ex99_1.htm

EXTORRE GOLD MINES LIMITED

Cerro Moro Gold-Silver Project

Second Preliminary Economic Assessment Technical Report NI 43-101

Prepared by

GR Engineering Services Limited

Aug 11

Title Page

Project Name:	Cerro Moro Gold-Silver Project
Title:	Second Preliminary Economic Assessment Technical Report
Location	Santa Cruz Province, southern Argentina
Effective Dates:
Effective Date of Technical Report:	August 2nd, 2011
Effective Date of Mineral Resources:	May 31st, 2011

Qualified Persons:

1.	Bill Gosling, Senior Process Metallurgist, of GR Engineering Services Ltd (FAusIMM), was responsible for the information provided for the metallurgy and process plant design;

2.	David (Ted) Coupland (BSc DipGeoSc CFSG ASIA MAusIMM CPGeo MMICA) Director – Geological Consulting - Principal Geostatistician, Cube Consulting Pty Ltd, was responsible for resource estimation, exploration, drilling and data verification;

3.	Eduardo Rosselot, CEng The Institute of Materials, Minerals and Mining (CEng MIMMM, Membership Nº448843), with NCL Ltda, was responsible for the mining related studies and economic valuation.

Other Expert Persons:

●	Dominic Piscioneri, Mechanical Engineer, employed by GR Engineering Services Ltd, led the engineering definition study related to processing plant design;

●	Dante Cramero, Manager - Environmental for Extorre Gold Mines Ltd, was responsible for the information provided for the environmental sections of the report

●	Gabriel Valero, Ausenco Vector, was responsible for the information related to the design and costing of the tailings storage facility.

Sections updated from PEA issued December 2010:

●	Jorge Gomez, Electrical Engineer employed by Servicios de Ingenieria Electrica y Electromecanica SRL, led the preliminary studies on power line design and capital costs

●	Mario Cuello, Geologist, employed by Ausenco Vector SA, led the preliminary studies on hydrology and hydrogeology, infrastructure, mine closure plans and costs, and environmental and social issues.

●	Guillermo Albornoz, Senior Mining Engineer employed by Antonio Karzulovic & Asoc. Ltda, led the preliminary studies related to geomechanical characterization, pit slope angles, underground stope designs, and crown pillar calculations

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

● CONSENT OF AUTHOR

British Columbia Securities Commission

Alberta Securities Commission

Saskatchewan Financial Services Commission

The Manitoba Securities Commission

Ontario Securities Commission

New Brunswick Securities Commission

Nova Scotia Securities Commission

Prince Edward Island Securities Office

Securities Commission of Newfoundland and Labrador

Re: Extorre Gold Mines Limited1

I, Bill Gosling, Process Metallurgist, have prepared the technical report titled Cerro Moro Gold-Silver Project2, Second Preliminary Economic Assessment, Technical Report NI 43-101 dated August 2nd, 2011 (the “Technical Report").

a)	I hereby consent to the public filing of the Report, the written disclosure of the Report and to the use of my name, Bill Gosling, in reference to this Report.

Dated this 2nd Day of August 2011

____________________________

Bill Gosling

GR Engineering Services Limited

(FAusIMM)

1 Extorre Gold Mines Limited (Extorre) was established from a spinoff of Exeter Resources Corporation (Exeter) Argentinean assets in March 2010, all references to Exeter within this document refer to pre-2010. All references to Estelar Resources Limited refer to the 100% owned Argentinean subsidiary of Extorre.

2 All references to Cerro Moro or the Cerro Moro Project contained in this document refer to the Cerro Moro Gold-Silver Project.

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

TABLE OF CONTENTS

SUMMARY

2.	INTRODUCTION

3.	RELIANCE ON OTHER EXPERTS

4.	PROPERTY DESCRIPTION AND LOCATION

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

HISTORY

7.	GEOLOGICAL SETTINGAND MINERALIZATION

8.	DEPOSIT TYPES

9.	EXPLORATION

10.

DRILLING

11.	SAMPLE PREPARATION, ANALYSIS AND SECURITY

12.	DATA VERIFICATION

13.	MINERAL PROCESSING AND METALLURGICAL TESTING

14.	MINERAL RESOURCE ESTIMATES

15.	MINERAL RESERVE ESTIMATES

16.	MINING METHODS

17.	RECOVERY METHODS

18.	PROJECT INFRASTRUCTURE

19.	MARKET STUDIES AND CONTRACTS

20.	ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

21.	CAPITAL AND OPERATING COSTS

22.	ECONOMIC ANALYSIS

23.	ADJACENT PROPERTIES

24.	OTHER RELEVANT DATA AND INFORMATION

25.	INTERPRETATION AND CONCLUSIONS

26.	RECOMMENDATIONS

27.	REFERENCES

28.	DATE AND SIGNATURES

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

SUMMARY

1.1	Introduction

This technical report describes the Cerro Moro Gold-Silver Project (the “Project”), a mineral exploration and development area located in the Santa Cruz Province, Argentina which is held by Extorre Gold Mines Limited (“Extorre”).

Extorre had previously produced a Preliminary Economic Assessment (“PEA”) for the Cerro Moro Gold-Silver Project in December 2010. Subsequent to completion of this second PEA (“study”) and further exploration, the following opportunities to improve project economics were identified:

	●	Completion of infill drilling on existing ore bodies;
	●	Discovery of additional ore bodies enabling total resources to be increased.

This study is based on a revised NI 43-101 resource estimate completed in April 2011 by Cube Consulting Pty Ltd. The study was completed by GR Engineering Services Limited (“GRES”) through the compilation of information generated by consultants and specialists appointed by Extorre for the study. GRES was responsible for the compilation of information and the preparation of the overall study, in addition to the design and costing of the process plant and infrastructure.

1.2	Exploration

The exploration works at Cerro Moro that have been completed since December 2010 include infill drilling and exploration drilling, Table 1 chronologically summarises all exploration work undertaken since December 2010 up until 31st May 2011.

Date	Extorre Exploration Works
Jan 2011	Complete infill drilling Martina, Loma Escondida, Gabriela with the aim to convert Inferred resource category mineralization to higher confidence categories. Exploration drilling to extend Lucia, Gabriela Loma Escondida and Esperanza. Scout drilling at Carolene
Feb 2011 to end March 2011	Exploration drilling to extend Gabriela, Esperanza- Nini, Escondida far west, Michelle, Loma Escondida. Scout drilling Agostina Gabriela NW, Belen and Zoe (discovery of Zoe) Regional drilling Union domes
April 2011 to end May 2011	Infill drilling at Carla and Gabriela SE. Exploration drilling at Zoe, Tres Lomas and Gabriela. Scout drilling Agostina, Deborah parallel, Loma Escondida and Zoe extensions

Table 1 Summary of Exploration Work undertaken by Extorre to May 31, 2011

1.3	Geology and Mineralisation

The polymetallic gold-silver mineralisation at Cerro Moro is of the low sulphidation epithermal vein type. Individual prospects vary from simple, single veins to complex vein systems with spur and

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cymoid loop structures. Limited quartz stockwork veinlets are also present peripheral to the main veins.

The mineralisation is hosted in sub-volcanic and volcanic rocks, which are interpreted to belong to the Jurassic age Chon Aike and La Matilde Formations. In general, the project area is composed of many structural blocks produced by a conjugate set of north-west and north-east faults. Both fault directions contain mineralised quartz veins. In the eastern and central part of the project, north-easterly trending faults control the mineralisation at the Deborah, Belen, Maria, Michelle, Barbara and Ana prospects. In the central and western part of the project, north-westerly trending structures are the dominant control to the mineralisation at the Escondida, Carla, Esperanza, Nini, Gabriela, Carolene, Lucia, Dora, Moro, Florencia, Natalia, Tres Lomas and Lala prospects. The Escondida Structure has a much larger segment of east west strike in the area of the recently discovered Zoe Shoot. Several secondary east-west structures occur between the primary north-west structures, and these control the mineralisation at the Patricia, Loma Escondida and Carlita prospects. The width of the veins varies generally from 1 to 5 metres, but veins may locally attain widths of up to 10 metres. The strike length of individual veins is variable, generally between 200 metres and 1 kilometre. The Escondida structure has been defined by drilling for more than 8 kilometres, and remains open along strike and also at depth.

Improved geological, structural and mineralisation models developed over the past two years have led to the discovery of numerous high-grade gold-silver polymetallic veins at Cerro Moro. These base metal anomalous Au-Ag veins tend to be recessive and are represented on the surface by either very poorly outcropping quartz vein material or none at all. Extensive cover of Tertiary marine sediments and Quaternary gravels also hinder the location or extensions of some of the veins and targets. Detailed ground magnetic survey data, coupled with IP and resistivity data, has highlighted potential targets that are initially tested via trenching and/or scout drilling, as in the case of the Escondida prospect.

Presently structures in northwest-southeast and east-west orientation are the preferred exploration targets, the latter orientation being interpreted to be due to tensional, pull apart structures resulting from minor strike-slip movement of the larger north-west structures.

The discontinuous ore shoots within the larger structures appear to be locally controlled by changes in strike, which are thought to have produced dilational flexures and jogs allowing for greater fluid flow. Changes in wall-rock lithology, along the structures, may be an important factor in controlling the local strike direction of the structures. It has also been noted that the brittle, felsic units have a tendency to produce stronger stockwork style development around the main structures than the intermediate units.

1.4	Mineral Resource Estimation

The April 2011 mineral resource estimate is the third to be undertaken for the Cerro Moro gold-silver project in Argentina and is discussed in detail in Chapter 14 of this present report. The first mineral resource estimate for Cerro Moro was announced on July 8, 2009 based on work undertaken by

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Snowden Mining Industry Consultants. This was followed by an updated mineral resource estimate announced on April 19, 2010 based on work undertaken by Cube Consulting (Cube). The April 2011 resource estimate is the second mineral resource update undertaken by Cube and forms the basis of this study. This updated resource estimate contains mineral resources in both the Indicated and Inferred category (see Table 1.2 and Table 1.3). Based on the current resource inventory, Extorre’s internal assessment indicates that the Cerro Moro Gold-Silver Project has the potential to support a low capital cost high grade gold mining operation. Of particular importance are the high grade Escondida prospects.

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
Escondida	620,000	18.8	829.2	35.4	374,000	16,530,000	705,000
Loma Escondida	44,000	18.4	919.5	36.8	26,000	1,297,000	52,000
Gabriela	537,000	2.4	371.0	9.9	42,000	6,411,000	170,000
Total	1,201,000	11.5	627.5	24.0	443,000	24,238,000	927,000

Table 2 Cerro Moro Indicated Mineral Resources above 1 ppm Gold Equivalent

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
Escondida	508,000	4.3	164.8	7.6	70,000	2,689,000	123,000
Loma Escondida	13,000	9.7	595.4	21.6	4,000	256,000	9,000
Gabriela	390,000	2.3	394.8	10.2	29,000	4,948,000	128,000
Esperanza	371,000	2.6	175.0	6.1	31,000	2,090,000	72,000
Deborah	579,000	2.4	48.1	3.4	45,000	896,000	63,000
Total	1,861,000	3.0	181.8	6.6	178,000	10,879,000	396,000

Table 3 Cerro Moro Inferred Mineral Resources above 1 ppm Gold Equivalent

The primary focus of drilling activities for the April 2011 mineral resource statement was infill drilling on the Loma Escondida and Gabriela (Central) vein zones, with the objective of maximizing the conversion of existing Inferred Category mineral resources to Indicated Category. Additional inferred resources were also generated through the drilling of extensions to known mineralization at the Escondida and Gabriela SE vein zones.

By May 31, 2011, a total of 1,297 drill holes for 165,652.80 m had been completed at the Cerro Moro Gold-Silver Project. Of this total, approximately 128,043.20 m (77%) corresponds to diamond drilling and the remainder (37,609.60 m / 23%) to RC drilling. Approximately 78% (129,389 m) of the drilling completed to date has been conducted on the Escondida, Gabriela, Esperanza, Loma Escondida, and Deborah prospects.

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1.5	Preliminary Mining Studies

This study is based on Cube Consulting’s April, 2011, 43-101 compliant mineral resource statement for Cerro Moro which includes both Indicated and Inferred Category Mineral Resources.

The reader is cautioned that the mining study is a preliminary assessment and it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary assessment will be realized. No Mineral Reserves have been estimated.

The mining study utilizes an owner operated conventional open pit mining method with backhoe excavators and front end loaders, 50t mining trucks, and auxiliary mobile equipment to support a mining operation that will produce a total material between 3.5 and 6.5 million tonnes per year.

An average operating open pit mining cost of US$2.15 per tonne of mined material has been estimated.

In addition to the open pit study, a portion of the underground mineral resources directly below the open pits were considered to be mined by a cut and fill system. The mining analysis carried out indicate a total of 1.0 million tonnes at 4.79 g Au/t and 352.2 g Ag/t can be considered as underground mine resources.

The underground technique to be used at Escondida and Gabriela (and beneath the remaining open pits) will be a Bench and Fill mining method that requires a primary decline, stope access levels 12 m apart (with individual stopes being 8 m high), ore removal using remote-controlled LHD’s and infilling of the mined-out stopes using a mixture of waste material from the open pits and cement.

The study is categorized as a Preliminary Economic Assessment (“PEA”) instead of a Prefeasibility Study (“PFS”) to facilitate the incorporation of additional mineralized zones. The engineering inputs are to the confidence level of a PFS but a portion of the resource for the study is still inferred. Although the lower confidence resource would not be mined until the latter years of the project they do contribute to the mine life.

The daily throughput in this study is being modeled at 1,000 tonnes/day (Table 4.). An option at 750 tonnes/day was developed considering only measured and indicated resources, to be compared with October 19, 2010 study.

			pp	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	Total
Combined Open Pit / Underground Mine Plan	Escondida	ktonnes	109	280	211	266	142	65	23	1	–	–	1,096
		Au g/t	11.6	19.1	10.9	8.3	8.4	8.7	6.6	9.4	–	–	11.9
		Ag g/t	573.7	802.4	426.7	345.3	398.9	401.9	339.9	248.4	–	–	510.3

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Gabriela	ktonnes	–	38	58	62	222	282	220	149	66	15	1,112
	Au g/t	–	1.9	2.2	2.5	1.9	2.1	2.0	1.5	1.3	0.9	1.9
	Ag g/t	–	337.1	370.4	426.2	292.7	331.9	316.0	244.1	222.8	148.3	307.6
Esperanza	ktonnes	–	–	–	–	–	–	117	141	49	–	307
	Au g/t	–	–	–	–	–	–	1.5	1.2	1.1	–	1.3
	Ag g/t	–	–	–	–	–	–	56.4	111.0	105.6	–	89.4
Deborah	ktonnes	–	–	–	–	–	–	–	–	155	57	212
	Au g/t	–	–	–	–	–	–	–	–	3.0	3.0	3.0
	Ag g/t	–	–	–	–	–	–	–	–	66.6	67.9	67.0
TOTAL	ktonnes	109	318	269	328	365	347	360	291	270	72	2,728
	Au g/t	11.6	17.1	9.1	7.2	4.5	3.3	2.2	1.4	2.2	2.5	5.9
	Ag g/t	574.7	746.1	414.5	360.5	334.2	345.0	233.3	179.6	112.0	84.4	345.8

Table 4 Mine Schedule Summary

A general layout for the Project is shown in Figure 1

Figure 1 Conceptual Mine Development for Escondida

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1.6	Metallurgical Testwork Summary

Metallurgical test work has demonstrated that the various ore types are readily amenable to cyanidation. Given the high silver to gold ratio (~60:1) a Merrill Crowe circuit was selected for precious metal recovery from solution rather than a carbon adsorption and elution circuit.

Filtration of the leached residue resulted in low filtration rates due to ultrafine clay particles in the ore. The ore is however amenable to thickening and counter current decantation thickening has been selected to wash the leach residue.

The Inco cyanide destruction process proved successful in reducing the weak acid dissociable (WAD) cyanide in the process tailings to below 8 ppm. The expected metallurgical recoveries for the ore are 95% for gold and 87.4% for silver, further testwork has been recommended to confirm recoveries and for the incorporation of the Zoe prospect into the project.

1.7	Mineral Processing and Recovery Methods

The process design targeted simplicity consistent with maximising recovery of the precious metals and the utilisation of well established and proven technologies.

The comminution circuit design involves a three stage crushing circuit followed by a single stage closed circuit ball mill. Classifier overflow from the grinding circuit is to be thickened prior to being leached in a five stage leach circuit with a nominal residence time of 48 hours.

The leach residue is to be washed in a 5 stage counter current decantation (CCD) thickening circuit with precious metals recovered from the pregnant solution by a conventional Merrill Crowe circuit. The washed residue from the CCD thickening circuit is to undergo cyanide destruction prior to thickening and tailings disposal.

Flash flotation and centrifugal gravity have been incorporated in the grinding circuit design to recover precious metals into a concentrate that is to be re-ground prior to intensive cyanidation leaching on a batch basis. The leached concentrate is to be filtered and washed and the resulting pregnant solution fed to the Merrill Crowe circuit for precious metal recovery. The washed filter cake is to be repulped and returned to main leaching circuit to maximise the extraction of precious metals.

The processing plant has been designed to treat 1000tpd of ore with maximum head grades of 25 g/t of gold and 1000 g/t of silver. A design availability of 92% (8059 operating hours per year) for the process plant has been selected with standby equipment in critical areas. The process design criteria were developed from testwork and benchmarked data from similar projects.

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1.8	Mine Geotechnical

The geotechnical evaluation and determination of geomechanical parameters were developed by A. Karzulovic & Assoc Ltd (AKL). AKL provided geomechanical parameters for the slopes of the different prospects that would be mined, the underground mining parameters and recommendations; and the stockpiles and waste dumps designs geometries.

The stability analysis was performed using the generalized limit equilibrium method, GLE, and the software Slide 5.0 (Rocscience (2005)), for which there were analyzed all possible alternatives fail, that is, overall slope, interamp angles, combination more than one interamp angles, sectors related to local geological contacts and/or geological faults, etc. At the same time, the probability of failure, PF, was calculated by the method proposed by Duncan (2000).

AKL recommended slopes angle can be summarized as follows:

	●	Escondida West: South-West Wall 68 °
	●	North Eastern Wall 65 °
	●	Escondida Central: South-West Wall 68 °
	●	North Eastern Wall 65 °
	●	Loma Escondida: South-West Wall 65 °
	●	North Eastern Wall 68 °
	●	Esperanza: 68° overall
	●	Gabriela: 68° overall
	●	Deborah: 68° overall

1.9	Plant Site Geotechnical

Vector Argentina SA, undertook geotechnical investigations for Extorre in April 2010. Three test pits and four bore holes were used for the preliminary investigations into the geotechnical conditions for the design of foundations for the process plant.

The process plant area topsoil and surface has low growing shrubs and contains organic matter, the topsoil average depth is 0.40 m. At the bottom of the valley to the east of the proposed ROM pad location where the terrain opens into a plain, below the surface layer of silty sand there is a very wet clay layer starting at 0.80 m extending to 1.80 m below the surface. Below the clay layer there is weathered sedimentary rock sandstone that breaks very easily reaching a depth of 3.20 m.

The central sector of the valley below the layer of topsoil, there is a layer of sedimentary rock very hard limiting excavation to 1.40 m. In the upper valley below the top soil there is a silty sand layer that extends from 0.40 to 0.70 m deep below which there is weathered sedimentary rock sandstone of medium hardness that was able to be removed by backhoe up to 1.70 m.

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For the process plant foundation design all major foundations will require box cut to rock and backfill with suitably compacted engineered fill to provide satisfactory founding of concrete footings. Area slabs and general footings will be founded on gravel hardstands placed over the existing material once top soil has been removed.

