold mineralization along the open strike extensions of the deposit to the NW and SE.

In 2007, a total of 7,684 m were drilled in 18 core holes by Yamana. Three of these holes were drilled to investigate the extension of the deposit at depth, and the others being infill holes.

In 2008, Yamana drilled 9 core holes totalling 2,823 m to infill the depth extension of higher-grade mineralization at the SE and NW ends of the deposit.

Apoena conducted a program of exploration through 2009-2011 aiming to check the continuity of the mineralization near the mine. Seventeen drill holes were drilled totalling 3,070 m.

By March 5, 2011, the effective date of the drill hole database, the drilling at the Mine totalled 93,666 m in 524 drill holes. The exploration activities carried out on the property are listed in Table 9-1 below. Activities carried out by Santa Elina and Yamana are taken from Mello and Petter (2008).

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Table 9-1 - Exploration Program

Year	Company	Work Done	Metres	Objective
1990	Santa Elina	Diamond Drilling	1,102.4	Exploration
1993 - 1995	Santa Elina	Diamond Drilling	30,239.6	Infill Drilling
1996	Santa Elina/ Echo Bay	Diamond Drilling	6,649.3	Infill, Exploration, Metallurgical Drilling
1997	Santa Elina/ Echo Bay	Diamond/RC Drilling	12,199.1	Infill Drilling
1997 - 1998	Santa Elina/ Echo Bay	Metallurgical Testing		Heap Leach Viability
2003 - 2004	Yamana	Diamond Drilling	15,029.9	Targeted Infill Drilling
		Bulk Sampling/Pilot Plant		Establish Grade and Continuity
		Pre-feasibility Study		Establish Mineral Reserves
2005	Yamana	Diamond Drilling	10,232.4	Infill/Upgrade Drilling
2006	Yamana	Diamond Drilling	4,620.3	Define Zone Extensions
2007	Yamana	Diamond Drilling	7,684.1	Infill/Deep Exploration
2008	Yamana	Diamond Drilling	2,823.3	Infill/Deep Exploration
2010/11	Apoena	Diamond Drilling	3,070.77	Define Zone Extensions
Total	Drilling		93,666.46

9.1

Sampling Method and Quality

Concerns about correctly sampling the coarse visible gold (nugget effect) in the São Francisco mineralization has resulted in various sampling and sample preparation protocols being used to collect representative samples and minimize sampling error.

Santa Elina

From 1990 to 1997, core samples were taken by Santa Elina at 2 m intervals from one-half of the diamond-sawn core.

Yamana

From 2003 to 2008, samples from the Yamana drilling were collected, prepared, and analyzed using stringent protocols, as established by Francis Pitard, a sampling consultant from Broomfield, Colorado. The sampling and logging protocols used by Yamana are summarized below.

The core was placed in labelled boxes at the drill site and the boxes were transported to the logging facility. Sample intervals were then marked, generally two meters in length, but were shortened to respect geological boundaries. Sample numbers were then assigned to the intervals. In-house and certified standards and locally collected blank samples (barren material) were inserted into the sample stream. No duplicate drill samples were sent for analysis.

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The core was photographed in its entirety prior to splitting or sampling and then logged in detail. The amount of core recovered from each logged interval was recorded in the drill logs. Samples were collected at 2 m intervals continuously for the full length of the core.

Core samples were sawn in half, with one half sent for assay and the other half stored on site at a secure facility. The core samples were placed in plastic bags and then in 36 litre plastic containers to be shipped to the assay laboratory.

The NQ (47.6 mm diameter) core drilled during the initial program was submitted in its entirety for sample analysis to maximize sample size, approximately 8 kg per each 2 m sample, to try and reduce the coarse gold nugget effect which can occur with small samples.

Later, drilling was done using the larger HQ (63.5 mm diameter) sized core. This core was sawn in half using a diamond rock saw and one-half of the cores submitted for sample analysis and the other half retained for future reference and stored in a secure facility. A 2 m interval sample of HQ half-core yields a sample weighting approximately the same as a 2 m interval sample of NQ whole-core (i.e., 8 kg). The 2 m core samples were carefully labelled using replicate tear-off control tags and sealed in secure plastic bags for shipment to the independent analytical laboratories. Blank samples and standard samples were inserted for control purposes for every 20 samples submitted for analysis.

All sampling and bagging was personally carried out or closely supervised by a company geologist. Samples were then transported by a commercial trucking contractor directly to the laboratory; either Lakefield-Geosol or SGS do Brasil Ltda, both independent certified analytical laboratories located at Belo Horizonte, Brazil.

No overall core recovery statistics were reviewed, but from the inspection of a number of drill logs and visual inspection of split core from several drill holes, it is estimated that there is better than 95% overall core recovery.

Apoena

The core samples collected during 2009-2011 followed the procedures described below.

Apoena used HQ size core. Core samples are placed into wood or plastic core boxes, labelled with the drill hole number and box number. The core boxes are collected daily from the drill rig and transported to the core shack by Apoena’s employees, where a log of the core is completed. The core boxes are labelled with the start and end of the interval for that box, the range of sample numbers and photographed. Geotechnical logging is carried out, including measurements of total core recovery (“TCR”), rock quality designation (“RQD”), and fracture angle and type. A logging geologist monitors the core cutting to ensure that the samples are correctly taken. Samples are typically 2 m in length with a minimum sample length of 0.5 m. The fixed 2 m length samples are not locally adjusted for lithology or alteration. However, visually unmineralized rock is left un-sampled.

In general, the core recovery is excellent, averaging 95%. There are only isolated areas that show core recovery less than 90%.

The core is cut using a diamond saw blade and half the core is placed in a numbered bag that is securely sealed. The remaining half of the core is placed back into the core box and is retained off-site in a secure storage facility. After each sample, the diamond saw is cleaned to minimize contamination between samples.

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A core twin sample is taken by quartering one half of the core and putting each quarter into different numbered bags which are then sealed. The labelled and sealed bags are put into large sacks that are sealed and the sacks are sent overland to the preparation laboratory in Belo Horizonte using secure transport.

The samples were assayed using a metallic screen fire assay methodology at the SGS laboratory in Belo Horizonte. This laboratory is ISO 9001-2000 certified (MET-1 &FAA505 procedure details can be viewed at http://www.br.sgs.com/pt_br/geochem_services.pdf).

Apoena systematically inserted twin samples, certified standard samples, and blank samples in a non-sequential order in all batches of drill core samples sent for assay. The results of the QA/QC are provided in Section 12.

The authors are of the opinion that the core logging and recording procedures are comparable to industry standards. The authors have reviewed the data and methods and agree these are appropriate for the style of mineralization at the São Francisco Mine.

9.2

Densities

A systematic program of drill core density measurements has been completed at São Francisco Mine. A total of 1,425 pieces of drill core have been wax coated and bulk densities determined using the buoyancy method from 53 drill holes spread over 29 drill sections. The average of all the density measurements is 2.61 t/m³ and average of the density measurement in the mineralized material is 2.60 t/m³. Apoena chose an appropriate bulk density value of 2.60 t/m³for the high-grade HAZ, the low-grade HAZ, and the waste rock. A value of 2.46 t/m³was assigned to the saprolite lithology.

In the authors opinion the methodology used for the collection and calculation of the bulk density measurements is appropriate for this deposit type.

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10.0

DRILLING

Apoena conducted a program of exploration through 2009-2011 aiming to check the continuity of the mineralization near the mine. A total of 17 drill holes were drilled totalling 3,070 m and the results were available for the September 30, 2011 Mineral Resource estimate.

Figure 10-1 displays the location of the drill holes completed by Apoena for the Mineral Resource estimate and Table 10-1 provides a listing of the collar locations and the drill hole lengths. The core recovery was very good and averaged approximately 95%.

Significant intersections are presented in Table 10-3. The intersections are interval lengths, as true widths have not been determined. A gold cutoff of 0.15 ppm Au and a minimum length of 4 m have been used to define the significant intersections. Figure 10-2 displays a cross section at 11,775 NW showing the significant intersections for drill hole SF-520.

The database used for the Mineral Resource estimate of September 30, 2011 contained 524 holes totalling 93,666 m of drilling. The effective data of the drill hole database is March 5, 2011. The drill hole spacing in the area of the open pit is approximately 25 m by 25 m which is considered appropriate for the style of mineralization at the São Francisco Mine.

Note that Apoena continued to drill at the Mine and completed an additional 43 drill holes totalling 5,144 metres (SF-530 to SF-572). However, the results of these additional drill holes and assays were not available to be used in the September 30, 2011 Mineral Resource estimate.

Figure 10-1 - 2008 Drill Hole Locations

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Table 10-1 – Apoena Minerals Drill Hole Collar Locations

Hole ID	Easting (m)	Northing (m)	Elevation (m)	Length (m)	Year Drilled
SF-513	214,890.9	8,356,788.4	771.5	290.2	2009
SF-514	215,435.0	8,356,005.0	720.0	227.9	2009
SF-515	215,425.3	8,356,027.8	719.9	215.0	2009
SF-516	215,407.9	8,356,078.7	720.0	239.1	2009
SF-517	215,173.3	8,356,433.3	730.0	275.0	2009
SF-518	215,234.2	8,356,180.0	700.0	167.6	2009
SF-519	215,502.6	8,355,966.5	730.3	230.0	2010
SF-520	214,782.8	8,356,887.3	760.0	275.8	2010
SF-521	214,808.3	8,356,860.4	760.0	270.2	2010
SF-522	214,922.7	8,356,703.1	752.0	306.0	2010
SF-523	214,576.6	8,356,745.6	719.9	72.5	2010
SF-524	214,953.8	8,356,213.7	741.9	60.9	2010
SF-525	214,765.7	8,356,559.2	729.3	100.4	2010
SF-526	215,357.3	8,355,899.3	719.9	90.1	2010
SF-527	215,421.6	8,355,823.1	741.7	78.5	2010
SF-528	214,974.6	8,356,230.9	730.8	61.8	2010
SF-529	215,517.0	8,355,772.2	756.3	109.8	2010

Table 10-2 - Apoena Minerals Drilling Significant Intervals 2010-2011

Hole Number	From (m)	To (m)	Length⁽¹⁾ (m)	Au⁽²⁾⁽³⁾(g/t)
SF-513	167.0	179.0	12.0	0.88
SF-513	195.0	207.0	12.0	0.33
SF-513	219.0	223.0	4.0	0.82
SF-513	270.3	278.0	7.8	0.67
SF-514	98.0	118.0	20.0	0.76
SF-514	163.0	171.0	8.0	2.29
SF-515	102.3	115.0	12.7	0.35
SF-516	98.0	128.0	30.0	1.11
SF-516	138.0	144.0	6.0	0.45
SF-517	98.0	106.0	8.0	0.51
SF-517	126.0	134.0	8.0	0.56
SF-517	206.0	228.0	22.0	0.93
SF-518	2.0	16.0	14.0	0.20
SF-518	36.0	42.0	6.0	0.22
SF-518	56.0	62.0	6.0	0.47
SF-518	72.0	78.0	6.0	1.74
SF-518	86.0	94.0	8.0	0.76
SF-518	104.0	130.0	26.0	0.24
SF-518	142.0	162.0	20.0	1.86
SF-519	110.0	134.0	24.0	1.52
SF-519	154.0	160.0	6.0	0.28
SF-520	141.0	151.0	10.0	0.17

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Hole Number	From (m)	To (m)	Length⁽¹⁾ (m)	Au⁽²⁾⁽³⁾(g/t)
SF-520	157.0	189.0	32.0	1.18
SF-520	201.0	245.0	44.0	0.61
Hole Number	From (m)	To (m)	Length⁽¹⁾ (m)	Au⁽²⁾⁽³⁾(g/t)
SF-521	156.0	168.0	12.0	0.46
SF-521	202.0	214.0	12.0	1.31
SF-522	100.0	112.0	12.0	0.21
SF-522	140.0	150.0	10.0	0.74
SF-522	208.0	218.0	10.0	0.82
SF-522	236.0	272.0	36.0	0.74
SF-522	292.0	302.0	10.0	3.74
SF-523	8.0	12.0	4.0	0.33
SF-528	0.0	10.0	10.0	0.20
SF-529	0.0	20.0	20.0	0.17
SF-529	68.0	72.0	4.0	1.03
(1) True width has yet to be determined.
(2) Fire assay and atomic absorption finish.
(3) Assays not capped.

Figure 10-2 - Cross Section 11,775 NW

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11.0

SAMPLE PREPARATION, ANALYSES, AND SECURITY

11.1

Sample Preparation and Analysis

11.1.1

Drill Core Samples

The Santa Elina (1990 to 1997) core samples were crushed on site to quarter-inch and pulverized to minus 2 mm in a hammer mill. A fraction of this was saved for ore characterization, the rest was panned to remove the so-called “gravity” gold (plus 150 mesh), which was then amalgamated and assayed at the São Vicente Mine laboratory. The heavies that remained after amalgamation were assumed to contain no gold, but were captured and saved. The entire pan tailings were collected, dried, and split to a 1.8 kg sample and shipped to the Nomos laboratory in Rio de Janeiro for gold fire assay using a 50 g subsample and an AA finish. The value obtained from this assay was termed the fine or “chemical” gold assay. Therefore, each 2 m interval has two assays, gravity and chemical, and the sum of them being the total gold grade.

From 2003 to 2008 Yamana mainly used SGS (ISO 9001:2000) in Belo Horizonte for sample preparation and assaying of drill core from the São Francisco Mine. Core samples were delivered to the laboratory in sealed batches by truck to Cuiabá and then sent by air to Belo Horizonte. Certified standards and blanks were used for QA/QC purposes. Core duplicates, coarse reject samples, or duplicate pulp samples were not submitted.

In 2008, Yamana submitted some 40 certified standard samples and 40 blanks with the drill core samples from the São Francisco Mine at a rate of one standard and one blank for every 30 samples submitted. A total of 11 sample batches were submitted to the laboratory for that year.

To account for coarse gold in samples from areas with high hydrothermal alteration Yamana used a screen fire assay. A fire assay was used on samples from areas with low to medium alteration. The details of the sample preparation and assaying procedures used by SGS for São Francisco’s core samples are summarized in point form below.

A submittal form is filled out by a Mine geologist or technician and delivered with the samples to SGS.

Samples are sorted, logged in, opened, and dried at 110ºC as required.

The entire samples are crushed in a jaw crusher to better than 95% passing 10 mesh. Crushers are cleaned with compressed air between every sample and with quartz sand every 40 samples. Granulometric checks are done every 20 samples.

Fire Assay

A 1,000 g crushed subsample is taken by a riffle split. The split is pulverized using a steel ring mill to better than 95% passing 150 mesh. Granulometric checks were done every 20 samples. Pulverisers were cleaned with compressed air between every sample and with quartz sand every 40th sample.

Standard fire assay (FA) methods using 50 g pulp samples are used to determine total gold content.

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Screen Fire Assay

Each sample weighing approximately 6 to 8 kg is pulverized to 150 mesh and then sifted with a 100 mesh plastic sieve. The coarse fraction generally weighs between 30 g and 50 g. In this case, the coarse fraction is analyzed in total and the fine fraction is assayed in triplicate after homogenization.

Standard fire assay (FA) methods are used to determine total gold content.

The sample, fluxes, lead oxide litharge, and silver are mixed and fired at 1,050° C so that the precious metals report to the molten lead metal phase. The samples are removed from the furnace and poured into molds. The slag is then removed from the cooled lead button and the button is placed in a cupel and fired at 950° C to oxidize all the lead and render a precious metal bead.

The beads are weighed and digested in nitric and hydrochloric acid to take the precious metals into solution. The sample solutions are analyzed by ICP-ES. For metallic assays, the coarse fraction is assayed in total and an aliquot of the fine fraction is analyzed. The gold concentration of the entire sample is determined by weighted average.

Apoena

The Apoena core samples were assayed using a metallic screen fire assay methodology at the SGS laboratory in Belo Horizonte using the screen fire assay procedure described above. The SGS laboratory is ISO 9001-2000 certified (MET-1 &FAA505 procedure details can be viewed at http://www.br.sgs.com/pt_br/geochem_services.pdf).

The authors are of the opinion that the sample preparation and assay procedures used for the drill core samples are in keeping with industry standards and can be used in the estimation of a Mineral Resource and a Mineral Reserve.

Grade Control Sampling

Grade control samples at the Mine are collected from blast hole cuttings using a radial splitter (pie-shaped sample tray) placed at random around the cuttings pile. Samples in altered and mineralized material (ore) are spaced at approximately 3 x 3.5 m and samples in waste are spaced at approximately 4.5 x 6.0 m. All samples in ore and three to four rows of blast holes in waste are collected and sent for screen fire assay.

The QA/QC protocol for the blast hole samples consists of using both rig-duplicates (a second sample collected using a second sample tray) and barren material (blanks). The rig-duplicates are inserted to make up 10% of the samples submitted for assay. The blanks are inserted to make up 3% of the samples sent for assay. The QA/QC samples are analyzed periodically to assess the reliability of the blast holes sample assays.

A summary of the grade control sampling procedure is provided below.

Sample all blast holes in altered and mineralized material and selected blast holes in waste material

50 kg sample is collected from cuttings pile using a radial splitter

Geological description recorded for each blast hole sample

50 kg sample is crushed and split using a riffle splitter to a 1 kg subsample

1 kg subsample is pulverized and then assayed using screen fire assay

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The authors believe that the current grade control sample practices should provide reliable data for the short-term grade control block model and maintain operational efficiency by using an industry standard methodology of collecting grade control data.

11.1.2

Sample Security

Details of the sample security procedures for the pre-2003 sample preparation, analysis, and transportation are no longer available; however the remaining drill core from the São Francisco Mine is stored at a secure facility at the São Vicente Mine, as are the sample pulps and rejects. Drill core from the Apoena drilling are securely stored at the Mine.

For samples from 2003 and onward, the sampling and bagging was carried out or closely supervised by a company geologist. Samples were then transported by a commercial trucking contractor directly to the independent certified analytical laboratories located in or near Belo Horizonte in Minas Gerais State, Brazil. Upon receipt of the samples, the laboratories were instructed to report any apparent tampering or inconsistencies in sample numbering to the company; however, no discrepancies or tampering were encountered.

11.2

Author's Opinion on Sample Preparation, Security, and Analytical Procedures

The authors believe that the sample preparation and assay procedures used for the drill core samples are in keeping with industry standards and can be used in the estimation of a Mineral Resource and a Mineral Reserve.

Based on a review of security protocols and discussions with Apoena personnel, the authors are of the opinion that sample security procedures at São Francisco Mine are in keeping with industry standards.

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12.0

DATA VERIFICATION

12.1

Previous Exploration QA/QC Programs

The QA/QC results for the Santa Elina drilling programs were analyzed by Mr. K. Lovstrom and are reported to be acceptable.

Yamana submitted blank and certified standard material as part of their QA/QC protocol for their drilling programs. The analysis of the QA/QC results was carried out by International GeoSystems Inc. (“GSI”) and is documented in the reports, São Francisco Gold Project Resource Model Update prepared by GeoSystems International Inc. March 8, 2006 and São Francisco Mine, Serra da Borda Mineração e Metalurgia SBMM by Yamana and NCL Consultants, February 2008.

Overall, the QA/QC results are reported to be acceptable and the assays are considered sufficiently reliable for use in Mineral Resource and Mineral Reserve estimation. For details of the QA/QC please see the reports referenced above.

In 2009 MCB carried out a limited data review and held discussions with São Francisco Mine personnel, and based on this review, MCB was of the opinion that data entry and verification procedures of exploration data at the Mine are in keeping with industry standards.

12.2

Apoena Exploration QA/QC Program

The drill hole data collected by Apoena is managed using an acQuire database management system that has appropriate QA/QC procedures to ensure the quality and reliability of the drill hole data. Both Apoena and Aura personnel have reviewed the drill hole assays and carried out spot checks to ensure that drill hole data was imported correctly. It is the opinion of the authors that the data management meets industry standards and that the data is sufficiently reliable such that it can be used for Mineral Resource and Mineral Reserve estimation.

The 2009 to 2011 drilling campaign inserted blanks and standard samples to monitor laboratory performance. As well, twin samples and pulp check duplicates were taken to assess the variability of the gold grades. The results obtained from the analysis of the duplicates in São Francisco Mine show extreme difficulty in reproducing the gold grades of the original samples.

12.2.1

Blanks

For the Apoena drilling campaign, a total of 185 blanks (samples with a grade of 0.0 g/t Au) were sent to the laboratory to test for potential contamination issues. Of those, 6 samples returned a grade above 0.01 g/t Au, with 1 sample having higher values. The highest value for blanks was 0.41 g/t Au. This high result is probably due to sample bag swaps or other handling errors rather than contamination.

The remaining 179 samples returned a lower-than-detection limit value, as would be expected from blank material.

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Figure 12-1 - Blank Analysis

12.2.2

Standards

The 2010-2011 drilling campaign used the certified standards from Ore Research & Exploration PTY LTD listed in Table 12-1 to check the reliability of the assays.

Table 12-1 - Standards to Check 2009-2011 Drilling Results

Standard	Grade (g/t)
STD 2Pd	0.885
STD 6Pc	1.52
STD 18C	3.52
STD 18Pb	3.63

A total of 193 standards were inserted randomly into the sample stream; 70 from 2Pd, 70 from 6Pc, 45 from 18C and 8 from 18Pb resulting in a fairly homogenous distribution of standard values. Analysis showed that standards behaved well throughout the campaign, as can be seen in Figures 12-2 through Figure 12-5, with results falling consistently within 3 standard deviations of the mean. Only two samples fell outside the 3 standard deviation envelope: MTDS001375, and SFDS000964.

Based on the results of these standards, the authors believe that there has been no statistical bias introduced by the laboratory during the 2009-2011 drilling campaign.

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Figure 12-2 - Results from Standard 2Pd

Figure 12-3 - Results from Standard 18C

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Figure 12-4 - Results from Standard 18Pb

Figure 12-5 - Results from Standard 6Pc

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12.2.3

Twin Samples

Twin samples, assays of quartered core, were taken to assess the variability of gold due to the cutting of the core. Twin sample results can be highly variable because the amount of coarse gold contained in the core could be highly variable from one quarter of the drill core to the other. A scatter plot showing the twin sample results can be seen in Figure 12-6. Clearly, there is a significant amount of variability and there is a small bias in the mean grades. The twin sample results confirm the highly variable nature of the gold mineralization at the Mine.

Figure 12-6 - Twin Sample Assays

12.2.4

Duplicate Pulp Assays

Pulp check assays collected during the 2009-2011 drill programs were sent to the ALS Chemex Group Belo Horizonte Minerals Lab. (“ALS”) laboratory in Belo Horizonte for fire assay with an AA finish as an independent check of the SGS assay results. Due to the sampling procedure at SGS, all the coarse material for the screen fire assay is consumed during the assay process and only the fine mesh portion (passing through 150 mesh) was sent for check assay. The graphs below show the SGS assay results against the ALS assay results.

The correlation is poor and there is a significant bias between the two assay results. The cause of these poor results is not known, but it should be investigated.

However, the results of the blanks and standards show that the assay results from SGS are reliable. Therefore, it is the opinion of the authors that the SGS assays from the 2009-2010 Apoena drilling program are sufficiently reliable that they can be used for Mineral Resource and Mineral Reserve estimation.

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Figure 12-7- Pulp Check Assays

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13.0

MINERAL PROCESSING AND METALLURGICAL TESTING

Metallurgical testing results presented herein were first presented in the November 2004 São Francisco Technical Report prepared for Yamana by KCA while the test work was conducted during 2004.

13.1

Pilot Scale Column Tests

A series of large column tests were conducted at São Vicente by SEDM personnel to estimate recovery on leaching low grade, ROM material. A KCA technician was onsite to supervise the column test program start-up. Four separate pilot-scale column leach tests were set up on four separate ROM samples from the shaft developed at São Francisco. The columns were 1.8 m in diameter by 10 m tall.

Two of the samples represented material identified as PDC material. These samples came from the shaft between 9.10 m and 25 m. The other two samples were identified as FOX 1 material and came from the shaft between 57 m and 73 m in depth.

After leaching, the columns were rinsed with a combination of fresh water and hydrogen peroxide to neutralize the residual sodium cyanide. The columns were then discharged and transported to the gravity pilot plant facility at São Francisco where they were processed individually to determine residual gold values (tailings assays). The results of the leach tests are as outlined in Table 13-1.

Table 13-1 - ROM Large Column Test Results

Colum No.	Sample I.D.	Extracted (g/t Au)	Tailings (g/t Au)	Calc.Head (g/t Au)	Extraction (% Au)	Days of Leach
1	PDC,9.10-25m	0.11	0.148	0.258	43	126
2	PDC,9.10-25m	0.11	0.064	0.174	63	111
3	FOX1,57-73m	0.22	0.062	0.282	78	86
4	FOX1,57-73m	0.21	0.054	0.264	80	86

13.2

Direct Column Leach Tests

Bottle roll and column leach tests were conducted by KCA on five separate ore types from 15 separate core holes. The ore types tested were as follows: DCP, PDC (2 composites), FOX 1 (2 composites), FOX 2 and FSF. PDC, FOX1 and FOX2 make up a majority of the orebody.

Bottle roll leach tests were completed on each of the composites at a grind size of 100% passing 0.15 mm. These results are presented in Table 13-2.

Table 13-2 - Summary of KCA Bottle Roll Leach Tests

Ore Type	Calculated Head (g/t Au)	Extracted (g/t Au)	Extracted (% Au)	Crush Size (mm)	Days of Leach	Wt. in Leach (kg)	Cons. NaCN (kg/t)	Add Ca(OH)₂ (kg/t)
DCP	0.88	0.73	83	-0.15	4	1.0	0.42	1.00
PDC	0.54	0.42	78	-0.15	4	1.0	0.17	1.00
PDC	3.95	3.89	98	-0.15	4	1.0	0.27	1.00
FOX1	1.26	1.21	96	-0.15	4	1.0	0.19	1.00
FOX1	0.82	0.77	94	-0.15	4	1.0	0.19	1.00
FOX2	1.03	0.98	95	-0.15	4	1.0	0.10	1.00
FSF	0.45	0.38	85	-0.15	4	1.0	0.25	1.00

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The bottle roll leach test results indicate all ore types are amenable to cyanide leaching with moderate to high recoveries for the ore types tested.