1.10	Site Infrastructure

1.10.1

Power Supply

Grid connection has been selected as providing the best combination of capital and operating cost with the least environmental impact. The key parameters of the high-voltage network are:

	●	A new connection station near the intersection of roads 281 and 66 for the connection to the grid. 132 kV overhead transmission line, 71 km long, from this intersection to the substation at the Cerro Moro process plant (see Figure 2).
	●	The process plant substation will be equipped with a single 132/33/13.2 kV step-down transformer of 10 MVA capacity and associated switchgear.
	●	13.2 kV overhead transmission line, 3 km long, from the process plant to the mine.

Figure 2 Overhead Powerline Route to Cerro Moro

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The Cerro Moro Gold-Silver Project has an installed maximum power demand of 4.5MW for the process plant and mine at 1000tpd.

1.10.2

Site Access Roads

Cerro Moro site is currently accessed using existing upgraded farm tracks from provincial road 66. For permanent access to the mine site it was determined the most cost-effective option is to construct a new access road from provincial road 66 direct to the process plant. This access road is approximately 13km long and will be constructed from local gravels to provide suitable access for light vehicles and trucks for delivery of construction equipment and operating requirements.

1.10.3

Communications Link

Communications for telephone and data services will be via satellite link. A satellite dish will be established at Cerro Moro to establish connection to the national telephone network.

1.10.4

Water Supply and Sewerage Treatment

Water supply to the process plant and infrastructure will be provided from water bores located to the north of the plant site. A water treatment plant will be located at the at the plant site and will provide potable water for distribution to offices, amenities and the accommodation camp.

The current water supply to Cerro Moro is transported by truck to tanks on site, this will system will be held in place until such time the borefield and water treatment plant are in operation.

A central sewage plant will be located at the accommodation camp with pumps from site facilities utilised to transfer sewage to the central unit.

1.10.5

Mine Infrastructure

The mine workshop, change house, training center, mine warehouse and fuel storage facilities will be located near the Cerro Moro mine.

These facilities will be designed to undertake the servicing of all underground equipment, the trucks and loaders used for transport of ore to the process plant, the road maintenance fleet and light vehicles.

One magazine located at a convenient and safe location between the ore bodies will be established for the Cerro Moro Mine.

1.10.6

Cerro Moro Administration

The following administration and plant buildings will be newly constructed for the project:

	●	Main administration building with medical centre and training room
	●	Security office and gatehouse

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	●	Laboratory
	●	Metallurgical office/laboratory
	●	Plant crib room and training room
	●	Plant workshop and warehouse
	●	Reagent and sodium cyanide storage

Office space requirements have been defined by the proposed departmental staffing structures.

1.10.7

Accommodation Camps

An accommodation camp will be required to house the construction, exploration and operational personnel for the Cerro Moro Gold-Silver Project. The general philosophy of any requirements for accommodation that are not critical for construction or operation will be met in Puerto Deseado have been applied, with personnel to be bussed as required.

During operations, the process plant and general and administrative staff will be housed in a single-status accommodation camp approximately 1000 m east of the process plant.

The accommodation camp will house the majority of the construction workforce prior to the mobilization of the operational personnel late in the construction period. The accommodation camp has been sized to house 300 personnel during operations. During the construction phase, additional accommodation units will enable approximately 350 personnel to be housed.

1.10.8

Tailings Storage Facility

The process tailings will be stored in large shallow valley area with earth fill embankments providing the necessary enclosure. This tailing storage facility (TSF) will have a geosynthetic liner to prevent seepage into the groundwater system. Surface water will be diverted around the TSF.

The tailings will be pumped from the process plant to the TSF as a slurry, and water will be recovered and pumped back to the process plant to minimize water consumption. The initial capacity is approximately 700,000m³ with the first expansion providing a total storage capacity of 1,500,000m3. The potential final capacity of the TSF, subject to further expansions, is up to 2,540,000m3 for a total of 8.2 years capacity.

1.11	Project Implementation

The study proposes that the process plant, associated facilities and site infrastructure be executed under the management of Extorre’s project team and an EPCM engineer (acting on behalf of Extorre).

Equipment for the project will be sourced from recognized international vendors, while fabrication and installation works will typically be undertaken by experienced Argentinean contractors.

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The following critical key dates and milestones have been identified for the implementation of the project. All dates are subject to receipt of environmental approvals, financing, award of contracts and meteorological conditions. Delays in meeting these target dates would result in a delay to project completion.

Commence Engineering	Q4 2011
Project Financing confirmed	Q4 2011
Commence placement of orders for long lead items	Q1 2012
Environmental Approvals granted	Q1 2012
Commencement of Construction	Q1 2012
Mining operations commence	Q3 2012
Camp completed	Q3 2012
Civil works completed	Q3 2012
All long lead equipment received on site	Q4 2012
Mechanical Completion and Commissioning	Q2 2013

Of particular note is the requirement to place orders for long-lead equipment items, with the implementation schedule critical path determined by delivery of the ball mill. The schedule assumes that much of the work required for project and scope definition is already complete at the commencement of detailed engineering work. As such, after an initial minimum five week period to finalise and confirm plant layouts and project definition document, design works are shown progressing.

1.12	Environmental

Table 5 shows the schedule of the filing of the Environmental Impact Assessment (EIA), granting and notification of the Environmental Impact Statement (EIS) and consequently the expiry date of the environmental permits obtained for the Cerro Moro Gold-Silver Project. Further, it is indicated whether the permit corresponds to an exploration and/or development.

Company	Eia Filing	Eis Grant	Eis Notification	Eis Expiry	Project Stage
MINCORP EXPL. SA	23/05/97	23/06/97	15/07/97	15/07/99	Exploration
MINCORP EXPL. SA	29/06/99	N/D	N/D	N/D	Exploration
CERRO VANGUARDIA SA	08/04/05	20/02/06	24/02/06	24/02/08	Exploration
CERRO VANGUARDIA SA	08/04/05	20/02/06	24/02/06	24/02/08	Exploration
ESTELAR RESOURCES LTD.	09/10/08	30/03/09	28/05/09	28/05/11	Exploration
ESTELAR RESOURCES LTD.	October 2010	13/12/10	15/12/10	15/12/12	Advanced Expl.
ESTELAR RESOURCES LTD.	17/09/2010	16/05/11	16/05/11	16/05/13	Development

Table 5 Schedule of the filing of the Environmental Impact Assessment (EIA)

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1.13	Capital and Operating Cost Estimates

The total initial capital investment for the mining, process plant and infrastructure is summarised in Table 6

Description	Equip / Matl Cost $	Labour Cost $	Freight Cost $	TOTALS $
PROCESS PLANT DIRECT COSTS	38,804,983	11,220,239	2,239,125	52,264,347
INFRASTRUCTURE DIRECT COSTS	37,273,717	1,913,404	874,306	40,061,426
EPCM	6,026,030	14,224,670	0	20,250,700
COMMISSIONING, SPARES & TEMPORARY FACILITIES	3,365,655	3,098,274	182,979	6,646,908
OWNERS & PRE-PRODUCTION COSTS	3,639,667	0	0	3,639,667
TOTAL PROCESS PLANT COST ESTIMATE	89,110,051	30,456,587	3,296,409	122,863,048

CONTINGENCY				13,029,997
VAT				19,879,898
TOTAL INCLUDING VAT				155,772,943

Table 6 Process Plant and Infrastructure CAPEX Summary

The total process plant capital and operating costs are summarised in Table 7 and are shown for each year of operation due to the high variability depending on silver grades in the plant feed. The mining capital and operating costs are summarised in Table 8 showing sustaining capital also.

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Production	Y00	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Total
Tonnes Treated (t)		307,901	335,752	335,662	335,908	335,821	359,893	335,867	332,427	72,391	2,751,622
Gold (g/t)		16.7	9.8	7.4	5.0	3.6	2.4	1.9	2.3	2.5	5.9
Silver (g/t)		733.8	455.8	373.5	354.8	349.0	253.3	203.0	129.5	84.5	345.0
Cost Centre	Unit Cost $/t	Unit Cost $/t	Unit Cost $/t	Unit Cost $/t	Unit Cost $/t	Unit Cost $/t	Unit Cost $/t	Unit Cost $/t	Unit Cost $/t	Unit Cost $/t	Unit Cost $/t
Cost Centre Labour		17.7	16.2	16.2	16.2	16.2	15.1	16.2	16.4	75.2	17.8
Power		5.8	5.3	5.3	5.3	5.3	5.0	5.3	5.4	10.5	5.5
Regents and Grinding		44.6	35.9	34.4	33.8	33.6	30.4	29.5	27.5	31.3	33.5
Media Maintenance		3.8	3.5	3.5	3.5	3.5	3.2	3.5	3.5	16.1	3.8
Linings		1.4	1.3	1.3	1.3	1.3	1.2	1.3	1.3	1.2	1.3
Other		8.4	7.8	7.8	7.8	7.8	7.3	7.8	7.8	32.9	8.5
Total		81.66	69.98	68.44	67.85	67.68	62.22	63.57	61.87	167.13	70.3
Total Opex		25,143,196	23,495,925	22,972,707	22,791,358	22,728,365	22,392,542	21,351,065	20,567,258	12,098,708	193,541,125
Total Capex	155,772,943	(19,879,898)*		3,550,000.0			5,500,000.0				144,943,045
Overall Processing Costs	155,772,943	5,263,298	23,495,925	26,522,707	22,791,358	22,728,365	27,892,542	21,351,065	20,567,258	12,098,708	338,484,170

“ Denotes return of VAT

Table 7 Process Plant Total Capex and Operating Cost Summary

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Production	Y00	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Total
Open Pit Operations
Total CAPEX	30,632,754		0	0	0	0	0	0	0	0	30,632,754
Total Sustaining Capital		4,318,000	0	100,000	548,000	1,010,000	0	0	0	0	5,976,000
Total OPEX		13,515,619	11,358,609	11,251,994	12,622,595	12,856,405	11,181,923	8,851,534	7,217,181	778,946	89,634,805
Total open pit expenses	30,632,754	17,833,619	11,358,609	11,351,994	13,170,595	13,866,405	11,181,923	8,851,534	7,217,181	778,946	126,243,559
Underground Operations
Equipment	0	16,536,000	6,600,000	660,000	601,000	0	0	0	0	0	24,397,000
Other Investments	0	2,016,720	380,000	180,000	0	0	0	0	0	0	2,576,720
Development	0	5,695,827	9,485,269	1,809,022	842,319	0	0	0	0	0	17,832,436
Varios Contingency 10%	0	2,424,855	1,646,527	264,902	144,332	0	0	0	0	0	4,480,616
Total CAPEX	0	26,673,401	18,111,796	2,913,924	1,587,650	0	0	0	0	0	49,286,771
Total OPEX	0	4,692,281	12,494,355	7,569,129	6,598,746	5,577,923	5,063,420	3,672,880	1,773,261	392,644	47,834,638
Total underground expenses	0	31,365,682	30,606,151	10,483,052	8,186,397	5,577,923	5,063,420	3,672,880	1,773,261	392,644	97,121,409
Overall Mining Expenses	30,632,754	49,199,301	41,964,760	21,835,046	21,356,992	19,444,327	16,245,343	12,524,414	8,990,442	1,171,590	223,364,968

Table 8 Mining CAPEX and OPEX Summary

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1.14	Economic Analysis

The total estimated initial capital cost of the project is US$ 208 million which comprises US$ 153 million for plant, infrastructure, energy, tailings dam, EPCM costs and owners costs. The capital allowance for initial underground mine development and equipment is an additional US$ 20 million and US$ 36 million for open pit mine equipment and pre-strip. Adding sustaining capital and closure, the total capital cost for the life of the project amounts to US$ 255 million.

Extorre should bear in mind that NCL is not a financial adviser, and that these models are indicative only, based on NCL’s experiences. NCL recommends that Extorre seeks its own financial and tax advice before taking action in relation to the financial matters rose herein.

Financial Model	US $1320 Gold / US $26 Silver
NPV0 pre-tax	US $ 551.0 million
NPV0 pre-tax with VAT recovery	US $ 581.0 million
NPV0 Free Cash flow (after tax)	US $ 356.4 million
NPV5 pre-tax	US $ 435.7 million
NPV5 Free cash flow (after tax)	US $ 274.4 million
IRR pre-tax	89.3%
IRR Free cash flow (after tax)	58.4%
Period to payback from start of production (at 0% discount)	12 months
Period to payback from start of production (at 5% discount)	15 months

Table 9 Economical Evaluation Results Summary

Sensitivity analysis was undertaken to measure the effect of variations in gold price, discount rate, total operating cost and average tax rate, for after-tax and before-tax cases. The obtained results are very solid, with positive NPVs and attractive IRR for almost all the combinations as contained in Section 22.

Extorre will also seek to apply for tax credits of approximately $US 50 million that relate to pre-mining exploration expenditures, as such credits are permissible under Argentine legislation but have not been included in the current financial analysis.

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2.	INTRODUCTION AND TERMS OF REFERENCE

2.1	Introduction

Extorre Gold Mines Limited1 (“Extorre”) commissioned GR Engineering Services Limited (“GRES”) to prepare a Technical Report as an update to the December 2010 Preliminary Economic Assessment (“PEA”) for the Cerro Moro Gold-Silver Project. This second PEA (“study”) is based on a revised NI 43-101 resource estimate completed in April 2011 by Cube Consulting Pty Ltd.

The Cerro Moro Gold-Silver Project is located 70 kilometers southwest of the port city of Puerto Deseado, Santa Cruz Province, southern Argentina.

The study was completed by GRES through the compilation of information generated by consultants and specialists appointed by Extorre. GRES was responsible for the compilation of information from these experts and specialists and the assembly of the overall study.

This report and the resource estimate have been prepared in compliance with the disclosure and reporting requirements set forth in the Canadian Securities Administrators National Instrument 43-101 (“NI 43-101”), Companion Policy 43-101CP, and Form 43-101F1.

2.2	Qualified Persons

The following qualified persons have compiled this technical report:

	●	Bill Gosling, Senior Process Metallurgist, of GR Engineering Services Limited was responsible for the information provided for the metallurgy and process plant design;
	●	Ted Coupland, Director Geological Consulting, of Cube Consulting, was responsible for resource estimation, exploration, drilling and data verification;
	●	Eduardo Rosselot, CEng The Institute of Materials, Minerals and Mining with NCL Ltda., was responsible for the mining related studies and economic valuation.

Table 10 details the responsibility of the technical report sections to each of the persons involved.

Responsible Person	Company	Part
Ted Coupland	Cube Consulting	7,8,9,10,11,12,14,15, 26
Bill Gosling	GRES	13, 17, 21, 26
Eduardo Rosselot	NCL	16, 21, 22, 26

Table 10 Technical Report Responsibility Matrix

1Extorre Gold Mines Limited (Extorre) was established from a spin off of Exeter Resources Corporation (Exeter) Argentinean assets in March 2010, all references to Exeter within this document refer to pre-2010. All references to Estelar Resources Limited refer to the 100% owned Argentinean subsidiary of Extorre.

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2.3	Frequently Used Acronyms, Abbreviations, Definitions, and Units of Measure

Unless otherwise indicated, all references to currency in this report refer to United States dollars (US$). Frequently used acronyms and abbreviations are listed below.

Abbreviation	Meaning
AAS	atomic absorption spectrometry
Ag	Silver
As	Arsenic
Au	Gold
Cu	Copper
g/t	grams per metric tonne
ha	Hectare
Hg	Mercury
ICMC	International Cyanide Management Code
km	Kilometers
l	liters
m	meters
masl	meters above sea level
Pb	Lead
QA/QC	quality assurance and quality control
RC	reverse-circulation drilling method
RQD	rock-quality designation
Sb	antimony
ºT	degrees relative to true north
t	tonnes
t/a	tonnes per annum
Zn	Zinc

Table 11 Frequently used acronyms and abbreviations

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3.	RELIANCE ON OTHER EXPERTS

3.1	Qualified Persons

The following Qualified Persons have contributed to the Technical Report in the following areas:

	●	Bill Gosling, Senior Process Metallurgist, of GR Engineering Services Ltd, for mineral processing, metallurgy and process plant design.

	●	David (Ted) Coupland, Director Geological Consulting, of Cube Consulting, for mineral resource estimation,

	●	Eduardo Rosselot, CEng The Institute of Materials, Minerals and Mining (CEng MIMMM, Membership Nº448843), with NCL Ltda, for mining related studies and economic evaluation.

GRES has relied on data and information provided by Extorre and on previously completed technical reports. Although GRES has reviewed the available data and visited the site, these activities serve to validate the section of the report that GRES are nominated as the qualified person.

3.2	Other Expert Persons

Other expert persons relied upon for the development of the study:

	●	Dominic Piscioneri, Mechanical Engineer, employed by GR Engineering Services Ltd, led the engineering definition study related to processing plant design;

	●	Dante Cramero, Manager - Environmental for Extorre Gold Mines Ltd, was responsible for the information provided for the environmental sections of the report

	●	Gabriel Valero, Ausenco Vector, was responsible for the information related to the design and costing of the tailings storage facility.

3.3	Sections updated from PEA issued December 2010

The following persons were relied upon for information that was developed for the PEA issued in December 2010; the information provided by these persons has been updated in this study.

	●	Jorge Gomez, Electrical Engineer employed by Servicios de Ingenieria Electrica y Electromecanica SRL, led the preliminary studies on power line design and capital costs

	●	Mario Cuello, Geologist, employed by Ausenco Vector SA, led the preliminary studies on hydrology and hydrogeology, infrastructure, mine closure plans and costs, and environmental and social issues.

	●	Guillermo Albornoz, Senior Mining Engineer employed by Antonio Karzulovic & Asoc. Ltda, led the preliminary studies related to geomechanical characterization, pit slope angles, underground stope designs, and crown pillar calculations.

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Location

4.1

Location

The Cerro Moro Project is located some 70 kilometres (90 kilometres by road) southwest of the port city of Puerto Deseado, Santa Cruz Province, southern Argentina.

Figure 3 Location and Access Map (Source Exeter, 2009a)

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The project area is geographically centred at approximately 48° 01’ 55” south latitude and 66° 33’ 45” west longitude. Access to the Cerro Moro project is via 20 kilometres of paved road (Provincial Highway 281) from Puerto Deseado to the locality of Tellier, followed by 70 km of all weather gravel road (Provincial Route 47) to the project turnoff. The Cerro Moro project area is characterized by low altitudes (70 m to 150 m above mean sea level), flat to undulating relief, and a cool, dry climate typical of the Patagonian pampa (plains).

4.2

Land Area

The Cerro Moro Project currently comprises eighteen tenements covering an area of approximately 177 square kilometres (Figure 4). Two of these tenements are Cateos de Exploracion (“Cateo”), with the remainder being Manifestaciones de Descubrimiento (“MD”).

Figure 4 Exeter Property Tenement Map (Source Exeter, 2009a)

4.3	Company Ownership and Agreements

Mineral rights in Argentina are acquired through application to the government for concessions to either explore for or mine minerals located within a specified parcel of land (and not through ground staking). Generally, all persons or entities qualified to acquire and possess real estate can obtain mineral rights. There are 3 main types of mineral rights and titles, these being described in detail below:

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‘Cateo’ - Before work may commence in an area, an exclusive exploration permit known as a “Cateo” must be obtained. Once an application is submitted all rights to any mineral discoveries on a Cateo by third parties belong to the applicant. A Cateo is measured in 500 hectare (“ha”) units and can range in size from a minimum of 1 unit (500 ha) to a maximum of 20 units (10,000 ha). The approval of a Cateo specifies the area and the term of the Cateo. A one-time fee of $0.80 per ha is due on application for the Cateo. The rights of the Cateo holder are subject to surface rights.

During the term of a Cateo, which begins 30 days after approval, periodic relinquishment of ground is made such that after 300 days from the date of approval, 50% of the area in excess of 4 units must be relinquished and after 700 days, 50% of the remaining area must be relinquished. A Cateo of 1 unit has duration of 150 days and for each additional unit, its duration is increased by an additional 50 days.

‘Manifestacion de Discubrimiento’ – upon discovery of a mineral occurrence within a cateo, the owner can apply for a Manifestacion de Discubrimiento (MD) to protect the discovery. The application for a MD can be made at any time during the term of the cateo but must be made before the expiry of the cateo. The maximum area of one MD is 3,000 ha. Upon verification and approval of the mineral discovery by the authorities, the MD will protect the mineral discovery until such time as the “mensura” (measurement) process begins leading to the eventual grant of a Mina (mining lease).