A series of column leach tests were completed on each composite sample crushed to minus 50 mm. These results are summarized in Table 13-3.

Table 13-3 - Summary of KCA Direct Column Leach Tests

Ore Type	Calculated Head (g/t Au)	Extracted (g/t Au)	Extracted (% Au)	Crush Size (mm)	Days of Leach	Wt.in Leach (kg)	Cons. NaCN (kg/t)	Add Cement (kg/t)
DCP	1.417	1.101	78	-50	108	81.12	0.55	15
PDC	1.161	0.598	51	-50	108	80.93	0.65	10
PDC	0.806	0.379	47	-50	108	81.01	0.63	10
FOX1	0.716	0.386	54	-50	108	80.47	0.91	5
FOX1	1.002	0.602	60	-50	108	80.94	0.63	10
FOX2	2.014	0.788	39	-50	108	80.46	0.85	5

The weighted average gold recovery based on the applicable major ore types of the above tests was 52%. The tests were shutdown at the request of Yamana while they were still leaching which led to lower recoveries when compared to past column leach test results.

13.3

Pilot Gravity Tests

The report by Francis Pitard on sampling (“Review of Sampling Systems and Sampling Practices at the Yamana Gold Inc. Operations in Brazil”) recommended that bulk samples be processed through a pilot gravity plant to obtain a better understanding of the effect of coarse gold on the overall gold grade. The plant utilized in the tests included crushers, screens, hydro-cyclones, jigs, a Knelson concentrator and associated equipment. The plant had a throughput capacity of between 4 tph and 5 tph of ore and included a primary jaw crusher and secondary impact crusher. The secondary crusher was operated in closed circuit with a vibrating screen.

Two series of tests were conducted. The first series was on three bulk samples from the shaft. Each bulk sample was crushed to a nominal 25 mm and then screened at 6 mm. The second series of tests were on samples from distinct intervals from the shaft. These samples were crushed to a nominal 6 mm, and then screened at 2 mm. The minus fractions (2 mm or 6 mm) fed the gravity plant circuit. Two types of tailings from the plant were produced: the final gravity tails from Knelson concentrator and final slimes from the cyclone overflow. The plus 6 mm fractions and the minus 6 mm gravity tailings were sent to KCA for column leach testing.

The three bulk samples that were separately processed in the gravity pilot plant included the following:

Test 1: PDC zone – 25 m to 57 m;

Test 2: FOX-1 zone, upper level – 103 m to 130 m;

Test 3: FOX-1 deeper zone, high-grade – 130 m to 150 m.

The separate intervals from the shaft were processed through the gravity plant to obtain a comparison of shaft intervals with drill core data.

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The main conclusions from the pilot plant tests are that about 30% of total gold content is recoverable in a gravity plant. The gold remaining in the minus 6-mm material and the plus 6-mm material require heap leaching methods for recovery.

On average, the slimes tail has a grade of about 1 g/t Au, which represents about 5% of the total gold content. The dewatering cyclone used in the pilot plant was not optimized for this purpose, leading to a higher than expected slimes loss. The gold in the slimes is presently considered a loss, but proper sizing and operation of the equipment in the production plant is expected to reduce this. The slimes will be deposited into a separate pond for storage and possible future treatment.

The gravity plant tests provided confirmation of Mr. Pitard’s statement that bulk sampling will be required to detect and evaluate coarse gold particles. In general, the bulk sampling tests in high-grade intervals were significantly higher than (in some cases double) the original gold grade obtained in the drill hole samples.

13.4

KCA Column Leach Tests of Gravity Pilot Plant Ore

Samples of the minus 6-mm gravity tailings and the plus 6-mm screened material from the three special gravity plant tests were sent to KCA for column leach testing. The results of these tests are shown in Table 13-4. The plus 6-mm column tests appeared to have finished leaching at the time the tests were ended, while the minus 6-mm tests were still leaching. However, SEDM requested that the tests be ended to allow the tails to be analyzed, so that a final recovery could be calculated. The results given in Table 13-4 show the test results at the time the tests were ended.

It was apparent that the minus 6-mm tests were still leaching; therefore, KCA extrapolated from the end of the leach tests to estimate an ultimate recovery that would have been expected if the tests had been allowed to continue to completion. In extrapolating the data, KCA considered the leach performance for some of the other long-term leach tests from 1997/1998 as well as the 2004 tests. Table 13-4 also includes the extrapolated ultimate recoveries for the minus 6-mm tests.

The samples received by KCA for the plus and minus 6-mm products were relatively large and were loaded into larger than typical columns. Because of this, it was decided that the tails from the columns could be split, with one half being put back into a smaller column to continue leaching, while the other half was processed to determine the gold remaining in the tails. Most of the re-leach columns were showing additional recovery at the time of KCA’s report that indicates KCA’s extrapolated values were reasonable or possibly even conservative. In addition to assaying the tailings, KCA planned to run a bottle roll test on a portion of the tailings from each column tails sample. These additional test results were not available at the time of completion of KCA’s report.

Table 13-4 - Summary of the Gravity Plant Column Leach Tests

Ore Type	Test No.	Sample ID	Calc.Head (g/t Au)	Extracted (g/t Au)	Tails (g/t Au)	Rec. (% Au)	Crush Size (mm)	Leach Time (days)	NaCN Cons (kg/t)	Projected Ultimate Recovery
PDC	1	25x6	0.37	0.27	0.10	73%	25x6	64	2.48	73%
PDC	1	CycUF	0.61	0.44	0.16	74%	-6	66	0.27	80%
FOX1	2	25x6	0.29	0.22	0.07	76%	25x6	64	2.28	76%
FOX1	2	CycUF	0.76	0.31	0.45	41%	-6	66	0.28	43%
FOX1	3	25x6	0.73	0.53	0.18	75%	25x6	64	2.52	84%
FOX1	3	CycUF	0.89	0.63	0.26	71%	-6	64	0.28	75%

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14.0

MINERAL RESOURCE ESTIMATES

The block model used to estimate the September 30, 2011 Mineral Resource for the São Francisco Mine was prepared under the supervision of Aura personnel using the data and geological interpretation provided by Aura.

It has been noted that the block model used for the Dec. 31, 2010 Mineral Resource estimate has not reconciled well with mine production and grade control mapping and sampling. Different modeling and estimation approaches were considered in order to generate a block model that reconciles better with production and would therefore produce a more realistic estimate of the remaining Mineral Resource.

A new geological model was constructed using an indicator methodology that considered both the geological interpretation based on the drill holes and the geological interpretation based on blast holes and pit mapping. The gold mineralization was grouped into five geological domains that were used for grade estimation. The block model was generated using Datamine® mining software and used Ordinary Kriging (“OK”) to interpolate the block model grade estimates.

14.1

Geological Interpretation

The geological information was reviewed to determine a practical method of modeling the mineralization that would improve mine reconciliation. The geological information available for the São Francisco Mine comes from two sources. One source is the geology interpretation and wireframes constructed using the exploration drill hole data by Yamana geologists. The other source is the geological interpretation and wireframes based on the logging of blast hole cuttings and the mapping of pit walls and benches by the Mine geologists.

The geological interpretation based on the exploration drill holes has been referred to as the long-term (“LT”) model because this model is used for long-term and life-of-mine mine planning. The geological interpretation based on the blast hole logging and pit mapping is referred to as the short-term (“ST”) model because it is used for grade control and short-term mine planning.

14.1.1

Long-term Wireframes

The long-term geological interpretation was constructed by Yamana using the diamond drill hole database. Three geology wireframes and two topographic surfaces were constructed:

Saprolite;

Low Grade Hydrothermal Alteration Zone;

High Grade Hydrothermal Alteration Zone;

Pre-mining topography, and

Current “as-built” topography showing the mined out volumes and waste dumps.

The high grade hydrothermal alteration zone (“HG-HAZ”) wireframe was fully enclosed by the low grade hydrothermal alteration zone (“LG-HAZ”) wireframe and all material outside of the LG-HAZ wireframe and below saprolite and/or topography was coded as waste. The method used to code material as HG-HAZ, LG-HAZ or waste is described in technical reports written by Yamana and NCL Brasil. In summary the interpretation utilized a combination of alteration fields captured in the drill logs, along with the presence of visible gold and the gold grade, to code an interval as high grade, low grade or waste:

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HG-HAZ: material with moderate to strong sulphide content, moderate to strong quartz veining, visible gold or a sample grade >= 0.4 g/t Au.

LG-HAZ: material not previously classified as high grade in which sulphides were observed.

14.1.2

Short-term Wireframes

The short-term wireframes were built by the São Francisco Mine geology staff using open pit wall or “face” mapping along with blast hole and grade control drill hole sample results. Three geological wireframes were constructed:

ORE: higher grade material (> 0.4 Au g/t) to be sent to the mill for crushing and gravity recovery of coarse gold, before being placed on the leach pad,

DLO: direct leach ore shipped directly to the leach pad for processing by cyanide leaching,

Waste.

14.2

Exploratory Data Analysis

14.2.1

Exploration Drill Hole Database

The March 5, 2011 drill hole database used in this analysis consisted of 524 diamond drill holes totalling 93,666 m of which 43,527 intervals totalling 87,073 m were sampled. The database included holes numbered SF-001 to SF-246 and SF-252 to SF-529. The summary statistics of the Au (g/t) assays, separated by logged lithology are shown in Table 14-1 below. The lithology legend is summarized in Table 14-2.

Table 14-1 - Summary Statistics of Exploration Database - Au (g/t)

	Len. (m)	SAP	CGL	CMS	MSP	FMS	HAZ	Absent	TOTAL
Mean	2.00	0.45	0.01	0.04	0.03	0.13	0.36	0.08	0.28
Max	52	36.7	0.13	2.58	15.08	17.91	311.11	4.37	311.11
Q3	2	0.41	0.01	0.02	0.01	0.037	0.2	0.01	0.15
Median	2	0.15	0.005	0.005	0.005	0.01	0.05	0	0.026
Q1	2	0.04	0	0.003	0	0.005	0.01	0	0.005
Min	0.2	0	0	0	0	0	0	0	0
IQR		0.37	0.01	0.02	0.01	0.03	0.19	0.01	0.15
CV		3.03	1.96	4.16	7.45	6.18	7.83	5.55	8.51
Number	43,527	1,845	43	562	9,687	1,210	30,071	109	43,527
ns	3,274	230	87	115	2,483	27	331	1	3,274

Table 14-2 - Summary Exploration Lithology Code

LITHOLOGY	Code	Abbrev.
Decomposed Rock + Soil + Laterite	1	SAP
Metaconglomerates	2	CGL
Conglomeratic metasandstones	3	CMS
Metasandstones and meta-pelitic rocks	4	MSP
Feldspathic Metasandstones	5	FMS
Hydrothermal Alteration Zone	6	HAZ

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The summary statistics of the Au (g/t) assays, coded using the exploration wireframes, are shown in Table 14-3 below.

Table 14-3 - Summary Statistics of Exploration Database - Au (g/t) - Coded by Exploration Wireframes

	Len. (m)	SAP	LG-HAZ	HG-HAZ	All HAZ	Waste	Absent	Total
Mean	2.00	0.44	0.17	0.74	0.35	0.04	0.03	0.28
Max	52	36.7	311.11	117.4	311.11	30.67	0.74	311.11
Q3	2	0.41	0.09	0.444	0.190	0.011	0.02	0.15
Median	2	0.15	0.02	0.18	0.048	0.005	0.005	0.026
Q1	2	0.04	0.005	0.06	0.01	0	0.005	0.005
Min	0.2	0	0	0	0	0	0.005	0
IQR		0.37	0.09	0.38	0.18	0.01	0.02	0.15
CV		3.04	13.95	4.78	7.93	10.39	2.67	8.51
Number	43,527	1,902	21,332	10,203	31,535	9,990	100	43,527
ns	3,274	134	149	18	167	2,973	0	3,274

14.2.2

Grade Control Database

The Grade Control database, as of March 5, 2011, consists of 15,347 holes totalling 154,320.7 m from which 15,959 intervals representing 154,268 m of drilling were sampled. The summary Au (g/t) statistics of the samples coded using the grade control wireframes, are summarized in Table 14-4 below.

Table 14-4 - Summary Statistics of Grade Control & Blast Hole Database - Au (g/t)

	Len. (m)	ORE	DLO	ORE + DLO	WST	Subtotal In Pit	Other	Total
Mean	9.67	0.86	0.49	0.64	0.17	0.52	0.38	0.51
Max	20	51.37	65.20	65.20	18.69	65.20	5.22	65.20
Q3	10	0.86	0.46	0.63	0.18	0.51	0.53	0.51
Median	10	0.47	0.25	0.33	0.08	0.24	0.3	0.24
Q1	10	0.27	0.13	0.16	0.04	0.10	0.07	0.10
Min	0.1	0.01	0	0	0	0	0	0
IQR		0.59	0.33	0.47	0.14	0.41	0.46	0.41
CV		2.16	3.00	2.57	2.92	2.81	1.37	2.80
Number	15,959	4,746	6,664	11,410	4,209	15,619	340	15,959

14.2.3

Discussion and Interpretation

The LT and ST wireframes as well as the exploration and grade control databases were imported into Datamine®. A comparison of the two interpretations, represented by the two sets of wireframes, shows a reasonable correlation between the DLO, grade control outline and the HG-HAZ exploration outline. This indicates that the potentially mineralized material can be readily identified in drill core and in the open pit.