‘Mina’ - After the size and configuration of a Manifestacion de Discubrimiento are determined, a part or all of it is surveyed and the area applied for a ‘Mina’ or Mining Lease. This is usually done after the results of exploration indicate a potential ore body.

4.4	Title Agreements and Ownership

The Cerro Moro Project currently comprises eighteen tenements covering an area of approximately 177 square kilometres.

Two of these tenements are Cateos de Exploracion (“Cateo”), with the remainder being Manifestaciones de Descubrimiento (“MD”).

A listing of the Cerro Moro tenements is provided in Table 12.

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Number	Letter	Year	Title Holder	Type	Name	Area (Hectares)
407082	M	1993	Estelar (ex-CVSA)*	MD	Bárbara II	420
407083	M	1993	Estelar (ex-CVSA)	MD	Michelle	420
407084	M	1993	Estelar (ex-CVSA)	MD	Michelle II	420
407087	M	1993	Estelar (ex-CVSA)	MD	Bárbara I	420
407088	M	1993	Estelar (ex-CVSA)	MD	Bárbara	420
407101	M	1993	Estelar (ex-CVSA)	MD	Michelle I	420
407102	M	1993	Estelar (ex-CVSA)	MD	Nini	420
412988	M	1995	Estelar (ex-CVSA)	MD	Hansen I	3,000
412989	M	1995	Estelar (ex-CVSA)	MD	Hansen II	3,000
412990	M	1995	Estelar (ex-CVSA)	MD	Hansen III	3,000
412991	M	1995	Estelar (ex-CVSA)	MD	Hansen	2,500
412992	M	1995	Estelar (ex-CVSA)	MD	Nini I	402
412993	M	1995	Estelar (ex-CVSA)	MD	Nini II	408
404908	C	2002	Estelar (ex-CVSA)	MD	La Virginia	699
401961	E	2007	Estelar	MD	Robertino	976
411600	E	2004	Estelar	MD	Williams	74
402342	E	2007	Estelar	Cateo	Edward	185
402343	E	2007	Estelar	Cateo	Matthew	493
Total Hectares						17,677

Table 12 Cerro Moro Tenements

4.4.1

Exeter - CVSA Option Agreement

On the 30th of December, 2003 Cerro Vanguardia Sociedad Anonima (“CVSA”) and Exeter signed an Agreement, granting Exeter the right to undertake exploration and prospecting work on 39 CVSA properties (“Properties”) described in schedule “A” of the Agreement.

The Agreement provides Exeter with the exclusive right to acquire a 100% interest in the Properties contained in four projects (located in three Argentine provinces; Rio Negro, Chubut and Santa Cruz) by incurring exploration expenditures of US$3 million over five years, with minimum expenditures in years one and two of $250,000 and $500,000, respectively, followed by $750,000 in each of years three through five including completing 8,000 metres of drilling. Exeter completed a legal due diligence review over of the Properties in March, 2004. The Agreement does not require Exeter to issue shares or make any cash payment to CVSA, other than a signature fee of US$100,000 which was fully paid in September, 2004.

Exeter exercised its Option to acquire the Properties in 2007, having incurred US$3 million on exploration. Once Exeter completes 10,000 metres of drilling on any project, CVSA has the right to back into a 60% interest in that project by paying Exeter 2.5 times Exeter's expenditures on that project and by carrying Exeter to the completion of a bankable feasibility study. CVSA may earn an additional 10% project interest (to bring its total interest to 70%) by financing Exeter's share of mine

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development costs (to be repaid at an agreed rate). Should CVSA not elect to back-in to a project, its interest will reverts to a 2% net smelter return (NSR) from production on that project.

Exeter completed 10,000 metres of drilling at Cerro Moro at the end of July, 2007 and following data collation and the receipt of drilling assays, delivered to CVSA a detailed report containing all geological data in early September 2007. CVSA then had a 45 day period in which to exercise its back-in right or its interest would revert to a 2% NSR. On the 29th of October, 2007 CVSA notified Exeter that they would not to exercise their back-in right.

4.4.2

Exeter - Fomicruz Agreement

On the 5th of March, 2008 Exeter and Fomento Minera de Santa Cruz Sociedad del Estado2 (“Fomicruz”), announced it had signed a letter of intent (“LOI”) that set out the key terms of a strategic agreement between Exeter and Fomicruz for the future development of the Cerro Moro project which also provides access to Fomicruz’s significant land holding around Cerro Moro.

The LOI states that Exeter and Fomicruz SE will, subject to approval by the parties, enter into a detailed agreement which will include the following terms:

	●	Fomicruz SE will acquire a 5% interest in Exeter’s 177 square kilometre Cerro Moro project;

	●	Exeter will have the right to earn up to an 80% interest in 691 square kilometres of Fomicruz SE exploration properties (see Figure 5) adjoining the Cerro Moro project by incurring US$10 million in exploration expenditures over a number of years;

	●	Exeter will fund all exploration and development costs of the Cerro Moro project and Fomicruz SE will repay an agreed amount of those costs from 50% of Fomicruz SE’s share of net revenue from future operations; and

	●	Exeter will manage the exploration and potential future development on the properties.

Figure 5 Fomicruz Property Tenement Map (Source Exeter, 2009a)

4 Fomicruz is a company owned by the Government of Santa Cruz Province, Argentina

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4.4.3

History of Cerro Moro Acquisition

The Cerro Moro property was discovered in 1993 by Mincorp Explorations S.A. (“Mincorp”), while investigating Landsat Thematic Mapper (“TM”) satellite imagery colour alteration anomalies. The ensuing exploration program, comprising geological mapping, rock chip geochemistry, and drilling led to the discovery of widespread quartz vein structures, variably mineralized, over more than a hundred square kilometre area.

Mincorp collected a total of 2,982 surface samples: 2,163 of these were from trenches and rock chip channel samples, with the remaining 819 samples being select rock chip samples. In addition, Mincorp completed a total of 34 drill holes for 2,593 metres, comprised of 19 diamond drill holes for 1,016 metres and 15 reverse circulation (“RC”) percussion drill holes for 1,577 metres.

In 2001 the rights of the property were transferred to Cerro Vanguardia Sociedad Anonima (“CVSA”) following the corporate takeover of Mincorp.

On the 30th of December, 2003 CVSA and Estelar Resources Ltd (“Estelar”3) signed an Exploration and Option Agreement (“Agreement”) granting Exeter Resource Corporation (“Exeter”) the right to undertake exploration and prospecting work on 39 CVSA properties described in schedule “A” of the Agreement. The Agreement groups the properties into four projects; “Cerro Moro”, “Other Santa Cruz properties”, “Chubut properties” and “Rio Negro properties”. The Agreement provided Exeter with the option (“Option”) to acquire the properties upon incurring $3,000,0004 in expenditures, including completing 8,000 metres of drilling over five years subject to CVSA retaining the right to back-in to a 60% interest in any project once Exeter completed 10,000 metres of drilling on that project. In the event that CVSA exercises its back-in right, it is required to pay Exeter 2.5 times the expenditures, less certain CVSA expenditures, incurred on the project. CVSA can further increase its interest to 70% in the project by funding Exeter’s 30% development share at industry terms with repayment by Exeter on an agreed basis. If CVSA does not exercise its back-in right its interest reverts to a 2% net smelter royalty. This Technical Report details only the exploration activities conducted at Cerro Moro.

In May of 2007, Exeter served notice to CVSA that it was exercising its Option to acquire the properties, having incurred the required exploration expenditures, as detailed in the Agreement. Further, in September of 2007, Exeter served notice of the completion of 10,000 metres of drilling on Cerro Moro, and as required in the Agreement provided CVSA “all relevant geological and technical data and results obtained from the exploration and prospecting works conducted” on the project to that time. This notice triggered CVSA’s once only right to back-in to a 60% project equity interest in Cerro Moro. At the end of October 2007, CVSA gave notice of its decision not to exercise the back-in right and its interest in Cerro Moro has reverted to a 2% net smelter return royalty.

5The Agreement was made between CVSA, Estelar and Exeter. Whilst Exeter agreed to conduct exploration on, and potentially acquire an interest in, the mining properties indicated in the agreement, such exploration activities were carried out by its wholly-owned Argentine subsidiary Estelar Resources Ltd.

6All monetary units are in United States of America dollars (“US$”), unless specified otherwise.

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Since commencing work in June, 2003, Exeter has carried out diverse exploration activities at Cerro Moro including detailed geological mapping, geochemical sampling (rock, soil, LAG), geophysical surveys (ground magnetic and electrical methods), and diamond and reverse circulation drilling.

In March 2009, Exeter and the Fomento Minero de Santa Cruz Sociedad del Estado (“Fomicruz”), the Santa Cruz provincial mining company, signed a definitive agreement over ten Fomicruz concessions located adjacent to the Cerro Moro concessions. By spending a total of $10,000,000 on exploration on these concessions, Exeter has the option to acquire an 80% interest in the Fomicruz properties. In addition, Fomicruz will acquire a 5% participating interest in Exeter’s Cerro Moro project following the granting of all the required exploitation concessions and permits to commence mining.

On February 11, 2010 Exeter announced its intention to spin out its Argentine assets into a new company to be known as Extorre Gold Mines Limited (“Extorre”). This transaction was approved by Canandian regulatory authorities and completed on March 23, 2010. Extorre holds all of Exeter’s former interests in the Cerro Moro and Don Sixto Projects, in addition to its portfolio of exploration projects in Argentine, and received an initial C$25 million from Exeter. All work conducted prior to March, 2010 has been referenced to Exeter within this report.

On April 19, 2010 Extorre Gold Mines Limited (Extorre) announced the second NI 43-101 compliant mineral resource estimate of related to the Cerro Moro gold-silver project in Argentina. This updated resource estimate contains mineral resources in both the Indicated and Inferred category. The April 2010 resource estimate of the Cerro Moro Project is the first undertaken by Cube and follows a previous estimate undertaken by Snowden Mining Industry Consultants announced on 8th July, 2009 (Bargmann et al, 2009). Based on the current resource inventory, Extorre’s internal assessment indicates that the Cerro Moro project has the potential to support a low capital cost high grade gold mining operation. Of particular importance are the high grade Escondida prospects.

The primary focus of drilling activities for the April 2010 mineral resource statement was infill and extensional drilling on the Escondida prospect. Infill drilling to 20m x 20m was completed to upgrade the substantial portions of the Escondida prospect to the Indicated resource category, in those areas where open-pit and underground mining activity is likely to be scheduled in the early stages of a mine development. In addition, resource drilling was also undertaken to define resource extensions to the Escondida prospect. By the 12th of February, 2010 - the data cut-off date for April 2010 resource estimate - a total of 887 drill holes for 97,302.10 m had been completed at Cerro Moro.

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Additional infill drilling was also conducted at the Loma Escondida, Martina, Carla, and Gabriela prospects. An updated NI 43-101 compliant mineral resource estimation for Cerro Moro was completed by Cube Consulting in April 2011, and incorporated results from infill and extensional drilling completed on the Escondida, Loma Escondida, and Gabriela zones up until January, 2011. This resource is discussed in further detail in Part 14.

By May 31, 2011, a total of 1,297 drill holes for 165,652.80 m had been completed at the Cerro Moro Project. Of this total, approximately 128,043.20 m (77%) corresponds to diamond drilling and the remainder (37,609.60 m / 23%) to RC drilling. Approximately 78% (129,389 m) of the drilling completed to date has been conducted on the Escondida, Gabriela, Esperanza, Loma Escondida, and Deborah prospects.

4.5	Environmental Liabilities and Permitting

Based on the Environmental Impact Report (Exploration Stage) submitted by Mincorp Exploraciones SA (1997), five updatings of such document have been submitted between 1997 and 2010 according to File 403.897/M/97 – Province Mining Bureau, Ministry of Economy and Public Works, Province of Santa Cruz.

In September 2010, the Environmental Impact Assessment (“EIA”) for the proposed Cerro Moro Mine Development was submitted to Santa Cruz Authorities. This EIAconsidered the development of an initial 750 tpd mining/processing option for Cerro Moro, and was approved by the competent authority in May 2011; thus, permits were granted to start the construction of the infrastructure required for Cerro Moro Project development and the deposit mining/operation stage.

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

The following information about the Cerro Moro property is taken directly from the technical report by NCL (Dec 2010). This information has not been updated or verlified:

5.1

Access

The Cerro Moro Project is located 70 kilometres (90 kilometres by road) southwest of the port city of Puerto Deseado, Santa Cruz Province, southern Argentina.

The project area is geographically cantered at approximately 48° 01’ 55” south latitude and 66° 33’ 45” west longitude.

Access to the Cerro Moro project is via 20 kilometres of paved road (Provincial Highway 281) from Puerto Deseado to the locality of Tellier, followed by 70 km of all weather gravel road (Provincial Route 47) to the project turnoff.

5.2

Climate

The Cerro Moro Project is situated in a climatic zone that is classified as semi-arid, with temperatures ranging from minimums of minus 10 degrees Celsius to maximums of 20 degrees Celsius. The bulk of the rainfall and occasional snow fall during the winter months (May – September), and the area is affected by moderate to strong winds.

5.3	Local Resources and Infrastructure

The area surrounding the Cerro Moro project is sparsely populated. Farm stations or ‘estancias’ consisting of one to a few houses dot the country side, the majority occurring several kilometres apart. Traditionally, the main source of income in the Project area has been derived from the raising of sheep.

The operating Cerro Vanguardia Gold Mine (“Cerro Vanguardia”), owned and operated by CVSA, is located approximately 130 kilometres (approximately 200 kilometres by road) to the west-southwest of the project.

The nearest regional centres are Puerto Deseado (pop. 17,000), Puerto San Julian (pop. 6,000), Caleta Olivia (pop. 36,000; located 210 kilometres by road to the north-west), and Comodoro Rivadavia (pop. 200,000; located 290 kilometres by road to the north-northwest of the project area. The city of Rio Gallegos (pop. 79,000), the capital of Santa Cruz Province, is located approximately 500 kilometres by road to the south of the project. These major centres are able to provide basic goods and services, in addition to being connected to the national power grid.

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Comodoro Rivadavia and Rio Gallegos are serviced with national airports. A well maintained concrete airstrip is located at Puerto Deseado, serviced via small to mid-size charter aircraft.

5.4	Physiography

The Cerro Moro project area is characterized by low altitudes (70 m to 150 m above mean sea level), flat to undulating relief, and a cool, dry climate typical of the Patagonian pampa (plains).

Cerro Moro has very low relief with an average elevation of 100 metres above sea level. No permanent surface watercourses exist within the Project, and several saline lagoons and saltpans are scattered throughout the area.

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HISTORY

The following information about the Cerro Moro property is taken directly from the technical report by NCL (Dec 2010). This information has not been updated or verified:

6.1	Cerro Moro Project History

In 2001 the rights of the property were transferred to Cerro Vanguardia Sociedad Anonima (“CVSA”) following the corporate takeover of Mincorp.

On the 30th of December, 2003 CVSA and Estelar Resources Ltd (“Estelar”1) signed an Exploration and Option Agreement (“Agreement”) granting Exeter Resource Corporation (“Exeter”) the right to undertake exploration and prospecting work on 39 CVSA properties described in schedule “A” of the Agreement. The Agreement groups the properties into four projects; “Cerro Moro”, “Other Santa Cruz properties”, “Chubut properties” and “Rio Negro properties”. The Agreement provided Exeter with the option (“Option”) to acquire the properties upon incurring $3,000,0002 in expenditures, including completing 8,000 metres of drilling over five years subject to CVSA retaining the right to back-in to a 60% interest in any project once Exeter completed 10,000 metres of drilling on that project. In the event that CVSA exercises its back-in right, it is required to pay Exeter 2.5 times the expenditures, less certain CVSA expenditures, incurred on the project. CVSA can further increase its interest to 70% in the project by funding Exeter’s 30% development share at industry terms with repayment by Exeter on an agreed basis. If CVSA does not exercise its back-in right its interest reverts to a 2% net smelter royalty. This Technical Report details only the exploration activities conducted at Cerro Moro.

7 The Agreement was made between CVSA, Estelar and Exeter. Whilst Exeter agreed to conduct exploration on, and potentially acquire an interest in, the mining properties indicated in the agreement, such exploration activities were carried out by its wholly-owned Argentine subsidiary Estelar Resources Ltd.

8 All monetary units are in United States of America dollars (“US$”), unless specified otherwise.

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In March 2009, Exeter and the Fomento Minero de Santa Cruz Sociedad del Estado (“Fomicruz”), the Santa Cruz provincial mining company, signed a definitive agreement over ten Fomicruz concessions located adjacent to the Cerro Moro concessions. By spending a total of US$10,000,000 on exploration on these concessions, Exeter has the option to acquire an 80% interest in the Fomicruz properties. In addition, Fomicruz will acquire a 5% participating interest in Exeter’s Cerro Moro project following the granting of all the required exploitation concessions and permits to commence mining.

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Escondida prospect. By the 12th of February, 2010 - the data cut-off date for April 2010 resource estimate - a total of 887 drill holes for 97,302.10 m had been completed at Cerro Moro.

Between the 12th of February, 2010, and the 31st of May, 2011, exploration activities at Cerro Moro were focused primarily on testing the resource potential of extensions to the known mineralized structures (Escondida, Loma Escondida, Gabriela, Esperanza, Carla, Deborah, Deborah Parallel, Dora, Lucia, Michelle, Natalia, Nini, Patricia and Tres Lomas). New areas without previous drilling were also tested during this period and included Agostina, Belen, and notably Zoe, where a significant high grade gold-silver discovery has been made on the eastern part of the Escondida structure. Additional infill drilling was also conducted at the Loma Escondida, Martina, Carla, and Gabriela prospects. An updated NI 43-101 compliant mineral resource estimation for Cerro Moro was completed by Cube Consulting in April 2011, and incorporated results from infill and extensional drilling completed on the Escondida, Loma Escondida, and Gabriela zones up until January, 2011. This resource is discussed in further detail in Part 14.

6.2	Recent Exploration at Cerro Morro

During 2010, exploration activities at Cerro Moro were focused on the evaluation of the resource potential of the known mineralized veins, together with the identification of new target areas. On April 19, 2010, Extorre announced an updated NI 43-101 mineral resource estimate for the Cerro Moro deposit, with the bulk of these resources having been identified in the high gold-silver grade Escondida vein. At the time of writing, a total of four diamond drill rigs were operating at Cerro Moro with the primary task of delineating additional gold-silver resources.

The Cerro Moro Project is an advanced stage exploration project in which a number of discrete mineralised vein zones have been identified. Exploration activities are targeting mineralisation identified as having potential for extraction by both open pit and underground mining methods.

6.3	Mining History

Development at the Cerro Moro site is limited to farm buildings, portable accommodation, gravel roads and minor service infrastructure to support exploration activities. There is no historical or small scale artisanal mining on the property.

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GEOLOGY

7.1	Regional Geology

Cerro Moro is geologically located within the Deseado Massif, which is a tectonic block in the central-portion of the Santa Cruz Province, covering an area of approximately 60,000 square kilometres.

The oldest rocks within the Deseado Massif are the Upper Pre-Cambrian and Lower Palaeozoic metamorphics of the La Modesta Formation (also known as the Rio Deseado Complex). This formation is intruded by granites of Lower to Middle Palaeozoic age. These rocks are in turn unconformably overlain by continental sandstone of the La Golondrina and El Tranquilo Formations, which were deposited in a series of graben and half-graben structures.

During the Jurassic and Cretaceous Periods the region underwent extensional tectonics, which initially resulted in the epiclastic and pyroclastic Roca Blanca Formation, followed by the widespread mafic volcanic field of the Bajo Pobre Formation during the Mid-Jurassic. During the Mid and Upper Jurassic these rocks were overlain by felsic and intermediate volcanics and sediments of the Bahia Laura Group. The Bahia Laura Group includes the Chon Aike and La Matilde Formations. The Chon Aike Formation constitutes a thick sequence of rhyolitic ignimbrites, tuffs and volcaniclastics, and is interpreted to host the gold mineralisation at Cerro Moro.