An examination of the statistics of the exploration samples, coded using the exploration wireframes, shows that the HG-HAZ wireframe appears to capture the great majority of the drill hole intervals with potentially economic grades. However, the median grade of these data is 0.18 g/t Au. This indicates that over half of the samples have grades that are lower than 0.2 g/t Au, the historical cutoff for heap leach ore. Furthermore, the grades of almost all of the samples classified as LG-HAZ by the exploration wireframe are below the heap leach cutoff grade.

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A similar examination of the blast hole statistics shows that, of the material classified as either ORE or DLO by the pit mapping, 25% of the blast hole samples have grades less than 0.16 (Au g/t). Further analysis shows that 30% of the samples have grades that are less than 0.2 g/t Au.

It is apparent that neither set of geological wireframes adequately models the erratic “nuggety” nature of the potentially economic mineralized material and that a different approach is required.

14.3

Geological Modeling

A single indicator approach was chosen to model the higher grade material within the broad hydrothermally altered zone. The details of the model procedure are provided in Sections 14.3.1 and 14.3.2 below.

14.3.1

Block Model Parameters

To be consistent with previous block models generated by Yamana and the Mine, the same block model parameters were used. The block model was rotated, horizontally, 322° clockwise about the UTM coordinate point 8,354,915.5 m N, 215,243.59 m E and -110 m Elev. so that the block boundaries are parallel to the strike of the mineralization. Uniform blocks of 10 m in each dimension were used and sub-celling was only permitted at the pre-mining and current topographic surfaces. The block model parameters are summarized in Table 14-5.

Table 14-5 - São Francisco Block Model Parameters

	East (m)	North (m)	Elev. (m)
Rotation	0	0	322
Origin	215,243.59	8,354,915.5	-110.0
	East (m)	North (m)	Elev. (m)
Minimum	215,243.59	8,354,915.5	-110.0
Maximum	216,693.59	8,358,815.5	820.0
Extent	1,450	3,900	930
Blocks	145	390	93
Block Size	10	10	10

14.3.2

Indicator Block Model

To better model the mineralization and the Au grades at the Mine, the Saprolite, LG-HAZ, HG-HAZ and surrounding rock were further divided into higher and lower grade volumes based on the probability of a block having a grade greater than an indicator Au cutoff. The steps involved in subdividing the geology model into high and low grade domains were:

5 m drill hole composites with a grade greater than the Au cutoff were assigned an Indicator value of one, otherwise the Indicator value was zero.

	·	Indicators were calculated for a range of Au cutoffs: 0.1, 0.15, 0.2, 0.25, 0.3, 0.35 and 0.4 g/t.
	·	Semi-variograms of each Indicator were calculated in the three orthogonal directions representing the strike, dip direction and thickness of the São Francisco deposit. (Table 14-6)

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Ordinary Kriging was used to interpolate the Indicators for each block. (Table 14-7)

A detailed analysis of the indicator kriging results was carried out and the 0.15 g/t Au Indicator was selected as the indicator result that best modeled the potentially economic mineralization. The 0.15 g/t Au Indicator generated cohesive groups of blocks that generally matched the DLO grade control wireframe and the HG-HAZ wireframe. Groups of blocks with a 50% or greater probability of being higher grade than 0.15 g/t Au were also observed to extend outside the HG-HAZ wireframe and within the LG-HAZ wireframe. This can be seen in Figures 14-1 and 14-2 below.

Figure 14-1 is a representative cross section through the 0.15 g/t Au Indicator model and Figure 14-2 is a plan view of the 0.15 g/t Au Indicator estimates The LG-HAZ and HG-HAZ wireframes are the dashed yellow and red outlines, respectively, while the DLO and ORE wireframes are the solid pink and red outlines, respectively.

Figure 14-1- São Francisco Cross Section 250S – 0.15 g/t Au Indicator

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Figure 14-2- São Francisco 735m Plan View – 0.15 g/t Au Indicator

The semi-variogram model of the 0.15 g/t Au Indicator was modeled using a nugget effect and three spherical structures. The semi-variogram parameters of the 0.15 g/t Au Indicator are summarized in Table 14-6 and a plot of the semi-variogram can be found in Figure 14-3.

Table 14-6 - São Francisco 0.15 g/t Au Indicator Variogram Parameters

IND 0.15		Range			Proportion of Sill
Orientation -38/55/0		X	Y	Z	%	Cum.
C0	0.067	-	-	-	28%	28%
C1	0.056	8	15	29	24%	52%
C2	0.060	72	56	58	25%	77%
C3	0.054	204	291	95	23%	100%
Sill	0.237

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Figure 14-3 - São Francisco 0.15 g/t Au Indicator Variogram Model

The indicator kriging considered two passes, the first used a search radii of 46 m by 28 m by 19 m and required a minimum of five and a maximum of 15 5 m composites with a maximum of three from any one drill hole and two drill holes were required to make an estimate. The second pass used a larger search, with search radii of 92 m by 56 m by 38 m and required a minimum of three and maximum of 15 5 m composites with a maximum of two per drill hole. The search parameters for the indicator kriging are provided in Table 14-7.

Table 14-7 - Estimation Parameters for 0.15 g/t Au Indicator

IND 0.15				Search Orientation		-38/55/0 ¹
	Search Radii			Composites			DH
SVOL	Dip	Strike	Thk.	MIN	MAX	Max/DDH	Min
1	46	28	19	5	15	3	2
2	92	56	38	3	15	2	2

¹ The Datamine rotation convention is positive clockwise about the Z/Y/X axes.

After selecting the Indicator which best modeled the mineralization, it was decided to further subdivide the model. Blocks with an indicator estimate greater than or equal to 0.5 were coded as high grade and those with an indicator less than 0.5 were flagged as low grade, thus generating new Domain codes as shown in Table 14-8.

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Table 14-8 - São Francisco Estimation Domain Codes

LITH	GCLAS
LITH	0	1
Saprolite	10	11
LG-HAZ	20	21
HG-HAZ	30	31
Waste	90	90

An analysis of composites on either side of LG-HAZ/HG-HAZ contact shows a gradational grade change across the contact. The contact profiles across the LG-HAZ/HG-HAZ contact are shown in Figure 14-4 below. The low grade domains (20 and 30) are shown on the left and the high grade domains (21 and 31) on the right. It was decided therefore to consider this contact to be a “soft” contact for the purposes of grade interpolation.

Figure 14-4 - São Francisco Contact Profiles Across LG-HAZ/HG-HAZ Contact

The domain codes were then revised to reflect the soft contact between the LG-HAZ and HG-HAZ domains by combining them to create Domains 50 and 51, as shown below in Table 14-9.

Table 14-9 - São Francisco Estimation Domain Codes

LITH	GCLAS
LITH	0	1
Saprolite	10	11
LG-HAZ	50	51
HG-HAZ	50	51
Waste	90	90

The Indicator approach was selected because the resulting geological domains appear to better separate the gold grades, than do the HG-HAZ exploration wireframes. As discussed above in section 14.2.3, a significant proportion of the data within the HG-HAZ wireframe (over 50%) have a grade of less than 0.20 g/t Au, the historical cutoff for heap leach ore. The summary statistics of the 5 m composites coded using the 0.15 g/t Au indicator model (Table 14-10, below) show a better separation of gold grades between the HG-HAZ and LG-HAZ domains, 51 and 50 respectively.

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14.4

Composite Statistics and Capping

The drill hole samples were composited to a uniform 5 m length within the lithological domains contained in the original drill logs. Remnants greater than or equal to 2.5 m in length (½ the default composite length) were included in the composite database and remnants less than 2.5 m in length were added to the previous 5 m composite, if possible. Unsampled drill hole intervals were assigned a grade of 0.0 Au (g/t). The statistics of the uncapped composites are summarized below in Table 14-10.

Table 14-10 - Summary Statistics of São Francisco 5 m Composites - Au (g/t)

Au (g/t)	Saprolite			HAZ			Waste	Total
Domain	10	11	All	50	51	All	90	All Comps
Mean	0.09	0.73	0.43	0.11	0.71	0.34	0.03	0.26
Max	2.189	21.206	21.206	25.976	133.996	133.996	8.735	133.996
Q3	0.100	0.745	0.446	0.092	0.547	0.252	0.014	0.178
Median	0.039	0.375	0.166	0.029	0.279	0.080	0.005	0.035
Q1	0.005	0.203	0.039	0.008	0.140	0.015	0.000	0.005
Min	0	0.003	0	0	0	0	0	0
IQR	0.10	0.54	0.41	0.08	0.41	0.24	0.01	0.17
CV	2.16	2.07	2.68	4.93	3.97	5.36	6.59	5.99
Number	363	416	779	8,021	5,092	13,113	5,600	19,492

Figure 14-5 - São Francisco Histogram and Lognormal Probability Plot of 5 m Composites

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In precious metal deposits, the assays typically have a positively skewed lognormal looking distribution where the mean is greater than the median and a small number of high grade data have a large effect on the mean. This can be seen in the histogram and cumulative-probability plot of the 5 m composites shown in Figure 14-5. These high grade assays are often referred to as “outliers”, and can account for a significant proportion of the apparent metal content of a deposit. Dealing with them in a way that prevents them from having a disproportionate effect on the Mineral Resource estimate is a challenge when working with precious metal deposits. A common approach to handling high grade data is to cap them and the selection of capping thresholds can also be a challenge.

Capping thresholds were initially selected by looking for breaks in the cumulative probability plots while paying attention to the coefficient of variation (CV = Standard Deviation/Mean) and the metal content. A high CV, greater than 1.5, often indicates that capping should be considered. From this analysis, three potential capping grades were considered, 6, 10, and 15 g/t Au.

14.4.1

Reconciliation to Blast Hole Data

Since São Francisco Mine is in operation, production data were available in the form of the grade control blast hole assays, which can be used to reconcile the estimated block grades. The blocks within the ORE and DLO grade control wireframes were estimated using inverse squared distance (ID²) and using the blast hole data capped to 6 g/t Au. The 6 g/t Au blast hole capping grade was used because this was the capping grade used at the Mine for their grade control block model.

Common volumes were then estimated using OK and the 5 m composites of the high grade domains: 11 and 51 were capped at 6, 10, and 15 g/t Au. The 5 m composites of the low grade domains 10, 50, and 90 were capped at 1 g/t Au. An estimate using uncapped data was also generated for comparison purposes.

Grade and Tonnage curves of the ID² estimate using blast hole data and the four OK estimates using 5m composite data are shown in Figure 14-6. While none of the OK estimates were able to match the average grade of the blast hole estimates, over the entire range of cutoff grades displayed, the OK estimates using a capping grade of 15 g/t Au track very well over the range of cutoffs used for grade control at the Mine: 0.15 to 0.4 g/t Au. It was therefore decided to adopt a capping grade of 15 g/t Au for the 5 m composites for the high grade domains 11 and 51.

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Figure 14-6 - Comparison of Blast Hole Estimate and Exploration OK Estimates

The capping thresholds, and the number of data affected, are summarized in Table 14-11. The statistics of the capped 5 m composites are shown in Table 14-12.

Table 14-11 - Capping Summary of São Francisco 5m Composites - Au (g/t)

Domain	Au (g/t)	Data Capped			Threshold
Domain	Threshold	Number	Prop (%)	Metal (%)	(%)
10	1.0	2	0.66%	6.1%	99.34%
11	15.0	1	0.24%	2.1%	99.76%
50	1.0	99	1.33%	23.8%	98.67%
51	15.0	19	0.37%	9.5%	99.63%
90	1.0	9	0.24%	13.6%	99.76%

Table 14-12 - Summary Statistics of São Francisco - Capped 5m Composites - Au (g/t)

Au Cap (g/t)	Saprolite			HAZ			Waste	All
Domain	10	11	All	50	51	All	90	Comps
Mean	0.08	0.72	0.42	0.08	0.64	0.30	0.02	0.23
Max	1.0	15.0	15.0	1.0	15.0	15.0	1.0	15.0
Q3	0.100	0.745	0.446	0.092	0.547	0.252	0.014	0.178
Median	0.039	0.375	0.166	0.029	0.279	0.080	0.005	0.035
Q1	0.005	0.203	0.039	0.008	0.140	0.015	0.000	0.005
Min	0	0.003	0	0	0	0	0	0
IQR	0.10	0.54	0.41	0.08	0.41	0.24	0.01	0.17
CV	1.62	1.86	2.44	1.88	2.29	3.22	3.18	3.68
Number	363	416	779	8,021	5,092	13,113	5,600	19,492

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14.5

Grade Estimation

14.5.1

Variography

Pairwise relative semi-variograms were calculated in three orthogonal directions (X, Y, & Z) representing the dip direction, strike and thickness, respectively, of the mineralization. The semi-variogram model parameters are summarized in Table 14-13, and plots of the semi-variograms is shown in Figures 14-7 and 14-8.

Table 14-13 - São Francisco Semi-Variogram Parameters

Domains: 10, 50, 90		Range			Proportion of Sill
Orientation -38/55/0		X	Y	Z	%	Cum.
C0	0.370	-	-	-	40%	40%
C1	0.100	11	26	10	11%	51%
C2	0.319	24	45	18	35%	86%
C3	0.126	289	1000	241	14%	100%
Sill	0.915
Domains: 11, 51		Range			Proportion of Sill
Orientation -38/55/0		X	Y	Z	%	Cum.
C0	0.411	-	-	-	84%	84%
C1	0.010	24	17	31	2%	86%
C2	0.029	80	40	83	6%	92%
C3	0.038	151	400	1000	8%	100%
Sill	0.488

Figure 14-7 - São Francisco Pairwise Relative Semi-variogram and Model for Domain 50

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Figure 14-8 - São Francisco Pairwise Relative Semi-variogram and Model for Domain 51

14.5.2

Estimation Parameters

Ordinary kriging was used to interpolate the Au block grades in the São Francisco deposit. Three search passes were used for the block model estimation. The first pass used a search ellipsoid with radii of 37 m by 26 m by 16 m and a minimum of four and a maximum of 15 composites, with a maximum of 3 composites from a single drill hole and a minimum of two drill holes were required to make an estimate.

The second pass used a search ellipsoid with radii of 74 m by 52 m by 32 m and a minimum of four and a maximum of 15 composites, with a maximum of 3 composites from a single drill hole and a minimum of two drill holes were required to make an estimate.

The third and final pass used a search ellipsoid with radii of 111 m by 78 m by 48 m and a minimum of two and a maximum of 15 composites, with a maximum of 3 composites from a single drill hole and a minimum of two drill holes were required to make an estimate.

The estimation parameters are summarized in Table 14-14.

Table 14-14 - São Francisco Estimation Parameters for Ordinary Kriging

São Francisco				Search Orientation		-38/55/0 1
	Search Radii			Composites			DH
SVOL	Dip	Strike	Th	MIN	MAX	Max/DDH	Min
1	37	26	16	4	15	3	2
2	74	52	32	4	15	3	2
3	111	78	48	2	15	3	2
1-The Datamine rotation convention is positive clockwise about the Z/Y/X axes.

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14.6

Block Model Validation

The block model was validated using a series of statistical and graphical summaries including: a global mean and a global variance validation, swath plots, a change of support validation and a visual validation of the block grade estimates in plans and sections.

14.6.1

Global Mean and Global Variance Validation

A global bias validation was carried out by comparing the composited assay weighted mean grade (for the 15 g/t capped gold) with the average grade estimates (also using the 15 g/t capped Au) of the nearest neighbour (NN), and OK models. Histograms and summary statistics can be seen below in Table 14-15. The OK estimates are globally unbiased because the mean gold grade for the 5 m composites, the NN estimates and OK estimates are all very similar.

Table 14-15 - Composite, NN, OK and ILC Statistics for Domain 51

	Mean	Std. Dev.	CV
Comps	0.64	1.465	2.29
NN	0.65	1.365	2.10
OK	0.66	0.636	0.96
ILC	0.65	0.500	0.77

A global variance validation was carried out by calculating the CV ratio for domain 51 between the OK estimates and the NN Au estimates that have been adjusted using the indirect lognormal correction (“ILC”) to a 10 x 10 x 10 m selective mining unit (“SMU”). The SMU is the smallest practical tonnage that can be used to separate ore from waste. For domain 51, the CV ratio is 0.81, which falls inside the preferred range of 75-95%. The preferred range is based on experience with similar deposits.

When the block model variance is less than the predicted variance of the SMUs and the CV ratio is inside the preferred range, and then based on experience, an appropriate amount of internal dilution has been incorporated into the block model estimates.

14.6.2

Trend Check

Local trends in the grade estimates were examined using swath plots for domain 51. The average estimate of the NN and OK block model estimates for 20 m wide swaths of easting, 50 m swaths in the northing and 20 m wide swaths in elevation were plotted against the easting, northing and elevation block indices. Block indices were used because this block model is rotated.

Figures 14-9 to 14-11 display the swath plots of gold grade for domain 51. The OK estimates are shown as the red line and the NN estimates are shown as the cyan line. In general, there is good agreement between the NN and OK models.

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Figure 14-9 - Swath Plot for Domain 51 Comparing NN and OK Au (g/t) Grades in the East Direction

Figure 14-10 - Swath Plot for Domain 51 Comparing NN and OK Au (g/t) Grades in the North Direction

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Figure 14-11 - Swath Plot for Domain 51 Comparing NN and OK Au (g/t) Grades in the Vertical Direction

14.6.3

Change of Support Validation

An assessment of the amount of smoothing in the block model estimates was carried out using the ILC change-of-support method. The ILC method was used to predict the distribution of the SMU for block sizes 10 x 10 x 10 m, which is considered the size of the SMU at the mine given the size of the mining equipment and the mining method. Grade tonnage curves of the hypothetical change-of-support distribution were compared to the grade tonnage curve of the OK gold estimates to determine the level of selectivity in the estimated model.

Figure 14-12 shows a grade tonnage curve of the OK block model gold grade estimates as black lines and the ILC adjusted NN block model gold grade estimates as red lines. At the 0.23 g/t Au resource cutoff there is good agreement between the tonnes and grade of the two estimates. However, as the cutoff grade increases there are differences between the grade and tonnes, indicating that the block model may not honour the SMU over the full range of the mining cutoffs from 0.2 to 0.4 g/t Au.

However, recall that the grade tonnage curve in Figure 14-6 shows that there is good agreement between the OK block model estimates and the Grade Control block model estimates over the full range of mining cutoffs used at the São Francisco Mine. Therefore, it is the opinion of the authors that there is sufficient change of support to properly account for internal dilution.

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Figure 14-12 - Grade-Tonnage Curves of the OK Estimates (LT), and the Predicted 10x10x10 m SMU

14.6.4

Visual Validation

Detailed visual inspections of the gold grade estimates were conducted in both plan and section to ensure that interpolation results honoured the geological boundaries and the drill hole data. This validation included confirmation of the proper coding of blocks for each of the domains and the block grade estimates relative to the capped 5 m drill hole composites to ensure that the drill hole data were properly represented in the model.

Based on the examination of plans and sections and an interrogation of selected block grades, the estimates can be explained as a function of the surrounding composites, the variogram model used, and the kriging plan applied (Figures 14-13 and 14-14).

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Figure 14-13 - Section 11,175 NW – OK Block Model Grades and 5 m Composites

Figure 14-14 - Section 11,775 NW – OK Block Model Grades and 5 m Composites

14.7

Mineral Resource Classification

The mineralization at the São Francisco Mine has been classified in accordance with the definitions of NI 43-101 Regulations and the CIM Standards.

The relevant definitions for the CIM Standards are as follows:

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A Mineral Resource is a concentration or occurrence of natural, solid, inorganic or fossilized organic material in or on the Earth’s crust in such form and quantity and of such a grade or quality that it has reasonable prospects for eventual economic extraction. The location, quantity, grade, geological characteristics and continuity of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge.

The term “reasonable prospects for eventual economic extraction” implies a judgment (albeit preliminary) by the Qualified Person in respect of the technical and economic factors likely to influence the prospect of economic extraction, including the approximate mining parameters. In other words, a Mineral Resource is not an inventory of all mineralization drilled or sampled, regardless of cutoff grade, likely mining dimensions location or continuity. It is a realistic inventory of mineralization which, under assumed and justifiable technical and economic conditions, might become economically extractable.

An Inferred Mineral Resource is that part of a Mineral Resource, for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes.

An Indicated Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough for geological and grade continuity to be reasonably assumed.

A Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough to confirm both geological and grade continuity.

As discussed in many technical reports about the São Francisco deposit, the presence of coarse gold is a common “problem” in the deposit which results in a high nugget effect. For the semi-variograms of the gold grades in domain 50 (HAZ Low Grade), the nugget effect accounts for 40% of the variogram sill while in domain 51 (HAZ High Grade) it accounts for over 80% of the sill. These high nugget effects confirm that this is a significant amount of variability over short distances in the gold grades which are likely due to the coarse free gold, and this makes local estimation of mining block grades very difficult. As a result, the authors are of the opinion that none of the São Francisco mineral resource can be classified as Measured. This is consistent with recommendations made by Scott Wilson Roscoe Postle and Associates (RPA).

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Therefore, blocks in the block model that are estimated in the first 2 passes of the search described in Table 14-14 were classified as Indicated Mineral Resource. Blocks estimated in the third pass of the search were classified as Inferred Mineral Resource.

14.8

Mineral Resource Classification

The September 30, 2011 Mineral Resource is estimated by Aura to be 10.923 million tonnes of Indicated Mineral Resource at an average grade of approximately 0.95 g/t Au and an Inferred Mineral Resource of 0.083 million tonnes at an average grade of 0.47 g/t Au, using a long term $1,700 per ounce gold price, and a 0.23 g/t Au cutoff (Table 14-16). Note that the extent of the September 30, 2011 Mineral Resource has been limited by an optimized pit using a $1,700 per ounce gold price. Note that the Mineral Resources are inclusive of Mineral Reserves. Also note that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

Table 14-16 – September 30, 2011 Mineral Resource Estimate

Resource Category	Tonnes ('000)	Au (g/t)	Oz. ('000)
Measured	0	0	0
Indicated	10,923	0.95	334
Measured + Indicated	10,923	0.95	334
Inferred	83	0.47	1

*At 0.23 Au (g/t) cutoff. Numbers may not add up due to rounding.

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15.0

MINERAL RESERVE ESTIMATES

The estimate of Mineral Reserves is based on a LOM Plan and plant production schedule developed by Aura. The Economic criteria using the Lerch-Grossman algorithm for pit limit evaluations, including process recoveries and operating costs are provided in Table 15-1.

The pit was optimized using MineSight Software. A raw pit shell was created using the parameters listed and then smoothed and detailed to include ramps and honour the pit slope criteria. Other than the internal dilution inherent through the resource estimation process, no additional dilution or mining loss was factored into the reserve estimates.

The south east corner of the pit was reduced slightly from the optimized pit shell due to interference with dewatering lines and waste dumps. It is recommended that the area be reviewed to optimize the amount of recovered gold ounces. It is unlikely, however that the extra ore will result in significant recovered ounces of gold.

Other than the internal dilution through the resource estimation process, no additional dilution or mining loss was factored into the reserve estimates.

Table 15-1 – Pit and Cost Parameters

Pit and Cost Parameters
Bench height (m)	10
Road width(m)	15
Overall Pit Slope (°)	52
Bench face angle (°)	80
Minimum pit bottom(m)	10
Berm width (m) (Alternating	10 / 4.3
Ramp slope (%)	10
Gold Price(US$/oz)	1500
Gold Recovery (%)	80
Mining cost (US$/t)	3.34
CGO Plant Cost (US$/t)	7.32
G&A cost (US$/t)	2.76

The September 30, 2011 Mineral Reserve is estimated by Aura Minerals to be 10,890 million tonnes of Probable Mineral Reserve at an average grade of 0.91 g/t Au. The Mineral Reserve has been estimated using 0.25 g/t Au cutoff. A summary of the Mineral Reserve is provided in Table 15-2.

Other than the risks to production discussed below, there are no issues that could materially affect the reserve estimate:

Mechanical breakdown of plant equipment – unforeseen failure is always a risk, but through maintenance programs, and the availability of critical spares, this risk can be managed.

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Availability of consumables – Mine management must continually be aware of any shortfalls of consumables such as cyanide and explosives and quickly source other suppliers.

The Mineral Resource and Mineral Reserve are estimated in accordance with the CIM definitions and are considered NI 43-101 compliant.

15.1

Mineral Reserve Summary

The mineral reserve estimate is presented in table 15-2.

Table 15-2 - September 30, 2011 Mineral Reserve Estimate

Ore ^*	Tonne (t)'000	Au (g/t)	oz'000
Proven	-	-	-
Probable	10,890	0.91	318
Proven + Probable	10,890	0.91	318
Waste	36,512
W/O	3.35

*Reserves at a 0.25 Au (g/t) cutoff– Note numbers may not add due to rounding.

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16.0

MINING METHODS

Mining at the São Francisco property is by open pit method. The pit is accessed via a standard access road or ramp 15 m wide at a minus 8% grade which provides access to the various benches.

The open pit operations work two shifts per day at 10 hours per shift, with a six day work week, for 313 days per year, and the crusher circuit operates approximately 16 hours per day, seven days per week. Production starts at 6:00 pm on Monday and continues 24 hours per day until Sunday morning at 6:00 am at which time all of the mining and crushing departments are shut down for a weekly scheduled maintenance shift.

The plant is operated on a 24 hours by 7 days per week schedule, so a stockpile at the primary crusher is required to fed the plant by a front end loader when the mine is not working.

The ore is hauled by 35 tonnes haul trucks provided by a Brazilian contractor, Fagundes. Over 2011, the contractor has replaced the 25 t trucks with the larger more efficient 35 t trucks.

The Contractor can move approximately 2 Mt of material per month.

16.1

Drilling and Blasting

A contractor operates an emulsion explosives plant on site, with Apoena supplying the materials. The contractor is responsible for mixing the explosives, transporting them to the bench, and loading the holes.

16.2

Loading and Hauling

Ore and waste haulage and other material movement at the São Francisco Mine is currently carried out by Fagundes using 35 t trucks, mined by hydraulic excavators.

Fagundes has its own employees including maintenance staff, operating staff, and purchasing staff. They are responsible for all of their logistics such as shops, transportation, and accommodations. All of these services are provided at the current cost per distance of hauling rate. Extra activities are covered on a cost-plus rate that has to be authorized by Apoena.