During the Early Tertiary Period – Paleocene Epoch the region was draped with continental and marine sediments. More recently, during the Late Tertiary Period to Pleistocene Epoch, basaltic lava flows were extruded but these are not observed at Cerro Moro (refer to Figure 6).

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Figure 6 Regional Geology and Schematic Columnar Section (Source Exeter, 2009a)

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7.2	Project Area Geology

The Cerro Moro property is interpreted to be underlain by shallow south dipping Jurassic age volcanic stratigraphy broadly assigned to the Bahia Laura Group which is regionally documented to encompass:

the Bajo Pobre Formation; which comprises tuffs, flows and sub-volcanic intrusives of andesitic to basaltic composition,

the Chon Aike Formation; the two principal components of which are the rhyolitic ignimbritic pyroclastic facies plus a late intrusive/extrusive flow-domes facies, and

the Matilde Formation; characterized by more epiclastic and aerially deposited ash rich volcaniclastic deposits that are observed to interdigitate or form a lateral facies variation to the main Chon Aike Formation sequence.

A little over 30% of the project is covered by Tertiary marine sediments and Quaternary gravels.

In 2008 contract geologist Nick Callan mapped the entire Moro project at a scale of 1:10,000 (refer to Figure 7 to Figure 10 for an Interpreted Geology map and accompanying legends). The following geological description is largely taken from his report (Callan, 2008).

Some 14 stratified volcanic units, were defined, that form mappable entities at 1:10,000 scale. These units may be grouped into five broad packages based on inferred ages (refer to Figure 11), composition, lithological characteristics and spatial relationships:

The oldest of these packages is the P1 group, comprising an extensive pile of coarse rhyolite clast-bearing ash-flow tuffs and remnant flow-domes, the upper part of which is punctuated by several closely associated welded ignimbritic horizons of rhyodacitic to locally more dacitic composition. The P1 group was assigned to the Chon Aike Formation.

The P1 group forms an extensive litho-structural domain covering a large part of the property, the southern boundary of which is defined by an arcuate, linked normal fault system formed by the interaction of the northwest striking Escondida Fault and the informally named northeast striking Deborah South Fault. Post-P1 group stratigraphy, showing shallow south to southeast dips, is generally confined to the corresponding down-thrown structural domain lying to the south of this linked normal fault system. Field evidence suggests that this linked fault system was active during deposition of the post-P1 group volcanic stratigraphy.

Unconformably overlying the P1 group is a compositionally distinct volcanic-sedimentary assemblage of andesitic to basaltic composition comprising bedded volcanic sediments, coarse volcanic breccias, flows and possibly concordant sub-volcanic intrusive units. Tentatively this assemblage is inferred to represent a short-lived overlap of the Bajo Pobre Formation with the Chon Aike Formation which dominates the property geology.

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This more intermediate compositional interlude is overlain, again unconformably, by a bedded series of felsic air-fall tuffs and related epiclastic units, characterized by the presence of wood fragments, accretionary lapilli and local hot-spring “sinter”. The latter features are diagnostic of the Matilde Formation of the Jurassic age Bahia Laura Group.

The bedded tuff and epiclastic unit pass gradationally upwards via a thin “transitional” stratified sequence characterized by pumice-bearing vitric tuffs beds into the youngest mapped volcanic group, comprising a crudely stratified sequence of strongly welded ignimbritic tuffs of rhyolitic composition. Callan (2008) also considered this latter group to be part of the Chon Aike Formation.

Hosted within the stratified volcanic pile is a series of co-genetic, texturally diverse rhyolitic units forming simple and composite intrusive flow-domes, as well as dykes on a variety of scales. Emplacement of the rhyolitic units was strongly influenced by structure. Rhyolitic intrusives are for the most part discordant with host volcanics though some locally sub-concordant relations were observed by Callan (2008). Cross-cutting relationships with the younger mapped stratigraphic units indicate much of the rhyolite intrusive activity occurred quite late in the geological evolution of the sector.

The area mapped by Callan (2008) is characterized by complex structural architecture comprising a mesh of steeply dipping to sub-vertical fault structures showing principally northwest to west-northwest strikes, northeast strikes, west-southwest strikes and more northeast to north-northeast strikes. “Several of the major mapped fault structures (e.g.: the west-northwest to northwest striking Escondida Fault, the northeast striking Deborah South Fault, and the northwest striking Deborah-Belen Fault) record evidence of a significant normal component of displacement over much of their kinematic history. Normal displacement occurred on some structures (e.g.: Escondida Fault, Deborah South Fault) synchronously with deposition of the post-P1 sequence and these structures very likely exerted some control on the primary distribution of volcanic units and Jurassic palaeo-topography. Normal or oblique faulting and associated dilation on variably oriented fault structures also likely controlled emplacement and geometry of rhyolite flow-domes and related dykes, and more importantly provided sites for epithermal Au-Ag vein mineralization. Post-mineral movement also occurred, localizing late rhyolitic intrusive activity, offsetting and brecciating vein mineralization, and locally preserving high-level alteration types. Thus a long-lived extensional structural regime prevailed and played an active role during much of the geological development of the area. This extensional regime is now manifest as widespread block-faulting with generally minor dip-slip or oblique slip on most faults, but with a locally more significant normal component of displacement across master faults (e.g.: Escondida Fault, Deborah South Fault).” (Callan, 2008).

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Figure 7 Interpreted Geology and Vein Location Map

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Figure 8 Legend for Figure 7 (Part 1 of 3)

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Figure 9 Legend for Figure 7 (Part 2 of 3)

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Figure 10 Legend for Figure 7 (Part 3 of 3)

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Figure 11 Geological Summary Map

7.3	Deposit Geology and Mineralization

Polymetallic gold-silver mineralization is associated with epithermal veins. High-grade gold and silver mineralisation is strongly associated with the presence of sulphides such as: pyrite, sphalerite, galena, acanthite and chalcopyrite. Detrimental toxic elements such as arsenic and mercury are at relatively low levels. The presence of yellow sphalerite and adularia are indicative of low temperature ore formation. At least two mineralisation pulses have been observed. The first pulse deposited clean white quartz veins with low sulphide content and is generally poorly mineralised. This has been interpreted as the product of shallow, circulating meteoric dominated water with scarce mixing of magmatic water. A second, later pulse, consisting of black silica is rich in sulphides and hosts high-grade mineralisation. Precious metal deposition is interpreted to be the product of mixing of magmatic dominated water with meteoric waters. Boiling textures, vein breccias and repeated quartz overgrowths with sulphidic ginguro bands are also observed. Coarse silver sulphide (acanthite) and

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electrum have been observed in several ore shoots and are common in the high grade shoots hosted in the Escondida structure Figure 12 shows the distribution of the various veins at Cerro Moro.

Figure 12 Cerro Moro Prospect Locations

7.3.1

Escondida

At the Escondida prospect work has concentrated on six main areas, designated here from east to west as Zoe, Martina, Escondida Far East, Escondida East, Escondida Central, Escondida West and Escondida Far West. Exploration drilling commenced on the Central and Western areas. Initially these two areas were thought to be two separate veins, but they have proven to be separate shoots within one continuous structure which traverses the entire property.

The Escondida structure has been followed with drilling for approximately 8 kilometres, and remains open along strike and at depth. The width of the mineralization varies between 0.1 to 5 metres on average. Drilling has shown that the mineralisation continues to a minimum depth of 250 metres below the present land surface. The Escondida structure has in general a northwest-southeast strike where high grade shoots are localised where the structure changes strike to east-west or west-northwest. This is the case for the Central, West and Far West high-grade shoots. The newly discovered Zoe Shoot is hosted in a long segment with east west strike. The Escondida quartz veins occur within this major north-west, pre-mineral fault, dipping between 70 to 85 degrees to the south-west. The fault is interpreted to have pre-, syn- and post-mineralisation movement. To the north of the fault, the foot-wall

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rocks consist of volcanics, and to the south, the hanging-wall rock consists of sandstones. High-grade mineralisation is strongly associated with a relatively late, black silica pulse of quartz which crosscuts and fills breccia voids within an earlier white quartz-adularia phase. The black silica is rich in coarse sulphides including; galena, sphalerite, pyrite, chalcopyrite and acanthite.

Zoe is the name given to a new high grade shoot on the Escondida structure located approximately 3.4 kilometres and 2.4 kilometres to the southeast of the Escondida Far East and Martina shoots, respectively. The style of mineralization at Zoe is similar to the other high grade shoots on the Escondida structure. To date the significant mineralization has been traced for approximately 600 metres along strike and to a depth of approximately 250 metres from surface. The significant mineralisation itself does not outcrop but a small section occurs relatively close to the surface and achieves a greater lateral extent at depth. The main vein is developed in the Escondida fault contact between andesitic sandstone (hangingwall) and rhyodacitic ignimbrites and intrusive felsic breccias (footwall). Mineralization is related to quartz-adularia and black silica material, rich in base metal sulphides. Coarse electrum and gold is observed in some of the high grade intersections. The Zoe vein is a prominent structure with thicknesses varying between 0.5 to 6 metres. The upper near surface portion of the structure is characterized by a narrow low grade shear zone which can contain mineralised fragments brought up from deeper levels.

7.3.2

Gabriela

The Gabriela prospect is a north-west oriented quartz vein dipping between 60 to 80 degrees to the north-east. Recent drilling and discovery of a new shoot to the south east has extended the strike length to 1.3 km. The width of the ore zone varies between 10 to 20 metres in the near surface central-southeastern part of the structure, and 3 to 5 metres to the northwest and south east. Mineralization at Gabriela is substantially more silver-rich, compared to other prospects at Cerro Moro. The vein is mostly characterised by white quartz-adularia with disseminated fine pyrite (and possible acanthite?). The vein has been tested to a maximum depth of 350 metres from surface. Mineralisation cuts several volcanic units.

7.3.3

Esperanza

The Esperanza prospect is a north-west structure with one or more quartz veins, which dip steeply to the north-east. At the north-west end of the prospect the veins have been emplaced within coherent andesite. Towards the south east sector (also termed the Esperanza South-East prospect) the main structure displaces rhyodacite to the north and rhyolite to the south. The mineralisation to the south-east is characterised by a 15 metre wide stock-work zone between several higher grade veins. Recent drilling has extended the Esperanza structure to the north-west as far as the Nini vein and now has a known strike length of approximately 2 kilometres. The mineralisation has been drilled to a vertical depth of approximately 120 metres. The main veins are 1 to 3 metres in width. Gold-silver mineralisation is associated with disseminated fine pyrite, copper, lead, and zinc sulphides within quartz veins. Sulphidic ginguro banding, with pyrite, chalcopyrite and sphalerite within quartz veining is observed in high grade zones.

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The structure is presently open at depth at several points. The north-west extension of the Esperanza structure has recently been tested by drilling, confirming that the Esperanza and Nini prospects are hosted in the same structure. The host rock is a rhyodacitic ignimbrite with fiammes.

7.3.4

Loma Escondida

The Loma Escondida vein is an east-west structure dipping steeply to the north and is hosted entirely within andesite. This vein is located 500 metres north of the Escondida prospect and has been interpreted as a secondary dilational structure formed by strike-slip movement of the major north-west structures. The vein crops out poorly, but has been followed by drilling and trenching for more than 600 metres in length. It is a relatively narrow vein ranging in thickness from 0.30 to 2 metres. Pyrite, galena, sphalerite and acanthite are associated with high-grade shoots.

7.3.5

Silvia

The Silvia prospect consists of several east-west orientated breccia-vein sets located near the intersection of two major structures, the Esperanza-Nini structure and the Barbara structure. Outcrop is sparse, with scattered white quartz float over an area of 200 metres by 350 metres. The mineralisation is complex with interpreted repeated brecciation; pre, syn and post mineralization. There are early quartz-adularia-sulphide (pyrite-galena-sphalerite-electrum) veins which have been brecciated and infilled with dark grey or white silica with minor base metal and coarse marcasite. An interpreted late phase consists of a barren breccia which contains clasts of all the previous events and is filled with a clay rock flour matrix. Vein widths vary between 1 to 4 metres, and to date have been tested to a maximum depth of 125 metres from surface.

7.3.6

Nini

The Nini prospect occurs as a major north-west vein structure which is partially covered by Tertiary sediments and has been traced by drilling and trenching for approximately 1.2 kilometres in length. To the south-east the structure has been traced beneath Tertiary cover to the Esperanza structure indicating that these are the continuation of the same structure. In the north-west the vein dips shallowly to the north-east, with the hanging wall consisting of dacitic ignimbrite and the foot wall of andesitic lavas. In the central and south-eastern sectors of the structure the dip changes to 80 degrees to the north-east and the host rock is intermediate to rhyodacitic ignimbrite, in both the hanging wall and foot wall. The thickness is variable between 0.90 and 3.30 metres. The predominant strike of the structure is north-westerly; however there are indications of potential east-west trending zones, which are presently thought to be more prospective for higher grades and widths. The mineralisation is generally observed within a single quartz vein, however vein breccias and stock-works with disseminated sulphides have been observed in places.

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7.3.7

Moro

The Moro prospect is a north-west striking quartz vein structure with a north-east dip. It can be traced for at least 500 metres, with widths ranging between 0.60 and 2.20 metres. It is hosted in felsic to intermediate tuff and the mineralization is associated with a chalcedonic quartz veining with well developed lattice-blade textures and disseminated pyrite.

7.3.8

Patricia

The Patricia prospect is an east-west, narrow quartz vein structure dipping steeply to the north and emplaced in coherent andesite, and rhyodacitic ignimbrites at depth. The area is partly covered by Tertiary and Quaternary sediments. The structure outcrops for 300 metres, but is interpreted from the ground magnetics to continue for an additional 300 metres to the west and 200 metres to the east. High-grade gold and silver mineralisation has been intersected in trenches and drill holes and is more elevated to the west, where the vein disappears under cover. This mineralisation is associated with copper, lead and zinc sulphides in quartz veins, approximately 1 metre in width. It is currently interpreted that this east-west vein occurs within a similar structural setting to that of the Loma Escondida prospect to the south, and that both are secondary dilational structures formed by strike-slip movement of the major north-west structures such as Escondida and Esperanza.

7.3.9

Deborah

The Deborah prospect is a north-east trending structure with a moderate north-west dip. The hanging wall rocks consist of felsic ignimbrite, and footwall rocks consist of intermediate to rhyodacitic ignimbrite. The vein has a total length of 700 metres and mineralisation has been tested to approximately 80 to 100 metres vertical depth. The thickness is variable between 0.50 to 5 metres.

The south-east portion of the structure continues under cover but can be traced for an additional 250 metres in the ground magnetics. To the north-east the structure has been truncated by a north-west trending fault with apparent dextral displacement. The better gold-silver grades to date are located at the north-eastern end and are associated with white quartz vein breccias with a black silica matrix and disseminated sulphides (mostly pyrite). This truncating north-west trending structure (designated as the Deborah Termination Structure, “DTS”), clearly observed in the ground magnetic data, was originally interpreted as being post-mineral, however recent interpretations suggest that this could be a structure similar to that at Escondida, where recent drilling on the DTS has intersected anomalous gold values.

7.3.10

Michelle

The Michelle prospect consists of several parallel sub-vertical quartz veins, trending approximately north-south to north-northeast. The veins have been tested by drilling and trenching over approximately 400 metres in length and were found to be up to 2 metres wide. The low-grade veins contain crystalline quartz with poorly developed coarse banding and occur in rhyolite.

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7.3.11

Carla

The Carla prospect contains a quartz vein stockwork and breccia vein which is sub parallel to a major north-west fault. This major fault has been interpreted as the same regional structure that controls the mineralization at the Esperanza and Nini prospects, however mineralised clasts within a tectonic breccia are evidence that it has had been reactivated after the mineralisation. Patchy high grade mineralisation has been followed by drilling and trenching for over 150 metres and is truncated by the post mineral fault within 50 vertical metres of the surface. The thickness of the mineralisation is variable between 1 to 10 metres. High-grade mineralisation is associated with hydrothermal black silica breccias rich in sulphides; including pyrite, galena, sphalerite (and acanthite?). The mineralisation is emplaced close to a wide tectonic breccia zone that separates a felsic ignimbrite to the north and a coherent andesite unit to the south.

7.3.12

Natalia

The Natalia prospect is a major northwest trending structure, located south of the Gabriela prospect and north of the Esperanza structure. The vein is characterised by low grade white massive quartz with disseminated fine pyrite with an interpreted late brecciation stage filled with a quartz-hematite matrix. The outcrop has been mapped for 200 metres, but the ground magnetics and recent scout drilling suggest that the structure could be traced for 3 kilometres, particularly to the northeast. The vein’s width varies between 1 to 2 metres, but stock-work zones up to 4 metres have been observed. The vein is emplaced in an interpreted fault contact between the rhyolite and rhyodacitic ignimbrite.

7.3.13

Mosquito

The Mosquito prospect is characterized by the presence of possible sinter caps and eruption breccias, which have been interpreted as the upper portion of a low sulphidation system. Loma Mosquito is located in the northern part of the property near the Virginia prospect and is the only place where sinters have been recognized. Drilling in this area has confirmed the presence of anomalous gold associated with shallow, east dipping eruption breccias and silicified tuff.

7.3.14

Susy

The Susy prospect is a northeast trending structure, clearly visible in the ground magnetic data, which may displace the north-western portion of the Gabriela structure. The vein is typically narrow (up to 1.5 metres wide) and discontinuous with chalcedonic and minor crystalline quartz. The veining is hosted within dacitic, fiamme-bearing welded tuff and rhyolitic breccias. Initially traced by trenching and mapping by Mincorp for approximately 150 metres, the drilling and ground magnetic data infer a strike length of approximately 700 metres.

7.3.15

Cassius

The Cassius prospect is a clay filled, approximately 2 kilometre long, northwest structure. It is approximately 800 metres southwest of the Escondida prospect and trends sub-parallel to the Escondida structure. The Cassius structure juxtaposes the welded ignimbritic sequence to the north against the andesitic sequence to the south.

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7.3.16

Carolene

The Carolene structure incorporates the Loma Stockwork prospect which was referred to in the previous PEA and a new structure which is 250m north of and parallel to the Gabriela resource. Recent drilling has confirmed narrow low-grade mineralisation in pyritic quartz stock-work veining and a north-west trending vein up to 6 to 9 metres in width, on the contact of la Matilde Formation to the north against Chon Aike volcanics to the south. This is interpreted to be the northern edge of the Chon Aike horst block and like Escondida forms the bounding fault to the structural block. This vein, located beneath Tertiary cover, and is interpreted from the ground magnetic data and drilling to strike for a minimum of 1000m metres to the north-west.

7.3.17

Florencia

The Florencia prospect is located 250 metres toward the northeast of the Moro prospect and is characterised by sub-outcrop of veins and breccia-vein float, oriented approximately east-west. The Florencia structure has been traced for 300 metres and is interpreted to potentially connect with the Moro structure at its western end. The vein’s width varies between 1 to 4 metres. The presence of jarosite and anomalous arsenic values suggests that it has formed in an elevated crustal setting.

7.3.18

Virginia

The Virginia prospect has been interpreted as an eroded acid sulphate cap, commonly found in the preserved upper portion of low sulphidation system. Two zones, each 200 metres in length, have been tested by shallow drilling with some significant results. The trend of this silica-cap has a north-east strike and the disseminated mineralization is associated with pervasive silica-kaolin-alunite? alteration in a felsic tuff.

7.3.19

Laura

The Laura prospect is a north-south oriented white quartz vein located at the northern part of Cerro Moro. The structure dips 60 degrees to the east, and is situated at the lithological contact between rhyolite (hanging-wall) and rhyodacitic ignimbrite (foot-wall). The width of the vein varies between 3 to 6 metres, and is composed of massive white-quartz with scarce disseminated sulphides (mostly pyrite). The boundaries of the vein are characterized by a less than one metre zone of quartz-gossanous material.

7.3.20

Dora

The Dora prospect is a wide quartz filled structure with a north-northwest strike, dipping steeply to the east. Low-grade quartz has been traced for 240 metres and widths vary between 1 to 11 metres. The mineralisation is low grade and the white massive quartz vein is low in sulphides. The structure occurs in a felsic tuff with strong argillic alteration.

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7.3.21

Tres Lomas

The Tres Lomas prospect is named after three small hills aligned in a north-west orientation. Quartz veining on the hills can be traced for approximately 200 metres along strike and are hosted in rhyodacitic ignimbrite. The veins are mostly near vertical and have well developed stock-works with widths ranging between 0.3 to 5 metres. The gold-silver mineralisation is associated with chalcedonic quartz with some base metal rich zones associated with higher grade gold and silver, as well as disseminated pyrite.

7.3.22

Maria

The Maria zone comprises a number of irregular ptygmatic veins up to 1 metre in width hosted within a spherulitic rhyolite dome. The veins comprise massive milky quartz with large dog tooth crystals in cavities. The veins are anomalous in base metals. Host rock alteration is not apparent. The interpretation is that the veins were emplaced whilst the dome was hot, possibly semi molten. The dome has a high silica content with quartz crystals observed in lithophysae.

7.3.23

Carlita

The Carlita prospect is an east-west oriented, narrow, low-grade quartz vein structure, traced for 350 metres. The vein is composed of white quartz with scarce disseminated sulphides (mostly pyrite). The widths vary between 0.5 to 2 metres. The western end of the structure has split into two separate veins and is interpreted to be a splay of the major northwest trending Esperanza-Nini structure. The host-rock is rhyodacitic ignimbrite. To the south of the main vein there are outcrops of east-west sub-parallel veinlets close to the contact between the rhyolite and rhyodacite.

7.3.24

Lala

The Lala prospect is a north-northwest trending structure dipping to the north-east. Quartz vein outcrop can be observed for approximately 400 metres on the surface, and the structure is interpreted from the ground magnetic to continue for an additional 1.8 kilometres to the southeast under Tertiary marine sediments. The vein is emplaced at the contact between rhyodacitic ignimbrite with fiammes and rhyolite. Two pulses of mineralisation are interpreted; an early phase of white quartz-adularia, which is brecciated by a second later pulse of black silica material with fine disseminated pyrite. This black silica is in part chalcedonic. The vein’s width varies between 0.5 to 2 metres.

7.3.25

FMD

The FMD prospect is a strong east-west break in the ground magnetic data with an associated weak resistivity anomaly, that trend for at least 50 metres. Abundant quartz float in scree is observed at surface, hosted within rhyolitic units.

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7.3.26

Romina

The Romina prospect is an east-west orientated vein, located to the east of the Gabriela prospect and north of the Barbara and Lourde structures. The vein can be traced for approximately 500 metres as quartz vein sub-outcrop, but the ground magnetic data suggests the structure could be extended an additional 300 metres to the west and east. The vein dips shallowly to the north and is dominated by white milky quartz style veining, with a variable width between 1 to 3 metres. Minor fine disseminated sulphides (predominantly pyrite) have been observed. A post mineral breccia unit, with hematitic cement, cuts the vein in places.

7.3.27

Ornella

The Ornella prospect is represented by an east-northeast trending structure, located approximately 300 metres north of Escondida Central. The poorly outcropping vein has a strike of approximately 400 metres, and a variable width between 1 to 5 metres. The structure is clearly observed in the ground magnetic data, and dips steeply to the north-west, hosted within a coherent andesitic unit.

7.3.28

Deborah Termination Structure

The Deborah Termination Structure (“DTS”) prospect is a significant northwest trending structure that is interpreted to truncate the north-eastern end of the Deborah vein. The DTS structure can be clearly observed in the ground magnetic data and can be traced for at least 4.5 kilometres. In the Deborah prospect area the structure is completely covered by recent sediments, however, to the south-east recent regional mapping has discovered outcropping quartz veining and breccias that are parallel to the structure observed in the magnetics that could be interpreted to belong to the DTS structure. In this south-eastern sector, field mapping has identified hydrothermal breccias with a hematite matrix and fluidized textures, with widths varying between 1 to 4 metres.

7.3.29

Lourdes

The Lourdes prospect is a north-south trending quartz veining structure, with widths of over 5 metres, dipping steeply to the east. It is located to the south of the Romina structure and has an interpreted strike length of approximately 500 metres. The veining is dominated by white milky quartz with minor fine disseminated sulphides (predominantly pyrite) and is emplaced in a rhyodacitic ignimbritic unit.

7.3.30

Lucia

Lucia is a northwest orientated vein located approximately 10 kilometres from the Escondida prospect in the north-eastern corner of the Cerro Moro property. It was discovered in 2010 during an exploration program that included geological/structural mapping, surface sampling and ground geophysical surveys. Outcrops of the central portion of the structure have been mapped for over 200 metres however the geophysical and recent drilling data indicates a potential strike length of at least 2.5 kilometres. The vein dips steeply to the northeast and is hosted within a contact between rhyodacitc ignimbrites (hangingwall) and dome related felsic breccias (footwall). Multiple hydrothermal pulses

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have been observed in the vein zone, with very localised high grade gold and silver mineralization associated with the later events; consisting of coarse toothy quartz, white adularia and black silica material, rich in acantite. Although the thickness of the main mineralized structure is narrow (less than 2 metres), stockwork or zones of minor veinlets are commonly observed as a low grade halo to the main vein. Silicification and strong smectite alteration is related to the mineralization.

7.3.31

Gabriela South-East Extension

The Gabriela Southeast (“SE”) Extension prospect corresponds to a new shoot discovered on an extension of the main Gabriela structure. The vein does not outcrop and was discovered initially by a series of shallow exploration holes in an attempt to follow the Gabriela structure to the southeast. The Gabriela SE Extension is approximately 400 metres in length and has been drilled to a depth of approximately 200 metres from surface. The mineralization is associated with quartz-adularia, with typical lattice blade textures, hosted in a rhyodacitic ignimbrite (P1 unit). The thickness varies between 2 to 5 metres with some wider zones being observed related with lower grade stockwork zones. Fine disseminated base metal sulphides, acantite and electrum are observed in the highest grade gold and silver zones of the shoot.

7.3.32

Conceptual Targets

The Conceptual Targets represent an area of approximately 2 square kilometres that is completely covered by a thick sequence of Tertiary Marine sediments, located in the northwest portion of the Project between two major structures, Gabriela and Nini-Esperanza. A total of seven, predominantly north-west orientated structures, were identified in the ground magnetic data and targeted by a series of scout drilling fences, resulting in numerous anomalous auriferous intersections. Most of the structures tested dip steeply to the north-east; with veining comprised of chalcedonic-quartz veining with fine disseminate pyrite, and widths varying between 1 to 3 metres. The host rock is dominated by rhyodacitic ignimbrites.

7.3.33

Agostina

This is a recently discovered vein stockwork zone which has had some scout drilling. The zone has an east west strike and dips steeply to the north, and occurs at the contact of a siliceous rhyolite breccia to the south with felsic tuffs to the north. Localised high grade gold and silver mineralisation has been intersected, however follow-up drilling suggests it occurs in small discontinuous sections.

7.3.34

Belen

This west dipping vein was discovered and channel sampled by Mincorp prior to Extorre’s involvement with the project in December 2003. The prospect consists of at least 2 white quartz veins to 2m wide with a north-north-east orientation and a strike of 450m. Although previous sampling returned low grade precious metal results at surface it was decided that the grade may improve with depth and two holes were recently drilled to test this possibility. The deepest hole returned a maximum result of 4.93 g/t Au and 54 g/t Ag with anomalous base metals in a quartz vein over a 0.65m interval on the contact

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between the P4 sandstone and a felsic breccia. This has shown that precious metal values are increasing at depth and the relationship between the footwall and hangingwall stratigraphy is of interest as it is similar to other Escondida prospects albeit in a different strike direction.

7.3.35

Deborah Parallel

The northeast trending Deborah Parallel structure is located between the Deborah and Belen veins, and can be mapped over a strike length of 2.5km between the northwest trending Deborah Termination Structure and the east-west Zoe prospect on the Escondida structure. The Deborah Parallel structure dips steeply to the east and contains sections filled with quartz veins up to 9m wide. The Deborah Parallel structure is also considered to be significant as it is characterised by a downthrown block of P4 sandstone and andesitic lava to the east juxtaposed against a footwall of P1 Chon Aike material to the west, with a vertical displacement interpreted to be of the order of 200 metres. Like Belen this same juxtaposition occurs across the Escondida structure and is considered to be a significant fault possibly bounding the eastern side of the Chon Aike horst block. To date 6 scout holes have been drilled into the Deborah Parallel structure and results of that work are yet to be assessed.

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8.	DEPOSIT TYPES

The polymetallic gold-silver mineralisation at Cerro Moro is of the low to intermediate sulphidation epithermal quartz-adularia-sulphide vein type. Individual prospects vary from simple, single veins to complex vein systems with spur and cymoid loop structures. Limited quartz stock-work veinlets are also present around the main veins. Vein strike orientations vary from:

1.	Northwest to west-northwest with easterly flexures (e.g.: Escondida, Zoe, Esperanza-Nini, Gabriela, Natalia, and Dora), and

2.	Northeast (e.g.: Barbara, Deborah, and Maria-Michele);

And the less common:

3.	East-west to west-southwest (e.g.: Sylvia area, Loma Escondida, and Patricia), and

4.	North-northwest (e.g.: Lourdes, and Lala).

Veins are typically steeply dipping to sub-vertical. Outcropping veins locally reach widths up to 4 metres whilst associated zones of quartz stringers and stockworking may attain widths in the order of 10 to 15 metres. The strike length of individual veins is variable generally between 200 metres and 1 kilometre. The Escondida structure has been traced by drilling for approximately 8 kilometres, and remains open along strike and at depth.

Argillic and locally silicic alteration is present peripheral to principal veins and within dense minor vein swarms; however the alteration halos are not extensive and may extend only as much as 5 times the vein width into host rock.

The discontinuous ore shoots within the larger structures appear to be locally controlled by changes in strike, which are thought to have produced dilational flexures and jogs allowing for greater fluid flow. Changes in wall-rock lithology, along the structures, may be an important factor in controlling the local strike direction of structures. It has also been noted that brittle, felsic units have a tendency to produce stronger stock-work style development around main structures than intermediate units.

“Precise timing of the onset of mineralization is unclear, but the possibly long-lived vein mineralizing event certainly continued until quite late in the overall construction of the Jurassic volcanic pile since veins are observed to cut welded ignimbrites which comprise the youngest mapped stratigraphic group. Field evidence clearly indicates that vein mineralization overlapped in time with intrusion of rhyolite flow-domes and dykes. Intrinsically pyritic and strongly silica oversaturated rhyolites, characterized by internal development of irregular quartz veinlet stockworks and patchy hydrothermal breccias invite speculation as to a closer genetic link between vein mineralization and some of these rhyolitic intrusive phases.” (Callan, 2008).

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9.	EXPLORATION

9.1	Introduction

A comprehensive description of exploration activities prior to 2009 is detailed in Williams & Perkins (Exeter, 2009a). This document has been lodged with the Canadian Securities Administrators and is available for viewing on SEDAR at http://www.sedar.com.

Since commencing work in June 2003 Exeter has carried out diverse exploration activities at Cerro Moro that includes; geological mapping, various geophysical surveys, surface sampling and drilling. Table 13 chronologically summarises all exploration work undertaken on the Cerro Moro project prior to 2009.

Date	Exeter Exploration Works
Jun 2003	Revision of the CVSA Information
Feb 2004	Aster Image Interpretation
Apr 2004	Planning Reverse Circulation Drilling
Jun 2004	Reverse Circulation Drilling
Jun 2004	Induced Polarization Survey by Quantec
Aug - Dec 2004	Magnetic Susceptibility Survey of Reverse Circulation Holes
Sep 2004	Resistivity Survey by Akubra
Oct 2004 - Feb 2005	Rock Chip Sampling on Geophysical Anomalies
Aug 2004 - Apr 2005	Rock Chip Sampling
Apr - Oct 2005	Ground Magnetic Survey
Mar - Aug 2006	Ground Magnetic Survey - E-W Grid
Jun 2006	Reverse Circulation Drilling
Jul 2006	CVSA half yearly Technical Report
Oct 2006	Geology Mapping and Stratigraphy Interpretation on Carla Prospect
Nov 2006	Trench Sampling
Nov - Dec 2006	Diamond Drilling
Jan - May 2007	Ground Magnetic Survey - N-S Grid
Jan - Mar 2007	Trench Sampling
Jan - Feb 2007	LAG Sampling
Mar 2007	CVSA half yearly Technical Report
Mar 2007 – Dec 2008	Ongoing Diamond and Reverse Circulation Drilling
Apr - Jun 2007	Induced Polarization Survey by Quantec
May 2007	LAG Sampling
May - Jun 2007	Project Scale Mapping of Central - West area
Jul 2007	CVSA Half yearly Technical Report
Sep 2007	Joint venture report Notification of Completion of 10,000 metres of Drilling on the Co Moro Property
Oct 2007 –Jan 2008	LAG Sampling
Nov 2007	Comments on Au-Ag Mineralization Control and Drilling Target Definition
Jan – Mar 2008	Core Sampling for metallurgical test work

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Date	Exeter Exploration Works

Jan 2008	INTA* site visit to evaluate areas for rehabilitation/re-vegetation of disturbed areas
Feb – Jun 2008	Project scale mapping of the entire property at 1:10,000
Jul – Dec 2008	LAG Sampling
Mar 2008	PQ size diamond core drilling

* INTA = “Instituto Nacional de Tecnología Agropecuaria” of Argentina (National Institute of Agriculture Technology).

Table 13 Summary of Exploration Work undertaken by Exeter Prior to 2009

Work at Cerro Moro during 2009 and Quarter 1, 2010 concentrated on drilling the Escondida Prospect. Other regional exploration activities during this period included target generation through geophysical surveys and satellite imagery, minor RC scout drilling, mapping and sampling. Work during the remainder of 2010 was focussed on:

Resource Extension Drilling and Infill Drilling at the Known Prospects, where this work included:

	●	Peripheral drilling of the areas of Indicated resources at Escondida Far West, Escondida West, and Gabriela.
	●	Infill drilling at the Loma Escondida, Gabriela and Esperanza prospects, with the aim to convert Inferred resource category mineralization to higher confidence categories.
	●	Continual drilling along the Escondida structure away from the known mineralisation, both to the northwest and southeast. This drill program was successful in identifying new high grade gold-silver mineralisation to the southeast at Martina.
	●	Continual drilling to the southeast of the known Gabriela mineralisation, which has been successful in identifying a new zone of gold-silver mineralisation.

Drilling for the Preliminary Economic Assessment (PEA) and subsequent Pre-Feasibility Drilling, where this work included:

	●	The drilling of 22 RC percussion drill holes for 1,710 m as part of the hydrology and hydrogeology investigations.
	●	The drilling of 20 RC percussion drill holes for 2,018 m as part of a program to sterilize proposed infrastructure areas for the plant site, offices, accommodation, exploration camp, etc.
	●	The drilling of 6 diamond drill holes in the area of the proposed exploration decline at Escondida Far West, primarily to provide geotechnical data to assist with the design.

New Discovery Drilling, that included;

●

Drilling at the Carolene and Lucia prospects.

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Table 14 summarises in chronological order all exploration work undertaken on the Cerro Moro project during 2009 and 2010. Comprehensive descriptions of exploration activities undertaken by Extorre Gold Mines (after the split from Exeter Resource Corp in March, 2010) at Cerro Moro are detailed in three NI-43-101 technical documents dated March 18, 2010, June 2, 2010 and December, 3 2010; these documents are available for viewing on SEDAR at http://www.sedar.com.

Date	Exeter & Extorre Exploration Works
Jan 2009	Commencement of the first NI 43-101 compliant resource estimate
Apr 2009 - Jul 2009	Diamond Drilling at Escondida
Jun - Jul 2009	Magneto-Telluric and Transient Electromagnetic
Jul 2009	Public release of the first NI 43-101 compliant resource estimate
Aug 2009 - Feb 2010	Extension and Infill Diamond Drilling at Escondida
Jan 2010	Commencement of the second NI 43-101 compliant resource estimate
Apr 2010	Public release of the second NI 43-101 compliant resource estimate
May – Oct 2010	Exploration diamond drilling with the aim to extend high grade mineralization on known structures (Escondida, Gabriela and Loma Escondida) and test new targets (Caroline and Lucia).
May – Sep 2010	Ground magnetic and IP Gradient surveys at the Lucia and Martina prospects.
May – Jun 2010	Condemnation drilling program at the "Great Dyke" area, located 2 km south of Escondida. This area has been proposed for infrastructure (plant, offices, accommodation) for the future mine
Jul – Aug 2010	Drilling at the "Tres Lagunas" area located to the east of the property to test for ground water.
Oct – Dec 2010	Diamond In-fill drilling program at the Gabriela and Loma Escondida prospects with the aim to convert Inferred resource category mineralization to higher confidence categories. Completion of the maiden Preliminary Economic Assessment for the proposed Cerro Moro Mine Development.

Table 14 Summary of Exploration Work undertaken by Exeter Since 2009

To May 31, 2011, Extorre’s exploration activities were focused on increasing Cerro Moro’s global mineral resources through the conversion of new discoveries at Carla, Martina, Loma Escondida and Gabriela SE to Inferred Resource Category. In addition, a significant new high grade gold-silver discovery was made at the Zoe prospect, which is located approximately 2.5 km southeast of the Martina zone on an east-west portion of the Escondida structure. Exploration drilling also continued to test extensions to known mineralisation at Lucia, Gabriela, Loma Escondida, and Esperanza. Scout drilling has also been aimed at finding new deposits at Agostina, Tres Lomas, Belen, Deborah Parallel, and Carolene. Regional drilling was also undertaken at the Union Domes Prospect, located some 14.5 km south of Escondida on the Formicruz joint venture tenements.

Table 1 and 15 chronologically summarise all exploration work undertaken on the Cerro Moro project up until 31st May 2011.

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Date	Extorre Exploration Works
Jan 2011	Complete infill drilling Martina, Loma Escondida, Gabriela with the aim to convert Inferred resource category mineralization to higher confidence categories. Exploration drilling to extend Lucia, Gabriela Loma Escondida and Esperanza. Scout drilling at Carolene
Feb 2011 to end March 2011	Exploration drilling to extend Gabriela, Esperanza- Nini, Escondida far west, Michelle, Loma Escondida. Scout drilling Agostina Gabriela NW, Belen and Zoe (discovery of Zoe) Regional drilling Union domes
April 2011 to end May 2011	Infill drilling at Carla and Gabriela SE. Exploration drilling at Zoe, Tres Lomas and Gabriela. Scout drilling Agostina, Deborah parallel, Loma Escondida and Zoe extensions

Table 15 Summary of Exploration Work undertaken by Extorre to May 31, 2011

9.2	Geophysics

Geophysical surveys completed at Cerro Moro during 2010 (refer to Coupland, 2010 for geophysical work conducted during 2009) consisted of additional ground magnetic data acquisition as well as gradient array induced polarization (“IP”)/resistivity surveys. No new geophysical programmes have been completed since September 2010.

9.2.1

Ground Magnetic Surveys

Work completed in 2010 included a 4 km x 5.5 km block located in the north-eastern corner of the property covering the Lucia Prospect. This survey was undertaken utilizing north-south lines spaced 80 metres apart. Refer to Figure 9.1 for current ground magnetic coverage of the tenements and an example of one of the data products, reduced to the pole (“RTP”).