Currently, the open pit is being mined with an average haul distance to the primary crusher of 3,200 m. The distance from the pits to the primary crusher will vary with the pits evolution. The planning team optimizes the haul roads to reduce the average haul distances.

16.3

Pit Slope Stability

The pit slopes are designed at an ultimate slope of 52°. The design is based on the recommendations from VOGBR (April/2010) and the mining experience acquired in the last few years. The authors have examined the current pit walls and found them to be in reasonably good shape.

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16.4

Road Maintenance

The pit roads are maintained using two Caterpillar 140H graders. A water truck is also utilized to water the ramp and keep dust levels down, given the large number of trucks operating within the pit. Caterpillar D6 and D8 dozers are also used for spreading material on the stockpiles and any work required in the pit.

16.5

Dewatering

During the wet season, there is a need for high capacity pumping to keep the pit operational. Careful planning of a low elevation sump and related pumping infrastructure is required. During this period the lowest portion of the pit where the sump is located, is not accessible for mining.

16.6

Maintenance

Maintenance is carried out by mine employees at the maintenance facilities located at the mine. The contractor is responsible for maintenance of its fleet of vehicles.

16.7

Power

Power for the operation is supplied via the national power grid that delivers power to the site at 138 kV, 60 Hz. The average power cost for 2010 was approximately $142.71/MWh, which is a combination of both the sales and distribution contracts, while in 2011 to September averaged $168.47/MWh. There are also two standby generators on site with two megawatts of capacity each.

16.8

Communications

The site communications are provided via a microwave system that supplies telephone and Internet services for the operations and personnel.

16.9

Mine Plan and Schedule

The LOM Plan results are shown below by quarters:

Table 16-1 - 4th Quarter 2011 Material Movement Forecast

Period	Ore			Waste
Period	Tonnage (t)	Au (g/t)	Ounces (oz)	Tonnage (t)	W/O
2011 / 4th Quarter	1,784,151	0.66	37,629	4,557,036	2.55

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Figure 16-1 – 2011 4^th Quarter Material Movement Forecast

Table 16-2 – 1st Quarter 2012 Material Movement Forecast

Period	Ore			Waste
Period	Tonnage (t)	Au (g/t)	Ounces (oz)	Tonnage (t)	W/O
2012 / 1st Quarter	1,131,983	0.68	24,821	4,461,564	3.94

Figure 16-2 2012 1 Quarter Material Movement Forecast

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Table 16-3 - 2nd Quarter 2012 Material Movement Forecast

Period	Ore			Waste
Period	Tonnage (t)	Au (g/t)	Ounces (oz)	Tonnage (t)	W/O
2012 / 2nd Quarter	1,402,411	0.66	29,759	4,628,723	3.30

Figure 16-3 –2nd Quarter 2012 Material Movement Forecast

Table 16-4 - 3rd Quarter 2012 Material Movement Forecast

Period	Ore			Waste
Period	Tonnage (t)	Au (g/t)	Ounces (oz)	Tonnage (t)	W/O
2012 / 3rd Quarter	1,280,976	0.84	34,760	4,956,068	3.87

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Figure 16-4 –3rd Quarter 2012 Material Movement Forecast

Table 16-5 - 4th Quarter 2012 Material Movement Forecast

Period	Ore			Waste
Period	Tonnage (t)	Au (g/t)	Ounces (oz)	Tonnage (t)	W/O
2012 / 4th Quarter	1,128,457	0.93	33,596	4,904,653	4.35

Figure 16-5 –4th Quarter 2012 Material Movement Forecast

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Table 16-6 – 1st Quarter 2013 Material Movement Forecast

Period	Ore			Waste
Period	Tonnage (t)	Au (g/t)	Ounces (oz)	Tonnage (t)	W/O
2013 / 1 Quarter	1,133,350	1.00	36,584	4,198,698	3.70

Figure 16-6 –1st Quarter 2013 Material Movement Forecast

Table 16-7 - 2nd Quarter 2013 Material Movement Forecast

Period	Ore			Waste
Period	Tonnage (t)	Au (g/t)	Ounces (oz)	Tonnage (t)	W/O
2013 / 2nd Quarter	1,337,297	1.25	53,830	4,919,223	3.68

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Figure 16-7 –2nd Quarter 2013 Material Movement Forecast

Table 16-8 - 3rd Quarter 2013 Material Movement Forecast

Period	Ore			Waste
Period	Tonnage (t)	Au (g/t)	Ounces (oz)	Tonnage (t)	W/O
2013 / 3rd Quarter	1,691,297	1.23	66,948	3,886,314	2.30

Figure 16-8 –3rd Quarter 2013 Material Movement Forecast

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Table 16-9 - Mine Sequence Results – Total

Period	Ore			Waste
Period	Tonnage (t)	Au (g/t)	Ounces (oz)	Tonnage (t)	W/O
2011 / 4th Quarter	1,784,151	0.66	37,629	4,557,036	2.55
2012 / 1 Quarter	1,131,983	0.68	24,821	4,461,564	3.94
2012 / 2nd Quarter	1,402,411	0.66	29,759	4,628,723	3.30
2012 / 3rd Quarter	1,280,976	0.84	34,760	4,956,068	3.87
2012 / 4th Quarter	1,128,457	0.93	33,596	4,904,653	4.35
2013 / 1 Quarter	1,133,350	1.00	36,584	4,198,698	3.70
2013 / 2nd Quarter	1,337,297	1.25	53,830	4,919,223	3.68
2013 / 3rd Quarter	1,691,297	1.23	66,948	3,886,314	2.30
TOTAL	10,889,923	0.91	317,926	36,512,278	3.35

Figure 16-9 – Mine Sequence Results - Total

16.10

Mining Fleet and Machinery

A list of the contractor’s mine equipment is provided in Table 16-10 and haulage cost in table 16-11.

Table 16-10 - Mine Equipment

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Table 16-11 - 2010 Contract Haulage Cost

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17.0

RECOVERY METHODS

17.1

Process Plant Description

The São Francisco Mine’s mineral processing system consists of a gravity concentration circuit combined with a multi-lift, heap leach operation (Figure 17-1).

The CGO is processed by crushing and gravity concentration, followed by heap leaching of the tailings product from the circuit. The crushing circuit consists of three stages of crushing and has been designed to generate a product of 30% passing less than 6.0 mm to feed the gravity circuit. The remaining 70% is directed to the leach pads with the gravity tails. During 2011, the two categories of lower grade material were stockpiled separately. Sub-marginal grade is material between 0.15 g/t Au to 0.25 g/t Au, low grade is material between 0.25 g/t Au to 0.40 g/t Au. CGO or high grade is material above 0.40 g/t Au.

At the heap leach area, crushed ore is stacked using trucks and track dozers. The leach pad is a multi-lift, single-use type pad. The basic components of the leach pad include a compacted earth foundation (a compacted soil sub-base), a 1.5 mm thick geo-membrane liner, a cushion layer of material on top of the geo-membrane, and a series of perforated plastic drainpipes, which are placed on top of the geo-membrane, under the cushion layer during stacking, to collect leach solutions. Following stacking of the crushed and ROM ore, both ore types are irrigated with leach solution and the resulting gold-bearing solutions are collected in a pregnant solution pond prior to further processing for recovery of gold.

A carbon ADR facility is used for gold recovery. The ADR facility includes a train of five, cascade-type, open-top up-flow carbon adsorption columns, a carbon desorption system, a carbon acid wash circuit, a precious metals recovery circuit that utilizes electro-winning, and a complete smelting system for gold production. Gravity plant concentrates and the loaded cathode material from the carbon recovery plant are combined and smelted to produce a gold doré product.

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Figure 17-1 - General Flow Sheet São Francisco Mine

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Approximately 2.1 million tonnes of ore were processed in 2011. Table 17-1 shows a summary of plant production and estimated gold recovered for the sum of the gravity plant and the heap leach pads.

Table 17-1 - 2011 Monthly Plant Production for São Francisco Mine

Month	Tonnes Processed	Gold Grade (g/t)	Total Gold Recovered (oz)
January	0	0.00	0
February	0	0.00	0
March	0	0.00	0
April	238,576	0.66	2,357
May	336,908	0.86	4,351
June	399,014	0.75	6,169
July	412,027	0.53	7,206
August	433,033	0.45	6,482
September	348,745	0.48	4,977
Total	2,168,303	0.61	31,542

17.2

Plant Design, Equipment Characteristics, and Specifications

17.2.1

Crushing Circuit

The crushing circuit was designed with an operational efficiency of 90%. The design feed rate to the circuit is 850 tph.

ROM is fed through a fixed grizzly. A hydraulic rock breaker is available to assist with oversize material. Grizzly undersize is fed to the primary crusher via a vibrating feeder. Primary crushing is carried out using a jaw crusher with a closed side setting (“CSS”) of 6 - 1/2 inches. Primary crusher product is transported via conveyor belt to the primary two-deck screen.

The oversize from the primary screen is fed directly to the secondary cone crusher with a CSS of 1-1/4 inches. The primary screen undersize joins with the secondary crusher product, which is then fed to a stockpile.

Material from this stockpile is fed to a small feed bin. The bin feeds the two (2) secondary two-deck screens operating in parallel. The oversize from each screen feeds one (1) of the two (2) tertiary cone crushers, operating in parallel. Each tertiary crusher has a CSS of ½ inch.

The undersize of the secondary screens is combined with the tertiary crusher product and fed to the tertiary two-deck wet vibrating screen. The undersize of the tertiary screen becomes the feed to the gravity circuit. The oversize of the tertiary screen is fed to a silo where lime is added. The discharge from the silo is then hauled via truck to the heap leach pad.

The Figure 17-2 shows the Process Flow Diagram of the crushing circuit.

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Figure 17-2 - Process Flow Diagram ("PFD") of the Crushing Circuit

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17.2.2

Gravity Circuit

The gravity circuit was designed with an operational efficiency of 83%. The design capacity of the circuit is 457 tph. The estimated gold recovery in the gravity circuit is 60%.

The gravity circuit is fed by the crushing circuit’s tertiary screen undersize. The gravity circuit begins with a hydro-cyclone. The underflow feeds the jig circuit at a design rate of 343 tph. The jig circuit consists of a circular jig (eight units) followed by two (2) Pan American secondary jigs in parallel, which both feed an additional Pan American tertiary jig. The tailings from the Pan American secondary jigs are fed back to the circular primary jig. The concentrate from the tertiary Pan American jig is fed to the coarse gravity recovery circuit, which consists of a centrifugal concentrator and a shaking table. The combined tailings from the Pan American jig, the primary hydrocyclone’s overflow and the fines gravity recovery circuit are fed to a spiral classifier.

The fines gravity recovery circuit consists of a centrifugal concentrator and a series of shaking tables. This centrifugal concentrator is fed by the secondary hydro-cyclone’s underflow at a flow rate of 33 tph. The secondary hydro-cyclone is fed by the primary hydro-cylone’s overflow at a flow rate of 38 tph. The products of both the fines and the coarse gravity recovery circuits are the concentrates that are sent directly to the smelter.

The slime tailing fraction from the spiral dewatering classifier overflow and both the secondary and tertiary hydro-cyclones’ overflows are fed to the tailings dam. The coarse product of the spiral classifier is hauled via trucks to the heap leach.

The Knelson concentrators feed the designated shaking tables circuit. The product of the shaking table’s circuit is a concentrate that is sent directly to the smelter.

The Process Flow Diagram for the gravity circuit is shown in the Figure 17-3.

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Figure 17-3 - PFD of the Gravity Circuit

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17.2.3

Heap Leach Pads

At the heap leach area, both crushed ore is stacked using trucks and track dozers. The leach pad is a multi-lift, single-use type pad. The basic components of the leach pad include a compacted earth foundation (a compacted soil sub-base), a 1.5 mm thick geo-membrane liner, a cushion layer of material on top of the geo-membrane, and a series of perforated plastic drainpipes, which are placed on top of the geo-membrane, under the cushion layer during stacking, to collect leach solutions. Following stacking of the crushed and ROM ore, both ore types are irrigated with leach solution and the resulting gold-bearing solutions are collected in a pregnant solution pond prior to further processing for recovery of gold.

The Process Flow Diagram for the Heap Leach Circuit is shown in Figure 17-4.

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Figure 17-4 - PFD of Heap Leach Circuit

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17.2.4

Carbon ADR Facility

The operational efficiency of the carbon ADR facility is 90% and the design capacity of the ADR facility is 2,125 m³/hr with an estimated overall plant gold recovery to gold doré of 78%.

The ADR facility includes a train of five, cascade-type, open-top up-flow carbon adsorption columns.

The electro-winning circuit operates with cathodes and anodes, under effect of an electric DC current.

Gravity plant concentrates and the loaded cathode material from the carbon recovery plant are combined and smelted to produce a gold doré product.

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Figure 17-5 - PFD of the ADR facility

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18.0

PROJECT INFRASTRUCTURE

The existing Mine camp facilities include lodges, administrative offices, medical clinic, restaurant, and a gravel air strip.

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19.0

MARKET STUDIES AND CONTRACTS

The principal commodities of gold and silver in doré from São Francisco are freely traded, at prices that are widely known, so that prospects for sale of any production are virtually assured. An average gold price of US$ 1500 per ounce for the reserves base case has been used.

The company currently has several operational contracts such as with Fagundes (loading and hauling) and Nitronel (blasting).