Figure 13 Cerro Moro Ground Magnetics – Reduced to Pole (RTP) Image

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9.2.2

Gradient Array IP Surveys

Approximately 91 line kilometres of Gradient Array IP / Resistivity data were acquired in early- to mid-2010 along 70 survey lines in two areas. The first survey area, consisting of 59 line kilometres and 29 survey lines, was conducted to fill a gap in the previous data coverage over the Escondida trend and the Bella Vista area to the southeast. The Gradient Array IP / Resistivity survey data was reported as the Newmont standard for integration with previously acquired data over the project area. Data quality for this survey is considered to be good and the results of the survey provide a “coherent and reasonably accurate representation of the geo-electrical properties (apparent resistivity and chargeability) of the subsurface suitable for geological interpretation within normal limitations” (Scarbrough, 2010a). Results obtained from this survey are summarized in Figure 14 and Figure 15.

Figure 14 Escondida to Bellavista – Chargeability Data

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Figure 15 Escondida to Bellavista – Resistivity Data

The second survey area, conducted over the Lucia prospect, consisted of 32 line kilometres along 41 survey lines. The “Gradient Array IP / Resistivity survey data was processed with in-house software and chargeability data was reported as the Newmont standard. Data quality for this survey is considered to be good and the results of the survey provide “a coherent and reasonably accurate representation of the geo-electrical properties (apparent resistivity and chargeability) of the subsurface suitable for geological interpretation within normal limitations” (Scarbrough, 2010b). ). Refer to Figure 16 and Figure 17 for a summary of the results.

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Figure 16 Lucia Prospect – Resistivity Data

Figure 17 Lucia Prospect – Chargeability Data

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10.

DRILLING

10.1

Summary

Prior to Exeter being involved in the project, a program of 34 drill holes for 2,593 metres was undertaken by Mincorp, comprising 19 diamond drill holes for 1,016 metres and 15 RC percussion drill holes for 1,577 metres.

As at the end of 2008, Exeter had drilled a total of 472 drill holes for 50,257.15 metres, comprising 264 diamond holes for 20,876.85 metres and 208 RC percussion drill holes for 29,380.30 metres. During 2009 and Quarter 1, 2010 the drilling had focussed primarily on the Escondida prospect, with 327 diamond drill holes being drilled for 39,722.7 metres to the 12th February 2010 (data cut off date for the April 2010 resource estimate). In addition, an RC percussion drill rig arrived on site during August 2009 and by the end of December 2009 had drilled 44 holes for a total of 4,729.3 metres at Cerro Moro. Of these, 35 holes were sited at 8 new exploratory targets with the remaining 9 holes drilled for water.

As at May 31, 2011, a total of 1,255 drill holes for 161,824.5m had been completed on both exploration and infill drilling at Cerro Moro (Table 16). An additional 22 holes for 1,801.3m had been drilled for water exploration, and a further 20 holes for 2018.0m were completed for sterilization of the proposed processing plant and camp site. Approximately 76% of the holes have been drilled at Escondida (50.38%), Gabriela (14.58%), Zoe (6.08%), and Esperanza (5.38%), reflecting the closer spaced drilling at the Escondida and Gabriela ore zones.

Prior to 2009 the majority of drilling at Cerro Moro had been undertaken by Major Perforaciones Argentina; a subsidiary of the Major Drilling group based in Canada, with a limited number of diamond holes drilled by Patagonia Drill Mining Services S.A. Experienced expatriate drilling supervisors supervised the drilling operations. All drilling during 2009 and Quarter 1, 2010 was undertaken by Connors Argentina, S.A., a wholly owned subsidiary of Boart Longyear, with corporate headquarters in Utah, U.S.A. In mid-2010, two Major drill rigs arrived on site to complement the two Boart-Longyear diamond drill rigs, and as of May 31, 2011, all 4 drill rigs continued to operate on site.

All diamond core drilling has been HQ3 size utilising triple tube equipment. The majority of diamond core holes were orientated using the Ballmark orientation system to provide accurate core orientations, although this system was replaced in September, 2010, by the Reflex ACT2 electronic system. Where water was intersected in RC percussion holes, resulting in lower sample recoveries, the holes were completed using diamond core to test the target zones.

RC percussion holes and percussion pre-collars to diamond drill holes were drilled with face-sampling hammers with hole diametres between 13.0 and 14.0 centimetres.

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Drilling Program	Prospect	Holes	RC (m)	DDH (m)	Total (m)	%
MINERALIZATION	Escondida inc Martina	575	12,384.50	69,147.70	81,532.20	50.38
	Gabriela	153	2,363.05	21,223.10	23,586.15	14.58
	Zoe	39	0.00	9,841.80	9,841.80	6.08
	Esperanza	74	2,155.80	6,552.80	8,708.60	5.38
	Loma Escondida	66	1,090.00	3,164.55	4,254.55	2.63
	Lucia	20	0.00	2,877.45	2,877.45	1.78
	Caroline	18	710.00	1,961.30	2,671.30	1.65
	Carla	33	426.00	2,099.30	2,525.30	1.56
	Patricia	21	640.00	1,503.25	2,143.25	1.32
	Nini	25	1,011.50	1,082.90	2,094.40	1.29
	Conceptual Target	23	1,899.45	94.80	1,994.25	1.23
	Agostina	17	0.00	1,715.50	1,715.50	1.06
	Natalia	15	624.00	1,078.65	1,702.65	1.05
	Michelle	15	471.00	1,200.80	1,671.80	1.03
	Silvia	14	921.00	724.15	1,645.15	1.02
	Moro	15	1,124.00	358.95	1,482.95	0.92
	Deborah	24	965.00	501.10	1,466.10	0.91
	Tres Lomas	12	276.00	636.00	912.00	0.56
	Cassius	6	474.00	335.30	809.30	0.50
	Deborah Parallel	4	269.00	447.50	716.50	0.44
	Dora	10	342.00	360.50	702.50	0.43
	Mosquito	9	470.00	230.65	700.65	0.43
	Susy	7	526.00	93.00	619.00	0.38
	Escondida North	4	310.00	191.15	501.15	0.31
	Nini Esperanza gap	4	440.00	50.50	490.50	0.30
	Florencia	4	304.00	168.00	472.00	0.29
	FMD	4	447.00	0.00	447.00	0.28
	TEM anomaly	3	420.00	0.00	420.00	0.26
	Virginia	11	411.00	0.00	411.00	0.25
	Laura	4	330.00	80.00	410.00	0.25
	MT	3	330.00	0.00	330.00	0.20
	DTS	3	321.00	0.00	321.00	0.20
	Maria	2	284.00	0.00	284.00	0.18
	Carlita	4	265.00	0.00	265.00	0.16
	Lala	3	264.00	0.00	264.00	0.16
	Marina	4	264.00	0.00	264.00	0.16
	Belen	2	0.00	222.50	222.50	0.14
	Romina	2	142.00	0.00	142.00	0.09
	Ornella	2	127.00	0.00	127.00	0.08
	Lourdes	1	80.00	0.00	80.00	0.05
Subtotal		1255	33,881.30	127,943.20	161,824.50	100.00

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GROUND WATER	Escondida	8	625.30	100.00	725.30	40.07
	Tres Lagunas	7	578.00	0.00	578.00	31.93
	TEM anomaly	2	133.00	0.00	133.00	7.35
	Lechuzo	1	90.00	0.00	90.00	4.97
	Esperanza	1	82.00	0.00	82.00	4.53
	Great Dike	1	80.00	0.00	80.00	4.42
	Carla	1	61.00	0.00	61.00	3.37
	Gabriela	1	61.00	0.00	61.00	3.37
Subtotal		22	1,710.30	100.00	1,810.30	100.00
STERILIZATION	Great Dike	20	2,018.00	0.00	2,018.00
TOTAL		1297	37,609.60	128,043.20	165,652.80

Table 16 Drilling Details for Cerro Moroat May 31, 2011

All Mincorp and Exeter RC percussion and diamond drill collars up until MD0268 (31-1-2008) were surveyed by a surveying contractor using a total station EDM theodolite and a differential GPS. The remaining drill collars to 31st May 2011 have been positioned by a surveying contractor using a differential GPS. Down-hole surveys were performed at the time of drilling initially utilising an Ausmine (“Eastman-type”) down-hole camera, and since March 2007, using a digital Reflex down-hole camera.

Figure 18 Drill Collars and Prospects Location Map

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10.2

Escondida

As of May 31, 2011 a total of 575 drill holes for 81,532.20 metres (comprising 69,147.70m of diamond drilling and 12,384.50 m of RC drilling) had been completed along the Escondida structure. This figure includes RC percussion metres drilled as pre-collars to diamond drill holes. Drilling has traced the Escondida structure for approximately 8 kilometres, with mineralisation having been identified in the following sectors (from west to east); Escondida Far West, Escondida West-Central-East, Escondida Far East, Martina, and Zoe.

The focus of the extensive drilling activities at Cerro Moro during 2009 and Quarter 1, 2010 was infill and extensional drilling on the Escondida prospect. Infill drilling to 20m x 20m was undertaken to upgrade substantial portions of the Escondida prospect to the Indicated mineral resource category. Since Q1, 2010, drilling has been focused on both the Martina zone (where drilling has been completed on a 60 x 60 m grid with the objective of defining an Inferred Category mineral resource) and on the newly-discovered Zoe sector, which is located approximately 2.5 km east of Martina. Longitudinal projections, constructed in the plane of the main vein, have been prepared and colour coded based on estimated true widths multiplied by the gold equivalent values. The gold equivalent value is calculated by dividing the silver grade (ppm) by 50 (approximate historical ratio of gold/silver US$ value) and adding it to the gold grade (ppm).

Sectors within the Escondida structure with defined mineral resources are shown in Figure 19; selected schematic cross-sections within the Escondida prospect are also shown in Figure 20 to Figure 22.

10.2.1

Escondida Far West

The Escondida Far West discovery was the result of scout fence drilling designed to locate the potential north-western extension of the main Escondida structure, in an area with extensive post mineralisation cover. The detailed ground magnetic data was utilised to estimate the position of the structure. None of the mineralisation outcrops at surface. In addition to being a blind mineralised body, near surface holes, as well as holes on the margin of the higher grade shoot, intersected only low grade to anomalous gold and silver values. This has implications over the entire property, where many of the veins are low grade near surface, and were previously assigned a low priority. Drilling at Escondida Far West during 2009 and Quarter 1, 2010 had focused on infill drilling of the higher grade mineralised shoots to 20 x 20 metre spacing to support an Indicated mineral resource category. Since Q1 2010 minor drilling has been done to test deeper mineralisation and extensions to the ore shoots at approximately 40 x 40 and 80m x 80m metre spacing.

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10.2.2

Escondida West, Central and East

Following the discovery of the two higher grade shoots at the central and western portions of Escondida, drilling focussed on tracing these plunging zones. Initially Escondida Central and Escondida West were interpreted to be two separate veins, or represented a faulted offset of the structure. However, diamond drilling demonstrated that the two zones are in fact one continuous structural zone.

Drilling at Escondida West, Central and East during 2009 and Quarter 1, 2010 has focused on infill drilling of the higher grade mineralised shoots to 20 x 20 metre spacing to support an Indicated mineral resource category. In addition, some drilling to test deeper mineralisation and extensions has been conducted at approximately 40 x 40 metre spacing. Since Q1 2010 only minor drilling has been done to test deeper mineralisation and extensions to the ore shoots at approximately 40 x 40 and 80m x 80m metre spacing.

10.2.3

Escondida Far East

A significant event for Cerro Moro was the “bonanza” intercept in MD373 located at Escondida Far East in mid-2008. It was a “blind discovery” below shallow RC percussion holes that were drilled to test the south-eastern extension of the Escondida structure beneath gravel cover which was predicted by geophysical surveys. A number of additional holes have been drilled around MD373 all of which have returned lower tenor results. This area of Escondida Far East appears to be complex and additional drilling and interpretation is required prior to inclusion in any reportable resources.

10.2.4

Martina

The Martina zone was discovered by following the Escondida structure to the east of the Escondida Far East shoot until an easterly flexure was discovered below stratigraphic cover of the P5 unit (La Matilde felsic ignimbrite). The high grade gold-silver mineralisation at Martina is completely blind at surface with no geophysical or geochemical evidence to guide drilling. The mineralisation at Martina is also much deeper than the other known orebodies to the west, commencing at 200m below surface. Structurally, Martina is more complex than the other Escondida targets in that there are several areas where the veins split into the hangingwall and footwall. On the basis of drilling completed to date, Martina has a higher gold to silver ratio (and lower base metal content) than other mineralised sectors at Escondida.

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10.2.5

Zoe

The high grade Zoe gold-silver zone is located approximately 2.5 km east of Martina and was first intersected by drill hole MD1196 in March, 2011. Drill hole MD1196 was designed to test a small area of quartz rubble which returned 0.34 g/t Au and 115 g/t Ag from surface rock sampling. The overall geology at Zoe is identical to other Escondida ore zones, with the mineralised vein being hosted at the contact between a P4 sandstone (hangingwall) unit and P1 ignimbrites (footwall). However, the Escondida structure trends east-west in the Zoe sector rather than northwest-southeast. Drilling is currently being undertaken at Zoe on a relatively broad spacing (160 m x 160 m and 80m x 80m) in order to define the limits of the mineralisation. The Zoe shoot appears to have larger dimensions than the other Escondida shoots, with significant mineralisation having been intersected over a strike length of 600m and to a depth of 300m. The Zoe mineralisation appears to pinch out into a shear zone as it nears the surface, hence explaining the very poor surface expression of this shoot.

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Figure 19 Escondida Total Project Longitudinal Projection - April 2011 Drilling (Local Grid)

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Figure 20 Escondida Far West Shoot - Schematic Cross-Section 29502E (Local Grid)

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Figure 21 Escondida West Shoot - Schematic Cross-Section 30117E (Local Grid)

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Figure 22 Escondida Central Shoot - Schematic Cross-Section 30475E (Local Grid)

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10.3	Other Prospects

10.3.1

Gabriela

Recent drilling at the Gabriela prospect has focused on i) infill drilling of the Gabriela Central shoot on a 40 m x 40 m spacing, with the primary purpose being to upgrade the existing Inferred mineral resources to a higher confidence level category, ii) infill drilling at a 40 m x 40 m spacing on the Gabriela SE ore shoot, and iii) a program of 4 deep drill holes to test for down-depth extensions of the Gabriela Central ore shoot. Assay results from i) and ii) have been incorporated into the updated mineral resource statement for Gabriela Central and SE shoots as reported in Section 14.6.3, whereas assay results from the ongoing deep drilling program (iii) are pending.

10.3.2

Loma Escondida

During Q4, 2010, an infill drill program (40 m x 40 m spacing) was completed at the Loma Escondida prospect, with the primary purpose being to upgrade the existing Inferred mineral resources to a higher confidence level category. Assay results from the infill drill program have been incorporated into the updated mineral resource statement for Loma Escondida as discussed in Section 14.6.2. No further drilling has been completed at Loma Escondida since end-2010.

10.3.3

Esperanza

Recent exploration drilling at the Esperanza prospect has been limited to i) step-out and step-back drilling around the known gold-silver mineralization in the Esperanza Central shoot, and ii) scout drilling in the area of the “gap” between the Esperanza vein and the postulated northwestern extension at Nini. Further drilling is expected to be undertaken in these areas in 2011, with the primary aim being to increasing mineral resources in the Esperanza prospect.

10.3.4

Lucia

A total of 20 diamond drill holes (for 2,877 m) were completed at the Lucia vein zone during Q4, 2010. The Lucia drill program was designed to evaluate the strike and depth extent of surface outcrops of veins / breccias containing high grade silver-gold mineralization. In overview, significant silver-gold mineralization was intersected in drill holes located nearest to the mineralized outcrops, however, step-out drilling along strike failed to intersect any significant mineralization. A more detailed evaluation of the assays results (especially trace element geochemistry) will still need to be undertaken before a final decision is reached on the resource potential of this zone.

10.3.5

Carla

An infill drill program (20 m x 20 m spacing) was completed on the high grade central portion of the Carla vein zone in April 2011. A total of 700 m of drilling was completed in 7 holes. As of May 31, 2011, assay results were still pending.

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10.3.6

Natalia

Six scout drill holes (for 820m) were completed along both the northwestern and southeastern extensions to the Natalia vein, a northwest-trending vein occurrence located between Esperanza and Gabriela. The current drilling confirmed that the Natalia vein ranges in width from 2 m to 5 m and contains anomalous gold-silver values; further work is proposed.

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11.	SAMPLING METHOD AND APPROACH

11.1	Previous Sampling

No information is available on Mincorp´s sampling procedures and protocols.

11.1.1

Surface Sampling

The same methodology was used to collect the samples in the trenches and surface rock chip channel sampling. All the samples were taken by a field assistant under a geologist’s supervision utilising hammer and chisel. The samples to be collected are initially marked up by a geologist using a nominal 2 meter sample length with smaller lengths dependent on geological or mineralisation contacts. All of the sampling in the trenches and surface rock chip channels are continuously sampled. Representative chips of each sample interval are stored in plastic boxes for future reference. The samples are placed in marked plastic bags, sealed and transported to the assay laboratory.

11.1.2

Diamond Drilling

All diamond drilling at the Cerro Moro Project has been completed by internationally-recognized drill contractors such as Boart-Longyear and Major Drilling Services.

All diamond drill holes completed by Extorre are orientated using the Ballmark system. This system uses gravity to make a mark on an aluminium disk with a small ball, with the mark representing the bottom of the hole in the oriented core. An Extorre technician at the drill site measures core recovery and draws an initial orientation line on the core. The drill core is then placed in marked wooden core boxes at the drill site and transported to Extorre’s camp (currently located at the north-western corner of the property) for processing.

The orientation is verified at the camp before detailed geotechnical logging by trained specialist technicians. Geotechnical logging documents rock quality, which includes structural measurements of defects. All data is recorded on a meter by meter basis and includes; recovery, rock quality designation (“RQD”), strength, weathering, and the orientation of fractures. Utilising the orientation line (which represents the bottom of the hole) “alpha” and “beta” measurements are taken of the fractures. The alpha angle is the angle between the fracture plane and core axis. The beta angle is the angle between the axis of the fracture and orientation line measured in clockwise direction. Knowing the alpha and beta angles of the fractures and the orientation of the hole (dip and azimuth), the true spatial position of the fractures can be calculated (dip and dip direction) using Dips software.

Core recoveries are estimated both on a “per drill run” basis using the core marker blocks and also on a “per meter” basis. Good recoveries (>95%) have generally been obtained from diamond drilling at Cerro Moro.

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The core derived from each diamond drill hole is digitally photographed, wet and dry, by technicians before cutting and sampling. The geology is subsequently recorded by geologists, with both rock types and alteration / mineralization data being recorded and entered into the data-base to allow the construction of geological sections. Structural measurements of faults, veins and geological contact are taken with a Brunton compass in an orientation frame utilising the orientation line as a reference.

Upon completion of logging and photography, a geologist marks the core for sampling. Sample lengths through the obvious mineralised zones vary between 0.3 and 1.5 metres, depending on geological and structural contacts, and these samples are classified as “high priority”. The remainder of each drill hole, classified as “low priority”, is generally sampled every one metre. To date, generally the complete hole has been sampled. Extorre technicians utilise a diamond saw on-site to cut the core in half; one half of the core is sampled and sent to the laboratory for analysis and the remaining half is stored on-site in the core boxes. The core saw is cleaned with a brick or other abrasive stone between each high priority sample to eliminate any possible contamination between potential high grade samples. The samples are placed in marked plastic bags, sealed, and transported to the assay laboratory.

11.1.3

RC Percussion Drilling

RC percussion samples are collected using a cyclone attached to the drill rig at one metre intervals. The geological logging is performed by Extorre geologists at the drill site. On the completion of logging, a geologist determines the sample interval lengths of the high and low priority zones. The high priority, potentially mineralized, zones are sampled at one metre intervals and composite samples of three metres are collected through the low priority zones. Each of the one metre samples are stored in plastic bags and are weighed and recorded by technicians at the drill site. The geologist records the diameter of the drilling tools (bit and shoes) at the beginning and end of each hole and uses these measurements to calculate an estimated weight of each sample. In this way the recovery of each sample can be calculated, assuming a density of 2.5. The resulting recoveries average above 85 percent. The sampling is performed at the Extorre camp by technicians, utilising a riffle splitter, where each metre sample is split 3 times. The average 1 metre sample weight is approximately 3 kilograms, with an average of 9 kilograms for the 3 metre composite samples.

11.2	Density Determination

Specific gravity (“SG”) values have been calculated for core samples at Cerro Moro routinely since August, 2007. The methodology used is to calculate the ratio of the weight of sample in air to the difference between the weight in air and weight in water. Core samples, generally >10 cm in length are oven dried to 100° C, sprayed with waterproof sealant and weighed. The sample is then weighed whilst suspended in water and the SG is calculated as follows:

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SG = Wp _

Wp – Wps

Where:

Wp = Weight of the sealed sample

Wps = Weight of the sealed suspended sample in water

The SG data is recorded for all lithologies encountered and will be used as the basis for assigning density values in future scoping studies. Vein samples, which include all types of quartz vein lithologies, vein breccias and stockworks, have SG values that vary between 2.45 and 2.81. Non vein samples, which include all other stratigraphic units, have SG values that vary between 2.34 and 2.61. The variations of the vein results are interpreted to mostly reflect the sulphide content associated with the veins.

As part of the quality control process, a set of check samples were submitted to Mecanica de Rocas in Santiago, Chile for SG determination. Ninety five percent (95%) of the 106 samples returned an error of <5%.

11.3	Sample Preparation and Analysis

The only sample preparation conducted by Extorre employees is the diamond saw cutting of diamond drill core samples and riffle splitting for the RC percussion samples (as detailed in sections 11.1.2 and 11.1.3 respectively), together with the bagging and sealing of samples on site. No other part of the sample preparation or analysis process, as detailed in the following sections, is conducted by an employee, officer, director, or associate of Extorre.

Prior to February 2011, samples from the Cerro Moro Project were routinely sent to ACME Analytical Laboratories (Argentina) S.A. (“ACME”) in Mendoza, Argentina, for preparation (drying, crushing, and milling). Sample assaying is subsequently undertaken at ACME’s laboratory in Santiago, Chile. Historically, some diamond drill holes from Cerro Moro were also despatched to ALS Chemex’s sample preparation facility in Mendoza, Argentina, and then analyzed at ALS Chemex’s laboratory in La Serena, Chile.

11.3.1

ACME On-site Sample Preparation Facility

In February 2011, Acme Labs established a dedicated on-site sample preparation laboratory for Extorre at the Cerro Moro Project. Samples are prepared by experienced personnel and a pulp split is sent to the company’s ISO9001 certified analytical laboratory located in Santiago Chile for analysis. The sample preparation facility has a capacity of 150-300 samples per day. Activities carried out by the on-site sample preparation facility are as follows:

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	●	Drying (60°C);
	●	Crushing (80% < 10#);
	●	Splitting;
	●	Pulverizing of the split fraction.

The facility on site has a DCU (Dust Control Unit) that keeps the environment free of dust and contamination during processing.

The sample preparation facility is equipped with the following:

	●	Two crushers (Terminator TM);
	●	One sample splitter (Hebro);
	●	One drying oven (Made in house);
	●	Two disc pulverizers (LM2 Labtechnics);
	●	One air compressor (Compresores Hernandez de 7.5 HP; Local product).

11.3.2

ACME Procedures

All sample batches are identified on the sample dispatch order form as being either of “high priority” (obvious mineralized zones) or “low priority” (not obviously mineralised).

A description of high and low priority sample preparation and analytical methods for ACME are detailed below.

Procedures for High Priority Samples

	●	Samples are received at the ACME on-site sample preparation facility at Cerro Moro, Argentina;
	●	Sample dispatch details are checked against samples received and confirmed with client;
	●	Samples numbers are entered into the laboratory information system and bar coded sample number stickers matching the client sample numbers are generated;
	●	Each sample is weighed at sample reception;
	●	Samples are oven dried at 60°C for a minimum of 24 hours;
	●	Samples are crushed to 85% passing -2mm using a TM Engineering Terminator Jaw crusher;
	●	A barren quartz wash is inserted at the beginning and end of each job number and at every tenth sample during the crushing phase;
	●	A 500 gram split is obtained using a stainless steel riffle splitter;
	●	Pulverising to 85% passing 200 mesh (-75 micron) is achieved using a Labtechnics LM2 disc pulveriser (Code: R200-CRU85);
	●	A barren quartz wash is inserted at the beginning and end of each job number and at every tenth sample during the pulverising phase;
	●	Pulps are weighed prior to dispatch to ACME Santiago, Chile;

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	●	Au (50 gram) by fire assay and AAS (Code: G6);
	●	When the grade of Au is higher than 10 parts per million (“ppm”), the samples are re-assayed by fire assay with a gravimetric finish (Code: G6 GRAV);
	●	Ag (1 gram/100 millilitre) by 4 acid digestion and AAS (Code: 8TD);
	●	When the grade of Ag is higher than 1,000 ppm, the samples are re-assayed by fire assay with a gravimetric finish (Code: G6 GRAV);
	●	35 elements by hot aqua regia digestion and ICP-ES (Code: G1D + Hg + Tl);
	●	Samples with >1 ppm Au are repeated by fire assay and AAS (Code: G6);
	●	Au via screen fire assay, if required (Code: G6 SFA).

Procedures for Low Priority Samples

	●	Crush, split and pulverise drill core to 200 mesh (Code: R200-CRU85);
	●	Au (50 gram) by fire assay and AAS (Code: G6);
	●	33 elements by hot aqua regia digestion and ICP-ES (Code: G1D).

11.3.3

ALS Chemex Procedures

The two batches of drilling samples dispatched to the ALS Chemex laboratory were flagged, on the sample dispatch order form, as either “high priority” (obvious mineralized zones) or “low priority” (not obviously mineralised).

A description of high and low priority sample preparation and analytical methods for ALS Chemex are detailed below in Sections 13.1.2.1 and 13.1.2.2.

Procedures for High Priority Samples

	●	All samples to be crushed to –2 mm (laboratory code CRU-31) and followed before and after with a barren flush (laboratory code WSH-21);
	●	Whole sample to be pulverised to –75 micron (laboratory code PUL-21);
	●	Assay Au by laboratory code Au-AA26 (50 gram Au fire assay; 0.01 to 100 parts per million (“ppm”) detection.);
	●	Ag by laboratory code Ag-GRA21 (30 gram Ag fire assay with gravimetric finish; 5 to 10,000 ppm detection);
	●	Multi-elements analysis by laboratory code ME-MS61m (four acid near total digestion; 48 elements and mercury);
	●	Samples with >1 ppm Au, repeat assay by laboratory code Au-AA26 (50 gram Au fire assay; 0.01 to 100 ppm detection).

Procedures for Low Priority Samples

	●	Crush whole sample laboratory code CRU-31 (-2 mm);
	●	Pulverise whole sample –75 micron laboratory code PUL-21;

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	●	Analyse Au by laboratory code Au-AA24 (50 gram fire assay with AAS finish 0.05 to 10 ppm detection. Retain all residue);
	●	Ag by laboratory code AgAA-61 (four acid digest with AAS finish; 0.5 to 100 ppm detection);
	●	Samples >1 Au re-assay by laboratory code laboratory code Au-AA24.

11.3.4

Representivity of Metallurgical Sampling

Extorre have undertaken extensive metallurgical sampling from the various mineralisation styles at the Cerro Moro project. As of December 2009, a total of 114 metallurgical samples had been collected of which 113 are from the main potentially ‘economic’ deposits at Cerro Moro (Escondida, Loma Escondida, Gabriela and Esperanza). Metallurgical samples have been assembled from representative diamond drill core and sent to Metcon Laboratories in Sydney, Australia for testing. The results of this metallurgical sampling program form the basis of this PEA. Additional metallurgical testwork was ongoing at the time of writing this report.

The metallurgical sampling program has been undertaken in such a way as to ensure that sufficient representative material has been selected from the various styles of mineralisation likely to be encountered during mining. The sampling program includes a significant amount of material associated with the ‘bonanza’ grade sulphide mineralisation found at Escondida and Loma Escondida together with lower grade peripheral sulphide mineralisation and dilutant material. Specific samples have also been selected that are representative of the various oxidation conditions at Cerro Moro. Figure 23 to Figure 26 show metallurgical sample locations for Escondia, Loma Escondida, Gabriela and Esperanza as of December 2009.

It is the Author’s opinion that the metallurgical sampling currently available for the Cerro Moro project is well advanced and is sufficiently representative of all styles of mineralisation likely to be mined at the project.

Figure 23 Escondida –Metallurgical Sampling Locations

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Figure 24 Loma Escondida – Metallurgical Sampling Locations

Figure 25 Gabriela – Metallurgical Sampling Locations

Figure 26 Esperanza – Metallurgical Sampling Locations

11.4	Quality Control

Quality control procedures include the regular use of client initiated geochemical standards, sample duplicates (RC percussion and lag soil samples only), geochemical blanks, check assaying of samples returning greater than 1.0 g/t gold, and twins of existing drill holes. These are fully discussed in Section 12.0 Data Verification. The results from the quality control sampling demonstrate that there were no obvious errors or bias in the sample preparation or analysis, and as such no corrective actions were required.

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11.5

Security

Sampling of rock chips, trenches and rock chip channel samples, RC percussion chips, and drill core has been conducted on-site by Company personnel under the supervision of experienced Company geologists.

Regional exploration samples for assay are placed in sealed plastic bags with a numbered sample tag firmly stapled inside the bag. Depending on the individual sample size, two to six samples are then placed in larger woven plastic bags, which are then sealed with cable ties and numbered in preparation for transport to the laboratory.

Up until June, 2008, the majority of Cerro Moro samples were transported from site by Company vehicle to the bus station in either Caleta Olivia or to Comodoro Rivadavia. The samples were then transported by bus to the ALS Chemex preparation facility in Mendoza, Argentina. Commencing in July, 2008, samples were transported by a private contractor (30 tonne capacity truck) from site directly to the ACME sample preparation facility in Mendoza, Argentina. Regional exploration sample continue to be sent for sample preparation to ACME's sample preparation facility in Mendoza. Upon completion of sample preparation, pulps from regional exploration are sent by Jet Paq to ACME's laboratory in Santiago, Chile.

Commencing in February, 2011, sample preparation of resource definition samples has occurred on site. Pulps from the on-site sample preparation facility are transported by van truck operated by a private contractor directly to Jet Paq (Aerolineas Argentina) in Comodoro Rivadavia. The pulps are transported by Jet Paq to Mendoza where they are stored at ACME's sample preparation facility for a short period before being sent by Jet Paq to ACME's laboratory in Santiago, Chile.

11.6	Authors Statement

The author has witnessed all aspects of sample preparation and dispatch carried out by Extorre staff. In addition, the author has visited the ACME sample preparation facility in Mendoza, Argentina and found it to be a well organised, clean and high quality facility. The author has not yet had the opportunity to visit the on-site sample preparation facility at Cerro Moro. The author has not visited the ACME analytical facility in Santiago, Chile, however the author has no reason to doubt the integrity of this facility. In the author’s opinion, the sample preparation, security and analytical procedures employed by Extorre are consistent with standard industry practice particularly in the exploration for precious metals.

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12.	DATA VERIFICATION

The various steps taken by Extorre to ensure the integrity of analytical data are consistent with standard industry practice particularly in the exploration for precious metals. The sampling procedures are consistent with the authors’ understanding of the style of mineralization and structural controls in the various deposits. The authors’ examination of drill cores, particularly in regard to the recognition of mineralized intervals verified the soundness of the core sampling procedure.

Cube’s site visit to Cerro Moro between 12th and 20th December 2009 included field inspection and confirmation of the location of drill holes, the mineralised outcrops, and mineralisation in both drill core and RC percussion holes.

A comprehensive description of data verification at Cerro Moro prior to 2009 is detailed in Bargmann et al (2009), Williams & Perkins (Exeter, 2009a) and Williams (Exeter, 2009b). All three of these documents have been previously lodged with the Canadian Securities Administrators and are available for viewing on SEDAR at http://www.sedar.com. For completeness some of the sections from the previous reports are repeated below, either in full or part.

12.1	QAQC Mincorp

Extorre has no documents supporting the QAQC procedures implemented or conducted by Mincorp. To the best of the author’s knowledge, no pulps or coarse rejects from the original Mincorp laboratory assaying still exist. Similarly, none of the remaining samples from the RC percussion drill holes could be located on-site. With respect to diamond drilling, approximately 70% of half core is stored on-site in varying condition. It is Extorre’s opinion that re-drilling of Mincorp drill holes is warranted in areas of interest or importance.

12.2	QAQC Exeter / Extorre

12.2.1

Geochemical Standards

Geochemical standards have been used by Extorre in all geochemical sampling and drilling programmes at Cerro Moro. During the period June 2003 to March 2007, one standard was inserted for every 40 samples. From March 2007, one standard was inserted for every 20 samples. The majority of standards are provided by Geostats Pty. Ltd. (“Geostats”) of Australia. To date, a total of 2,698 standards have been submitted from over 20 different standard types with recommended assay values varying from 0.33 to 47.24 Au ppm. The assays provided by ACME have been statistically analysed and separated into categories according to the recommended standard values. This analysis demonstrated that the greatest irregularities are associated with the lowest grade standards, however, standards with higher than recommended values have returned more acceptable levels.

During 2009 to early 2010, Extorre routinely inserted 19 different commercially Certified Reference Materials (CRM) for laboratory quality control. As of 12th February 2010 (data cut off date for April 2010 resource estimate), Extorre had submitted approximately 1,000 CRMs and 1,800 blanks since the beginning of 2009. Cube undertook a review of the 19 CRM results for the period 2009 to February, 2010 (Coupland, 2010). In general, blanks showed acceptable performance with only 5 values exceeding 0.03 Au ppm. During this period, the majority of gold and silver CRM's performed within the limits expected from a well managed and robust QAQC program and there does not appear to be evidence of systematic error or material bias. Cube made some random checks of the database for CRM values that were outside two standard deviations of the expected value and noted that in all cases Extorre had highlighted these for follow-up.

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During early 2010 to early 2011, Extorre has continued the regular insertion of CRM´s and blank samples as part of the company's standard QAQC procedure. Since the 12th February 2010 (data cut off date of April 2010 resource estimate, Coupland, 2010), Extorre has routinely inserted 29 different CRM's for laboratory quality control. As of 28th February 2011 (data cut off date for April 2011 resource estimate), Extorre had submitted approximately 2,200 CRMs and 2,200 blanks.

Cube undertook a review of the 29 CRM results for the period February 12th 2010 to February 28th 2011. As with the previous period, blanks showed acceptable performance with only 5 values exceeding 0.03 Au ppm. In general, apart from a small number of obvious errors, the majority of gold and silver CRM's have also performed within expected limits and there does not appear to be evidence of systematic error or material bias.

12.2.2

RC Percussion Duplicate Samples

A total of 2 duplicate RC percussion samples were submitted to ACME for the only RC percussion drill hole reported during 2009. Both samples originally assayed below the detection limit of the assaying method, as did the assays of the duplicate samples.

12.2.3

Diamond Drilling Duplicate Samples

A total of 251 duplicate diamond drill samples were selected for ACME Chemex. As demonstrated in Figure 27 the correlation between the original and duplicate samples for gold is within accepted limits.

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Figure 27 Diamond Duplicate Samples vs Original Sample – Au ppm

As demonstrated in Figure 28 to Figure 30 the correlation between the original and duplicate samples for silver for the three different assay methods are also within accepted limits; hot aqua regia digestion and ICP-ES (G1d-ES), by 4 acid digestion and AAS (8TD), and by fire assay with a gravimetric finish (G6 GRAV).

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Figure 28 Diamond Duplicate Samples vs Original Sample – Ag ppm by Code G1D-ES

Figure 29 Diamond Duplicate Samples vs Original Sample – Ag ppm by Code 8TD

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Figure 30 Diamond Duplicate Samples vs Original Sample – Ag ppm by Code G6 GRAV

12.2.4

Check Assaying of Samples Greater than 1ppm Au

As part of the assaying procedure, ACME re-assayed all samples that returned gold values of greater than 1.0 Au ppm with a second 50 gram fire assay and AAS finish (same method as original assay). Re-assaying is not routinely performed for samples returning gold values greater than 10 Au ppm as these samples are re-assayed by fire assay with a gravimetric finish. The plot of the original, versus re-assaying samples (a total of 1,354 pairs) shows reasonable correlation indicating relatively low variability within the pulp sample (refer to Figure 31).

This re-assaying procedure also includes the geochemical standards (a total of 1,341 pairs), which also shows reasonable correlation (refer to Figure 32).

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Figure 31 ACME Re-assays of diamond samples greater than 1 Au ppm

Figure 32 ACME Re-assays of geochemical standards greater than 1 Au ppm

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12.2.5

Blanks

Geochemical blanks prepared from barren quartz have been submitted into the geochemical and drilling sample stream. From June 2003 to March 2007, the blanks were inserted every 40 samples. From March 2007, one blank was inserted for every 20 samples. These were offset with the standards giving a control sample every 10 samples.

Approximately 90% of the results assayed either below the detection limit (“BDL”) of 0.005 or the value 0.005 Au ppm.

12.2.6

Twinned Holes

As stated in Section 12.1, Extorre has no documentation of any QAQC procedures implemented or conducted by Mincorp. There are only three Mincorp drill holes at the Escondida prospect (all diamond drill holes), all located at Escondida West, and the two more significant drill holes were twinned.

Graphical logs of the 4 drill holes, displaying gold and silver results are presented in Coupland (2010).

12.2.7

Database

Since early 2007, the entering of new assay data has been the sole responsibility of Extorre’s Chief Draftsman and Database Manager in Argentina. The Cerro Moro ‘database’ originally consisted of a series of separate MS Excel files containing collar locations, down hole surveys, geological logging, and assays. A relational database utilising MS Access was developed during the last quarter of 2007, and came into general use during 2008.

Validation of the database has been a four stage process:

1. Perusal of existing Quality Control products (i.e.: blanks, geochemical standards and duplicates) to identify areas for further investigation.

2. Visual checking by the project geologist.

3. Detailed validation of all raw data.

4. Checking by senior Exeter personnel for simple errors utilising MS Excel and Micromine software packages.

12.3	Data Verification by Cube

12.3.1

Drill Hole Collar Location

Cube independently surveyed 69 drill hole locations using a hand held Garman 12XL GPS unit. Of the 69 holes only 60 holes had final survey locations recorded in the database. The remaining 9 holes were relatively recent and only had planned coordinates recorded in the database. All checked holes were drilled during 2009 and covered the full length of the Escondida prospect. No significant discrepancies were detected and all coordinates were within ±5m which is within the accuracy of the GPS unit used.

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12.3.2

Database Validation

Cube undertook an extensive verification of the drill holes database whilst on site. Database verification included:

	1.	A detailed overview of the database structure with Exeter's on-site database manager;
	2.	A detailed check of 1 in 10 drill holes drilled in 2009 including:
	a)	Cross-validation of sample numbers from sample cutting sheet, sample dispatch tag book, Exeter dispatch number, laboratory job number and database;
	b)	Cross-validation of certified reference material numbers between sample dispatch book and database;
	c)	Cross-validation of original signed and scanned ACME lab assay reports against database;
	d)	Check of downhole survey between original digital Reflex records and database;
	e)	Check of database drill hole collar coordinates against original contract surveyor records.

12.3.3

Independent Laboratory Checks

Cube requested the retrieval of a list of pulp or coarse reject material from ACME for selected mineralised intervals for approximately 1 in 10 holes covering all 2009 diamond drilling. This material was shipped directly to Genalysis Laboratory Services in Perth, Western Australia. An independent analysis of 213 pulp/reject samples was carried out by Genalysis using similar analytical techniques to that employed by ACME. Figure 33 and Figure 34 shows excellent correlation across all grade ranges for gold and silver between the ACME and Genalysis laboratories.