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20.0

ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL IMPACT

The ore processing plants and support facilities will be removed and the heap leach pads will be neutralized, re-contoured, and the land re-vegetated. The concrete pads and foundations will be removed and the roads and other operating surfaces will be reclaimed and re-vegetated. Reclamation activities will be initiated as soon as portions of the mine are no longer required. The current mine budget includes approximately US$ 11.3 million for reclamation and closure.

The São Francisco Mine operation has a valid environmental license. There are no current environmental liabilities or non-compliance issues for the property.

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21.0

CAPITAL AND OPERATING COSTS

The sustaining capital expenditures (“CAPEX”) are estimated by Aura to be US$ 15.9 million, as shown in Table 21-1.

Table 21-1 - Capital Expenditures

Area (US$ 000)	2012	2013	2014	2015	Total
Mine Equipment , dewatering	600	600	150		1,380
Plant Sustaining Projects	800	800	200		1,840
ADR Sustaining capital	600	600	150		1,380
Reclamation and Closure			8,365	2,910	11,275
Total Capital	2,000	2,000	8,865	2,910	15,875

The 2011 unit operating cost breakdown for the São Francisco Mine is shown in Table 1-10.

Table 21-2 - 2011 Operation Cost Breakdown

Area	Unit Costs
Mine (US$/t moved)
Drilling	0.39
Blasting	0.35
Haulage/Loading/Auxiliary (Contractor)	2.09
Engineering and Operations Management	0.11
Total Mine (US$/t moved)	2.94
Process ($/t processed)
Crushing and Stacking	2.58
Gravity	0.81
Leaching / Cyanidation	4.11
ADR/Electrolysis/Smelting	1.65
Total ($t processed)	9.15
G&A ($/t processed)	2.77

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22.0

ECONOMIC ANALYSIS

A post-tax Cash Flow Model has been developed for the LOM production schedule and capital and operating cost estimates. For all periods the after tax cash flow is positive, supporting the reserve designation. A sensitivity analysis has been completed that examined gold price, and operating cost sensitivity, and none of the analysis show a negative net present value.

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23.0

ADJACENT PROPERTIES

There are no adjacent properties to the São Francisco Mine.

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24.0

OTHER RELEVANT DATA AND INFORMATION

The São Francisco Mine carries out a monthly reconciliation between the LT model estimated using the drill hole data and the geological model (based on that data) and the ST block model estimated using the grade control samples (blast hole assays) and a geological model based on face mapping and blast hole logging.

The grade control practices employed at the São Francisco Mine consists of the following activities:

Face Mapping of the walls and exposed areas in the pit

Blast hole sampling and assaying (as described above)

Geological interpretation and modeling of the face mapping and blast hole logging

Short-term block model estimation in MineSight® software using inverse distance methodology

Ore control boundaries drawn to define mining units and control dilution.

Figure 24-1 graphically displays the results of the monthly reconciliation. The ST data are shown as red lines, with the tonnes shown as a dashed line and the metal shown as the solid line. The LT data are shown as black lines, with the tonnes shown as a dashed line and the metal shown as the solid line.

The LT tonnes tracks moderately well compared to the ST tonnes. While some months do show a bias of up to 20% in the tonnes between the two models, the last three months compare quite well, especially December 2011.

The LT metal tracks moderately well compared to the ST metal. While some months do show a bias of up to 15 to 20% between the two models, the last three months compare quite well, especially December 2011.

Overall, there appears to be an improvement between the tonnes and metal predicted by the LT and that of the ST when compared to past reconciliation results. This improvement in reconciliation is likely the result of changes made to the estimation parameters used for the estimation of the September 30, 2011 block model.

It is recommended that the São Francisco Mine continue to perform a monthly reconciliation between the LT and ST and monitor the performance of the LT and, if required, update the estimation parameters to improve the reconciliation between the models.

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Figure 24-1 – Monthly Reconciliation between LT and ST Block Model Results

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25.0

INTERPRETATION AND CONCLUSIONS

The economic results show that the deposit can be mined at a profit. Significant risks to the project’s economics, along with mitigation measures include:

Mechanical breakdown of plant equipment – unforeseen failure is always a risk, but through maintenance programs, and the availability of critical spares, this risk can be managed.

Availability of consumables – Mine management must continually be aware of any shortfalls of consumables such as cyanide and explosives and quickly source other suppliers.

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26.0

RECOMMENDATIONS

As mining progresses, continued reconciliation needs to be reviewed and if parameters change, an update of the mine plan should be developed.

Operating costs should be reviewed on a regular basis to ensure operating cutoffs are still valid.

It is recommended that the south east area of the open pit be reviewed to optimize the amount of recovered gold ounces. The pit was reduced in size slightly to account for a waste dump and dewatering lines.

It is also recommended that the LT model be updated to incorporate the additional drilling (43 drill holes totalling 5,144 m: holes SF-530 to SF-572) by Apoena that were not available for the September 30, 2011 Mineral Resource estimate.

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27.0

REFERENCES

GeoSystems International, 2006, São Francisco Gold Project, Resource Update, March 8, 2006.

Journel, A G, 1983. Nonparametric estimation of spatial distributions, Math. Geol Vol 15, No. 3, p. 445-468.

Kappes, Cassiday & Associates, 2004, São Francisco Mine Feasibility Study, Gravity/Heap Leach Gold Mine, prepared for Yamana Desenvolvimento Mineral, S.A., November 2004.

Moreno, Rogerio. Technical Report – Resources and Reserves on the São Francisco Mine Located in Mato Grosso state in Brazil, March 30, 2011, 65 pp.

Mello, R., and Petter, R., 2008, São Francisco Mine, Serra da Borda Mineração e Metalurgia S.A. (SBMM), as at February 2008, NI 43-101 Technical Report by NCL Brasil for Yamana Gold Inc., February 2008, 125 pp.

Moore, C. And Lecuyer, N., 2009, Amended and Restated Technical Report on the São Francisco Mine, Mato Grosso State, Brazil, by Scott Wilson RPA, Amended July 27, 2009.

NCL Brasil and Yamana Gold Inc., 2008, São Francisco Mine, Serra Borda Mineração e Metalurgia, SBMM, NI 43-101 Technical Report, February 2008.

Pitard, F.F., 2004, Review of Sampling Systems and Sampling at the Yamana Gold Inc. Operations in Brazil, An Internal Report to Yamana Gold Inc. by Francis Pitard Sampling Consultants, March 12, 2004, 18 pp.

Pitard, F.F., 2007, Review of Sampling Systems and Sampling Practices at the Yamana Gold Inc. São Francisco Operations in Brazil, An Internal Report to Yamana Gold Inc. by Francis Pitard Sampling Consultants, November 16, 2007, 12 pp.

São Vicente Technical Report for an Open Pit, Gravity Concentration and Heap Leach Mine in Brazil – Rezende Eng. Minas S/S Ltda. November 2006.

Silva, A.H., 2008, Geostatistical Simulation for Grades Uncertainty Assessment in São Francisco Mine, Understanding the Variability and the Coarse Gold Effect by a Simulation, An Internal Presentation to Yamana Gold Inc., 15 pp.

Smee, B.W., 2008a, A Review of the Commercial and Mine Laboratories Used by Yamana Gold Inc., An Internal Report prepared for Yamana Gold Inc. by Smee and Associates Consulting Ltd., June 2008, 106 pp.

Smee, B.W., 2008b, Results of a Grade Control Sampling Review and Mine Laboratory Audit São Francisco Mine, Mato Grosso, Brazil, An Internal Report prepared for Yamana Desenvolvimento Mineral S.A., by Smee and Associates Consulting Ltd., October 2008, 61 pp.

MICON, 2011, São Francisco Mineral Resource Estimate, prepared for Aura Minerals Inc, August 24, 2011.

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28.0

DATE AND SIGNATURE PAGE

Signature Date:	January 31, 2012
Effective Date:	September 30, 2011

“J. Britt Reid” (Signed)

J. Britt Reid P. Eng.
Executive Vice President and Chief Operating Officer of Aura Minerals Inc.

Responsible for sections 1.1, 1.2, 1.5, 1.11, 1.12,1.13, 1.15, 1.16, 2 - 5, 13, 17 - 20, and 22 - 27

“Bruce Butcher” (Signed and Sealed)

Bruce Butcher P. Eng.
Vice President, Technical Services of Aura Minerals Inc.

Responsible for sections1.8, 1.9, 1.10, 15, 16 and 21

“Chris Keech” (Signed and Sealed)

Chris Keech P. Geo.
Principal Geologist, CGK Consulting Services Inc.

Responsible for sections 6, 7, 8, 9, 10, 11, 12, 14 and 24, as well as those portions of sections 1, 25 and 26 that pertain to the foregoing sections;

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CERTIFICATE OF QUALIFIED PERSON

As an author of this Technical Report on the São Francisco Mine for Aura Minerals Inc. (“Aura Minerals”), I, J. Britt Reid, do hereby certify that:

This certificate pertains to the Technical Report dated January 31, 2012 and entitled “Resource and Reserve Estimates on the São Francisco Mine in the Municipality of Vila Bela da Santíssima Trindade, State of Mato Grosso, Brazil.” The effective date of the report is September 30, 2011 which is the date of the mineral resource and mineral reserve estimates;

I am employed as the Executive Vice President & Chief Operating Officer of Aura Minerals Inc. Suite 1950 – 777 Dunsmuir Street, Vancouver, British Columbia, Canada V7Y 1K4, tel. 604 669-4777, email breid@auraminerals.com;

I am a member of the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC, member #12061). I graduated from the University of British Columbia with a Bachelor of Applied Science in Mineral Engineering in 1974;

I have practiced my profession continuously since 1974 and have been involved in the design, construction and operation of mines in Canada, United States, Mexico, Brazil and Chile;

I am familiar with Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and, by reason of education, experience and professional registration, I fulfill the requirements of a Qualified Person as defined in NI 43-101;

My most recent personal inspections of the São Francisco Mine were from September 29 to 30, 2011 and November 3 to 4, 2011;

I have been employed by Aura Minerals as Executive Vice President and Chief Operating Officer since November 2009. In addition, I currently hold securities in Aura Minerals. Accordingly, I am not independent of Aura Minerals as defined by NI 43-101;

I am responsible for the preparation and review of sections 1.1, 1.2, 1.5, 1.11, 1.12,1.13, 1.15, 1.16, 2 - 5, 13, 17 - 20, and 22 - 27 of this Technical Report;

As of the date of this certificate, to the best of my knowledge, information and belief, the portions of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make this Technical Report not misleading; and

10)

I have read the Canadian National Instrument 43-101 and the portions of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

Dated this 31st day of January, 2012

“J. Britt Reid” (Signed)

J. Britt Reid, B.A.Sc., P.Eng.
Executive Vice President & Chief Operating Officer of Aura Minerals Inc.

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CERTIFICATE OF QUALIFIED PERSON

As an author of this Technical Report on the São Francisco Mine for Aura Minerals Inc. (“Aura Minerals”), I, Bruce Butcher, do hereby certify that:

I am employed as the Vice President, Technical Services of Aura Minerals Inc., Suite 1950 – 777 Dunsmuir St. Vancouver, British Columbia, Canada, tel. (604) 669-4777, e-mail bbutcher@auraminerals.com

I am a member of the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC, member #29970). I graduated from Queens University with a Bachelor of Applied Science degree in Mining Engineering in 1988.

I have practiced my profession continuously since 1988 and have been involved in the design, construction and operation of mines in Canada, Chile, Mexico, Honduras, Brazil, the United States, and Peru.

My most recent personal inspection of the São Francisco Mine was from February 21 to 22, 2011.

I have been employed by Aura Minerals as Vice President, Technical Services since September 2010. In addition, I currently hold securities in Aura Minerals. Accordingly, I am not independent of Aura Minerals as defined by NI 43-101;

I am responsible for the preparation and review of sections 1.8, 1.9, 1.10, 15, 16 and 21 of this Technical Report;

10)

I have read the Canadian National Instrument 43-101 and the portions of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

Dated this 31st day of January, 2012

“Bruce Butcher” (Signed and sealed)

Bruce Butcher, B.Sc., P.Eng.

Vice President, Technical Services of Aura Minerals Inc.

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CERTIFICATE OF QUALIFIED PERSON

As an author of this Technical Report on the São Francisco Mine for Aura Minerals Inc. (“Aura Minerals”), I, Christopher Gerald Keech, do hereby certify that:

I am employed as the Principal Geologist by, and carried out this assignment for, CGK Consulting Services Inc., 15410 95A Avenue, Surrey, British Columbia, Canada, tel. (604) 721-9745, e-mail chris.g.keech@gmail.com;

I am a member of the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC, member #27185). I graduated from M^cMaster University with a Bachelor of Science in Geology degree in 1980;

I have practiced my profession continuously since 1981 and have been involved in the estimation of mineral resources of precious and base metal deposits in Canada, United States, Australia, Mexico, and Chile;

My most recent personal inspection of the São Francisco Mine was from October 19 to November 5, 2011;

Prior to the completion of this Technical Report, I held the position of Manager of Geostatistics with Aura Minerals for more than two years, from Sept 8, 2009 to October 15, 2011. I directly supervised the work pertaining to the mineral resource estimate for the São Francisco Mine that was carried out by Aura Minerals employees. In addition, I currently hold securities in Aura Minerals. Accordingly, I am not independent of Aura Minerals as defined by NI 43-101;

I am responsible for the preparation of sections 6, 7, 8, 9, 10, 11, 12, 14 and 24 of this Technical Report, as well as those portions of sections 1, 25 and 26 that pertain to the foregoing sections;

10)

I have read the Canadian National Instrument 43-101 and the portions of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

Dated this 31st day of January, 2012

“Christopher G. Keech” {Signed and sealed}

Christopher G. Keech, B.Sc., P.Geo.
Principal Geologist of CGK Consulting Services Inc.

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APPENDIX 1

GLOSSARY OF MINING AND RELATED TERMS AND ABBREVIATIONS

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GLOSSARY OF MINING AND RELATED TERMS AND ABBREVIATIONS

The following is a glossary of certain mining terms and abbreviations that may be used in this Technical Report.

A

AA	Atomic absorption – assay technique method

Adit	A horizontal underground opening, open to the atmosphere at both ends.

Assay	A chemical test performed on a sample of ores or minerals to determine the amount of valuable metals contained.

Au	Gold.

Aura Minerals	Aura Minerals Inc., including, unless the context otherwise requires, the Company's subsidiaries.

B

Blasthole	A drill hole in a mine that is filled with explosives in order to blast loose a quantity of rock.

Bulk sample	A large sample of mineralized rock, frequently hundreds of tonnes, selected in such a manner as to be representative of the potential mineral deposit (orebody) being sampled and used to determine metallurgical characteristics.

Bullion	Precious metal formed into bars or ingots.

By-product	A secondary metal or mineral product recovered in the milling process.

C

CSS	Closed side setting – setting size for a crusher, which represents the maximum particle size that can pass through the crusher.

Channel sample	A sample composed of pieces of vein or mineral deposit that have been cut out of a small trench or channel, usually about 10 cm wide and 2 cm deep.

Chip sample	A method of sampling a rock exposure whereby a regular series of small chips of rock is broken off along a line across the face, back or wall.

CIM	The Canadian Institute of Mining, Metallurgy and Petroleum

CIM Standards	The CIM Definition Standards on Mineral Resources and Mineral Reserves adopted by CIM Council from time to time. Last revised on December 11, 2005 and currently undergoing a review process for a further update possibly in 2011.

Concentrate	A fine, powdery product of the milling process containing a high percentage of valuable metal

Contact	A geological term used to describe the line or plane along which two different rock formations meet.

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Core	The long cylindrical piece of rock, about an inch in diameter, brought to surface by diamond drilling.

Core sample	One or several pieces of whole or split parts of core selected as a sample for analysis or assay.

Custom smelter	A smelter which processes concentrates from independent mines. Concentrates may be purchased or the smelter may be contracted to do the processing for the independent company.

Cutoff grade	The lowest grade of mineralized rock that qualifies as ore grade in a given deposit, and is also used as the lowest grade below which the mineralized rock currently cannot be profitably exploited. Cutoff grades vary between deposits depending upon the amenability of ore to gold extraction and upon costs of production.

CV	The standard deviation divided by the mean.

Cyanidation	A method of extracting exposed gold or silver grains from crushed or ground ore by dissolving it in a weak cyanide solution. May be carried out in tanks inside a mill or in heaps of ore out of doors.

Cyanide	A chemical species containing carbon and nitrogen used to dissolve gold and silver from ore.

D

Dacite	The extrusive (volcanic) equivalent of quartz diorite.

Dilution	Rock that is, by necessity, removed along with the ore in the mining process, subsequently lowering the grade of the ore.

Diorite Dip	The angle at which a vein, structure or rock bed is inclined from the horizontal as measured at right angles to the strike.

° C	Degrees celcius.

E

Epithermal	Hydrothermal mineral deposit formed within one kilometre of the earth’s surface, in the temperature range of 50–200°C.

Epithermal deposit	A mineral deposit consisting of veins and replacement bodies, usually in volcanic or sedimentary rocks, containing precious metals or, more rarely, base metals.

Exploration	Prospecting, sampling, mapping, diamond drilling and other work involved in searching for ore.

F

Fault	A break in the Earth's crust caused by tectonic forces which have moved the rock on one side with respect to the other.

Fold	Any bending or wrinkling of rock strata.

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Footwall	The rock on the underside of a vein or mineral deposit.

Fracture	A break in the rock, the opening of which allows mineral-bearing solutions to enter. A "cross-fracture" is a minor break extending at more-or-less right angles to the direction of the principal fractures.

G

g/t	Grams per metric tonne.

Galena	Lead sulphide, the most common ore mineral of lead.

gpt	Grams per tonne.

Grade	Term used to indicate the concentration of an economically desirable mineral or element in its host rock as a function of its relative mass. With gold, this term may be expressed as grams per tonne (g/t) or ounces per tonne (opt).

Gram	0.0321507 troy ounces.

H

Hanging wall	The rock on the upper side of a vein or mineral deposit.

High grade	Rich ore. As a verb, it refers to selective mining of the best ore in a deposit.

Host rock	The rock surrounding a mineral deposit or orebody.

HQ	Drill core with a 63.5 millimetre diameter.

Hydrothermal	Processes associated with heated or superheated water, especially mineralization or alteration.

I

Indicated Mineral Resource	An Indicated Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough for geological and grade continuity to be reasonably assumed.

Inferred Mineral Resource	An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes.

Intrusive	A body of igneous rock formed by the consolidation of magma intruded into other

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in	inch.

K

km	Kilometre(s). Equal to 0.62 miles.

kg/t	kilograms per tonne.

kg	kilograms (metric).

L

Leaching	The separation, selective removal or dissolving-out of soluble constituents from a rock or ore body by the natural actions of percolating solutions.

Limestone	A bedded, sedimentary deposit consisting chiefly of calcium carbonate.

M

m	Metre(s). Equal to 3.28 feet.

Marble	A metamorphic rock derived from the recrystallization of limestone under intense heat and pressure.

Measured Mineral Resource	A Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough to confirm both geological and grade continuity.

Metallurgy	The science and art of separating metals and metallic minerals from their ores by mechanical and chemical processes.

Metamorphic	Affected by physical, chemical, and structural processes imposed by depth in the earth’s crust.

Mill	A plant in which ore is treated and metals are recovered or prepared for smelting; also a revolving drum used for the grinding of ores in preparation for treatment.

Mine	An excavation beneath the surface of the ground from which mineral matter of value is extracted.

Mineral	A naturally occurring homogeneous substance having definite physical properties and chemical composition and, if formed under favourable conditions, a definite crystal form.

Mineral Claim	That portion of public mineral lands which a party has staked or marked out in accordance with federal or state mining laws to acquire the right to explore for and exploit the minerals under the surface.

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Mineral Deposit	A mass of naturally occurring mineral material, e.g. metal ores or non metallic minerals, usually of economic value, without regard to mode of origin.

Mineralization	The process or processes by which mineral or minerals are introduced into a rock, resulting in a valuable or potentially valuable deposit.

Mineral Resource	A concentration or occurrence of natural, solid, inorganic or fossilized organic material in or on the earth's crust in such form and quantity and of such grade or quality that it has reasonable prospects for economic extraction. The location, quantity, grade, geological characteristics and continuity of a mineral resource are known, estimated or interpreted from specific geological evidence and knowledge. The term mineral resource covers mineralization and natural material of intrinsic economic interest which has been identified and estimated through exploration and sampling and within which mineral reserves may subsequently be defined by the consideration and application of technical, economic, legal, environmental, socio-economic and governmental factors. The phrase reasonable prospects for economic extraction imply a judgment by the Qualified Person in respect of the technical and economic factors likely to influence the prospect of economic extraction. A mineral resource is an inventory of mineralization that under realistically assumed and justifiable technical and economic conditions might become economically extractable. The term mineral resource used in this Technical Report is a Canadian mining term as defined in accordance with NI 43-101 – Standards of Disclosure for Mineral Projects under the guidelines set out in the Canadian Institute of Mining, Metallurgy and Petroleum (the CIM), Standards on Mineral Resource and Mineral Reserves Definitions and guidelines adopted by the CIM Council on December 11, 2005 (the CIM Standards).

Mt	Millions of tonnes, with a metric ton of 1,000 kilograms (2,205 pounds).

N

NaCN	Sodium Cyanide.

Net Smelter Return (NSR)	A payment made by a producer of metals based on the value of the gross metal production from the property, less deduction of certain limited costs including smelting, refining, transportation and insurance costs.

NI 43-101	National Instrument 43-101is a national instrument for the Standards of Disclosure for Mineral Projects within Canada. The Instrument is a codified set of rules and guidelines for reporting and displaying information related to mineral properties owned by, or explored by, companies which report these results on stock exchanges within Canada. This includes foreign-owned mining entities that trade on stock exchanges overseen by the Canadian Securities Administrators (CSA), even if they only trade on Over The Counter (OTC) derivatives or other instrumented securities.

NQ	Drill core with a 47.6 millimetre diameter.

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O

OK	Ordinary kriging – geostatistical method of estimating values.

Open Pit	A mine that is entirely on surface. Also, referred to as open-cut or open-cast mine.

Ore	Mineralized material that can be extracted and processed at a profit.

Orebody	A term used to denote the mineralization contained within an economic mineral deposit.

Ounce	A measure of weight in gold and other precious metals, correctly troy ounces, which weigh 31.1 grams as distinct from an imperial ounce which weigh 28.4 grams.

Outcrop	An exposure of rock or mineral deposit that can be seen on surface, which is, not covered by soil or water.

Oxidation	A chemical reaction caused by exposure to oxygen that results in a change in the chemical.

oz	Ounce.

P

Plant	A building or group of buildings in which a process or function is carried out; at a mine site it will include warehouses, hoisting equipment, compressors, maintenance shops, offices and the mill or concentrator.

Ppm	parts per million.

Pyrite	A common, pale-bronze or brass-yellow, mineral. Pyrite has a brilliant metallic luster and has been mistaken for gold. Pyrite is the most wide-spread and abundant of the sulphide minerals and occurs in all kinds of rocks.

Q

Qualified Person	Conforms to that definition under NI 43-101 for an individual: (a) to be an engineer or geoscientist with at least five years' experience in mineral exploration, mine development or operation or mineral project assessment, or any combination of these; (b) to have experience relevant to the subject matter of the mineral project and the technical report; and (c) to be a member in good standing of a professional association that, among other things, is self-regulatory, has been given authority by statute, admits members based on their qualifications and experience, requires compliance with professional standards of competence and ethics and has disciplinary powers to suspend or expel a member.

QA/QC Quality Assurance / Quality Control	Systems and methodologies to ensure samples and data have good confidence of accuracy.

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NI 43-101 Technical Report


R

Reclamation	The restoration of a site after mining or exploration activity is completed.

Recovery Rate	A term used in process metallurgy to indicate the proportion of valuable material obtained in the processing of an ore. It is generally stated as a percentage of the material recovered compared to the total material present.

Refining	The final stage of metal production in which impurities are removed from the molten metal.

Refractory ore	Ore that resists the action of chemical reagents in the normal treatment processes and which may require pressure leaching or other means to effect the full recovery of the valuable minerals.

Rod mill	A steel cylinder filled with steel rods into which crushed ore is fed. The rod mill is rotated, causing the balls to cascade and grind the ore.

RQD Rock Quality Designation –	A system which ranks rock strength. From the RQD index the rock mass can be classified as follows:

	RQD	Rock mass quality
	<25%	very poor
	25-50%	poor
	50-75%	fair
	75-90%	good
	90-100%	excellent

S

Shoot	A concentration of mineral values; that part of a vein or zone carrying values of ore grade.

Skarn	Name for the metamorphic rocks surrounding an igneous intrusive where it comes in contact with a limestone or dolostone formation.

SMU	Is short for selective mining unit - the smallest tonnage of material that can be separated from waste and mined as ore.

Sphalerite	A zinc sulphide mineral; the most common ore mineral of zinc.

Stockpile	Broken ore heaped on surface, pending treatment or shipment.

Strike	The direction, or bearing from true north, of a vein or rock formation measure on a horizontal surface.

Stringer	A narrow vein or irregular filament of a mineral or minerals traversing a rock mass.

São Francisco Mine

NI 43-101 Technical Report


Sulphides	A group of minerals which contains sulphur and other metallic element such as copper and zinc. Gold and silver is usually associated with sulphide enrichment in mineral deposits.

T

Tailings	Material rejected from a mill after most of the recoverable valuable minerals have been extracted.

Tailings pond	A low-lying depression used to confine tailings, the prime function of which is to allow enough time for heavy metals to settle out or for cyanide to be destroyed before water is discharged into the local watershed.

Tonne	A metric ton of 1,000 kilograms (2,205 pounds).

t	A metric tonne.

t/h	Tonne per hour.

tph	tonnes per hour.

t/m3	Tonnes per cubic metre.

t/d	Tonnes per day.

tpd	tonnes per day.

Tunnel	A horizontal underground opening, open to the atmosphere at both ends.

V

Vein	A fissure, fault or crack in a rock filled by minerals that have travelled upwards from some deep source.

W

Wall rocks	Rock units on either side of a mineral deposit (orebody). The hanging wall and footwall rocks of a mineral deposit (orebody).

Waste	Unmineralized, or sometimes mineralized, rock that is not minable at a profit.

Z

Zone	An area of distinct mineralization.