Figure 33 Laboratory Checks - ACME vs Genalysis - Au ppm

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Figure 34 Laboratory Checks - ACME vs Genalysis - Ag ppm

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12.3.4

Cube Comments

The author has undertaken reasonable endeavours assess the veracity of drilling data for the Cerro Moro project. It can be concluded that all logging, sampling and data QAQC procedures during 2009 and Quarter 1, 2010 have been carried out to a high industry standard and record keeping and database management is excellent. Since the author's previous site visit in December, 2010, all logging, sampling and data QAQC procedures have been consistent.

The author believes that the current database provides an accurate and robust representation of the Cerro Moro project and is appropriate for ongoing resource evaluation.

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13.	MINERAL PROCESSING AND METALLURGICAL TESTING

13.1	Metallurgical Testwork

13.1.1

Introduction

The concentrator feed for the Cerro Moro Project comprises ore from several deposits – Escondida (which includes the Central, East, West and Far West areas), Loma Escodida, Esperanza, Gabriela, and Deborah. The ores can be broadly described as silver - gold in nature and have been categorised by oxidation state. Both open pit and underground mining will be undertaken to extract the ore.

The test work conducted to date has concentrated on the high tonnage ore types listed in Table 17 with flowsheet development based on results for Escondida as outlined in the following reports:

	●	Comminution Test Work Conducted Upon Samples of Ore from Cerro Moro Gold Project, AMMTEC Report No A11721, December 2008;
	●	Initial Metallurgical Test Work on the Cerro Moro Silver/Gold Project, Argentina; Metcon Report M1668, February 2009 including Roger Townend & Associates mineralogy report 22301;
	●	Cerro Moro Au-Ag Leach Residue Mineralogy; MODA, June 2009;
	●	Metallurgical Test Work Conducted Upon Samples of Ore from Cerro Moro Gold Deposit; AMMTEC Report No A12671, May 2010 and Report No A12791, August 2010;
	●	Metallurgical Characterisation Tests - Cerro Moro Mineralised Shoots - Escondida Far West, Metcon Report M2026, March 2010;
	●	Flowsheet Development Test Work on the Cerro Moro Au/Ag Project; Metcon Report M2014, November 2010;
	●	Flowsheet Development Test Work on the Cerro Moro Au/Ag Project; Metcon Report M2254, 2011;
	●	Thickening of Cerro Moro Gold Leach Feed, Outotec Report S1474TB, February 2011;
	●	Metallurgical Test Work Conducted Upon Test Work Products for Cerro Moro Project, ALS AMMTEC Report No A13274, February 2011;
	●	Cerro Moro Pressure Filtration Test Work, Ishigaki Oceania Pty Ltd, February 2011;
	●	Transportable Moisture Test Report TML002-96, Australian Testing Sampling & Inspection Services Pty Ltd, March 2011.

The high grade of silver relative to gold indicates that the use of the Merrill Crowe process is favoured over carbon adsorption for the recovery of the precious metals after cyanidation leaching.

13.1.2

Representivity of Metallurgical Sampling

The Representivity of Metallurgical Sampling is discussed in section 11.3.4 of this document.

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13.1.3

Testwork Samples

The major domains for metallurgical test work were identified as:

	●	Escondida Central Fresh;
	●	Escondida Central Oxide;
	●	Escondida East;
	●	Escondida West;
	●	Escondida Far West;
	●	Loma Escondida;
	●	Gabriela;
	●	Esperanza; and
	●	Esperanza South East.

Diamond drill hole (DDH) core samples were selected from the various ore types to represent the major domains. As noted in Metcon Report M1668 (February 2009) the selection of intercepts to generate composites for metallurgical testing of the individual domains was done by Exeter Resource Corporation metallurgical and geological management on the following basis:

“The drill logs, assays and descriptions were reviewed by metallurgical and geological management and the potentially economic intercepts were selected on the basis of grade, grade continuity, rock type, alteration type and width. Generally a minimum down-hole width of 2 metres was used and a cut-off grade of 0.5g/t Au was applied however these were not rigorous criteria and discretion was applied on some occasions”.

Quarter core was used to generate the test composites for each domain.

After the initial phase of testing the majority of the subsequent test work was conducted on a composite representing the Escondida ore type. The makeup of this composite on a weight for weight basis was as follows:

●	Escondida Central Fresh	28.3%
●	Escondida Central Oxide	26.1%
●	Escondida East	26.1%
●	Escondida West	19.5%

13.1.4

Testwork Water

Test work was conducted with tap water from Perth and Sydney.

13.1.5

Testwork Analyses

Gold analyses on solid samples were conducted by fire assay with an AAS (atomic absorption spectroscopy) finish while the silver analyses on solid samples were conducted by acid digestion and

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ICPms (inductively coupled plasma mass spectroscopy) or AAS finish. Gold and silver analyses on solution samples were analysed by ICPms (inductively coupled plasma mass spectroscopy). Sulphur analyses on solid samples were conducted by Leco. Base metal analyses on solids were by acid digestion and ICPms finish and base metal analyses on solutions were by ICPms.

All analyses were conducted by Ammtec in Perth, Western Australia in accordance with their standard quality control procedures.

13.1.6

Mineralisation

The deposits are classified as low sulphidation (sulphide minerals are about 5% of the ore) epithermal gold – silver ore where the mineralisation is associated with sulphides (pyrite, sphalerite, chalcopyrite and galena) in quartz veins that typically occur along fault zones. Alteration of the surrounding rock by the hydrothermal fluids and faulting has produced illite clays. The host rocks include:

	●	Rhyolite – essentially a fine grained granite, typically felsic (silica rich) and sometimes brecciated with clasts of quartz and feldspar, and accessory biotite/mica and hornblende;
	●	Rhyodacite;
	●	Andesite – porphyritic texture with phenocrysts of feldspar but little or no quartz;
	●	Felsic tuff – silicate minerals (quartz, mica, muscovite, feldspar) often brecciated;
	●	Volcaniclastic sediment.

As the veins in the deposits are narrow with widths ranging from 1 m to 5 m (Gabriela has veins up to 20 m wide in the central and south-eastern part of the deposit), the feed to the plant will include the host rocks and the illite clays from the alteration zone adjacent to the veins and these will impact on the operation and performance of the treatment plant.

The main minerals present based on the geology and the mineralogy ofleach tails from oxide and fresh Escondida ore, mineralogy of flotation concentrates and gravity concentrates from Escondida and Gabriela ore, and head sample assays are:

	●	Quartz (SiO2) – The quartz is generally fine grained and occurs as chalcedony in places. Phenocrysts/clasts of quartz are also present and these would form pebbles in a SAG mill and rock scats from ball mills. The quartz will cause high abrasion in crushing and grinding equipment’;
	●	Illite/mica - Illite is a very fine grained phyllosilicate sheet mineral (a form of mica) that will increase the viscosity of the slurry in the plant and reduce the settling and filtration rates in dewatering. The platy nature of most forms of mica usually causes a coarse grind and a very high work index to be reported due to the large two dimensional size of the particles however, illite grains range from 0.1 to 0.3 µm in diameter and are only 30 Å thick compared to most phyllosilicates that have grains up to 2.0 µm thick and are up to 4.0 µm wide. Therefore breakage during grinding into composite grains of apparent coarse platy

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particles does not inflate the work indices measured when illite is present. Illite is a non-expanding mica/clay in that it does not absorb water with expansion of its layer structure (as smectites such as bentonite and fuller’s earth do) and so excessive viscosity is not expected at most slurry densities that will be used in the treatment plant. Nevertheless the fineness of the particles will blind cloths in filtration and result in low settling rates in thickening;

	●	Kaolin/kaolinite (Al4Si4O10(OH)8) - Clay particles 0.3 to 4.0 µm wide and 0.05 to 2.0 µm thick will affect the viscosity, thickening and filtration rates;
	●	Chlorite ((Mg,Fe,Al)6(Si,Al)4O10(OH)8 – Chlorite can increase the viscosity and yield stress of slurry;
	●	Feldspar/microcline (KAlSi3O8).

The feed will therefore comprise predominantly silicate minerals that are fine grained which will increase the power required for size reduction and the wear life of equipment linings and grinding media. The mineralised veins ofquartz – adularia (KAlSi3O8 a variety of microcline or orthoclase feldspar) contain gold and silver and associated sulphides as ginguro bands that alternate with the quartz and other silicates. Ginguro is a term used to describe a fine grained aggregate, usually dark in colour that consists of electrum, silver (sulphide) minerals and sulphide minerals such as chalcopyrite (CuFeS2), sphalerite ((Zn,Fe)S), galena (PbS) and pyrite (FeS2). The texture within the veins is also often brecciated.

Gold occurs as native gold, argentian gold (up to 23% Ag) or electrum (an alloy of gold and silver with silver contents reported by Roger Townend to range from 40% to 76% Ag in the Escondida and Gabriela ores). The electrum and argentian gold occur as discrete particles and are often occluded within pyrite, acanthite (Ag2S) or sphalerite, or as mineral grains in association with galena, acanthite and/or chalcopyrite – see Figure 35 to 13.6. Native gold can be readily recovered by gravity, flotation and cyanidation processes. Electrum does not respond well to flotation and as for argentian gold, has a lower rate of dissolution in cyanide than native gold (and requires a higher concentration of cyanide to achieve good recovery in equivalent leach times) but electrum is recovered readily by gravity means. Fine gold minerals were seen in binary, ternary and quaternary composites and inclusions within other sulphide minerals indicating that recovery of gold (and silver) will be maximised by producing a sulphide concentrate from gravity and flotation circuits or by ultrafine grinding prior to leaching.

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Figure 35 20 µm electrum(white) at galena/pyrite contact, also sphalerite

Figure 36 50µm argentian gold within coarse pyrite also containing galena

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Figure 37 Electrum of 6 µm in chalcopyrite enclosed by 100 micron acanthite

Figure 38 Two electrums (63% silver) 5 and 10 µm in pyrite, composite with sphalerite and jalpaite, also titanium oxide.

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Silver is present as acanthite (Ag2S containing 87% Ag), tetrahedrite/freibergite ((Cu,Ag)10(Fe,Zn)2Sb4S13 with up to 34% Ag content) and in electrum and argentian gold. Generally there will also be silver within the lattice of other sulphide minerals associated with epithermal deposits – in galena, chalcopyrite and pyrite. The cyanidation rate of silver and silver minerals is slower than for gold and higher cyanide concentrations, finer grind sizes and longer leach times are generally required. The higher cyanide concentrations result in higher residual cyanide levels in the tailings stream which will require detoxification/neutralisation. The silver minerals are generally softer than the other minerals present and tend to be overground during grinding. Combined with their naturally fine grained nature and relatively lower specific gravity (7.5 for acanthite compared to 19.3 for gold), the silver recovery tends to be lower than gold in gravity circuits. The flotation kinetics of the silver sulphide minerals is moderately high although slimes losses will occur. As the silver minerals will preferentially grind to a smaller size than the majority of the other minerals in the ore, a proportion of the silver minerals will report more readily to the cyclone overflow – the amount of acanthite in the circulating load relative to the other sulphide minerals will be lower - and therefore may not be presented to gravity or flash flotation which are typically fed a portion of the cyclone underflow stream. This suggests that a portion of the mill discharge should be treated by unit cell flotation or by gravity.

Figure 39 Zoned electrum from 5 to 100 µm composite with acanthite, sphalerite and chalcopyrite.

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Figure 40 Electrum of 10 µm in acanthite in sphalerite

Table 17 summarises the mineral occurrence and associations found from optical and scanning electron microprobe analysis of flash flotation and gravity concentrates generated in laboratory tests.

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	Escondida. Central Fresh Flash Conc.	Escondida. Central Oxide Flash Conc.	Escondida. East Flash Conc.	Gabriela Flash Conc.	Escondida. Central Fresh Gravity Conc.	Escondida. Central Oxide Gravity Conc.	Escondida. East Gravity Conc.	Gabriela Gravity Conc.
Dominant Mineral	Pyrite 20-400µm 75% discrete	Pyrite 20 -400µm 50% discrete; composite with or disseminated in quartz	Pyrite 20 - 300µm discrete & with inclusions of galena & chalcopyrite	Pyrite 20 - 250µm discrete	Pyrite 50% free, 20% inclusions in quartz	Pyrite 50% discrete, 35% disseminated in pyrite	Pyrite 50 -150µm mostly discrete; fine inclusions in silicates	Goethite discrete
Major Mineral	Sphalerite mostly discrete	Sphalerite <100µm mostly discrete	–	–	Sphalerite discrete & composite with galena	–	–	–
Major Mineral	Galena composite with sphalerite	–	–	–	Galena	–	–	–
Minor Minerals	-	Galena<100µm; chalcopytite <100µm; sphalerite mostly discrete 150µm	–	–	–	–	–	Hematite micro-crystalline
Accessory Minerals	Chalcopyrite, galena, acanthite	Acanthite; galena associated with acanthite	Sphalerite <100µm mostly discrete; chalcopyrite	Acanthite rarely discrete, sphalerite composite with & in pyrite; galena	Acanthite discrete to 200µm, composite with pyrite, sphalerite, galena. jalpaite	Sphalerite discrete & composite with galena; galena mostly composite; acanthite	Sphalerite discrete <150µm	Pyrite disseminated euhedral in silicate; magnetite; sphalerite composite with pyrite, jalpaite, quartz
Trace Minerals	Electrum, freibergite, jalpaite	Pearcite, jalpaite, electrum, argentite gold, chalcopyrite, covellite; stromeyerite 50µm rimmed with native silver as rims on pyrite	Galena, acanthite, bornite, electrum, argentite gold, jalpaite, marcasite	Chalcopyrite & composite with galena or included in sphalerite; digenite; covellite rims pyrite; acanthite; electrum; jalpaite; argentite gold; native silver	Bornite, chalcopyrite included in acanthite; electrum ; argentite gold	Jalpaite; electrum; argentite gold; covellite; stromeyerite <50µm rimmed & veined by chalcopyrite	Galena composite with sphalerite&in pyrite; chalcopyrite composite with pyrite & silicate; bornite	Galena; chalcopyrite

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Escondida. Central

Fresh

Flash Conc.

Escondida. Central

Oxide

Flash Conc.

Escondida. East

Flash Conc.

Gabriela

Flash Conc.

Escondida. Central

Fresh

Gravity Conc.

Escondida. Central

Oxide

Gravity Conc.

Escondida. East

Gravity Conc.

Gabriela

Gravity Conc.

Electrum occurrence	40% silver, inclusions in pyrite, composite with galena & sphalerite	52% silver composite with pyrite or acanthite, & within pyrite	76%silver composite with acanthite & pyrite; disseminated in acanthite	Sub-µm in pyrite & acanthite	Composite with pyrite, sphalerite & galena; 50% silver	In pyrite	Discrete	In sphalerite
Argentian Gold occurence	Inclusions in pyrite	In chalcopyrite& pyrite	90µm discrete	23%silver very fine inclusions in pyrite; native silver very fine with acanthite in pyrite	Very fine in galena	23% silver; in or with jalpaite, acanthite, chalcopyrite	20µm with galena in sphalerite; composite with pyrite	<5µm in pyrite & acanthite
Acanthite occurrence	100µ discrete, inclusions in pyrite & sphalerite; composite with pyrite & galena	coarse discrete & in pyrite	Discrete to 200µm; composite with or in pyrite, galena, & silicate	Discrete to 200µm; composite with pyrite, chalcopyrite, sphalerite, galena; fines in pyrite			Included in pyrite & composite with galena or feldspar	Up to 300µm; fine in silicate associated with jalpaite& pyrite
Jalpaite occurrence	70µm discrete; composite with freibergite	<200µm discrete		100µm composite with pyrite or chalcopyrite		Composite with sphalerite, quartz, pyrite	With galena as composites in pyrite & quartz	100µm hosting finer acanthite in quartz; 25-50µm in sphalerite & chalcopyrite

Table 17 Summary of flash flotation & gravity concentrate mineralogy findings (Escondida & Gabriela)

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The modal analysis of leach tailings indicated that the silver occurred mainly as acanthite (accounting for 96% of the silver in the fresh ore tailing and 93% in the oxide ore tailing). 3% of the silver was present in tetrahedrite for the fresh ore tailing and 6% in the oxide ore tailing. The liberation and composite particle analysis for acanthite is shown in Table 18 and indicates that liberation is higher for oxide ore.

	+53 µm Fresh	+53 µm Oxide	-53+20 µm Fresh	-53+20 µm Oxide	-20 µm Fresh	-20 µm Oxide	Overall Fresh	Overall Oxide
Weight %	26	21	27	23	47	56	100	100
% Liberated	48	62	19	60	92	93	67	76
% Binary with Pyrite	10	6	4	3	0	0	3	2
% Binary with Sphalerite	29	18	12	0	6	0	11	5
% Binary with Chalcopyrite	0	2	2	14	0	6	0	7
% Binary with Gangue	0	3	20	15	0	0	5	5
% Ternary	6	6	42	7	2	1	12	4
% Quaternary	8	3	0	0	0	0	1	1

Table 18 Liberation and composite details for Acanthite (Ag2S) in leach tailing (Escondida)

The mineralogy liberation analysis indicates significant compositing of the acanthite with other sulphide minerals and that a fine grind of 20 to 50 µm is required to liberate the acanthite especially for the fresh ore. It also indicates that regrinding of flotation and gravity concentrates to approximately 20 µm will liberate the majority of the acanthite that has reported to tailing. Without regrinding, upgrading of flotation and gravity concentrates using cleaning stages is likely to result in losses of gold and silver due to the complex associations and fineness of included grains. Based on the calculated assays for the particles analysed, 43% of the silver was lost in the -20 µm fraction and 31% in the +53 µm fraction of the fresh ore and 53% in the -20 µm fraction and 23% in the +53 µm fraction of the oxide ore – the highest losses were in the fines. Most acanthite particles in the coarse fractions showed corrosion on the surface indicating that leaching had occurred and that longer times or finer grinding may be required to increase silver leach recovery. The presence of a high proportion of liberated acanthite particles in the tailings (67% and 76% for the fresh and oxide ores respectively) also indicates there was insufficient time for the mineral to leach. It may also indicate that some passivation of the surface has occurred that slowed the leach rate or that insufficient cyanide was present.

The variety of minerals and the complexity of the mineral associations of the precious metals mineralogy (especially the proportion of ternary and quaternary particles containing tetrahedrite) therefore indicate that there is justification for both gravity and flotation processes to maximise recoveries with the processes tending to complement each other depending on the variation in mineral occurrence. For example, flaky or platelet gold formed during grinding of the ductile native gold tends to flutter into the tailing stream in gravity concentrators. This gold is able to be recovered in flotation.

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13.1.7

Comminution

The unconfined compression strength (UCS) of the ore ranges from 81 MPa to 222 MPa with an average of 120 MPa and is considered to be moderately tough. The measured abrasion indices ranged from 0.1 to 0.5. The ore with the highest toughness occurs with the finer grained volcanic rocks. As the majority of the gangue comprises silica, the abrasion index is high and tough rock will need to be broken, selection of a double toggle jaw crusher will minimise the wear rate of liners in the primary crusher however a double toggle jaw crusher is not available in the capacity range required so a single toggle jaw crusher was selected.

The SMC breakage data indicates the ore overall is of medium hardness relative to the data basewith individual samples ranging from soft to moderately hard. The Bond ball mill work indices ranged from a moderate 14.4 kWh/t to a high (hard) 19.9 kWh/t. The respective rod mill work indices were only slightly higher than the ball mill work indices with ratios ranging from 0.95 to 1.17 indicating there would be little build up of a critical size in a semi-autogenous or autogenous mill. It also indicates that sufficient competent media may be lacking at times and a high ball charge would be needed to effect grinding in a semi-autogenous (SAG) mill.

Variations in the feed ore breakage characteristics will result in fluctuating plant treatment rates in a SAG mill operation which will cause a significant effect on downstream processes as the relative effect is amplified by the low average throughput of 42 t/h. Therefore, a three stage crush and single stage ball mill comminution process has been selected to provide a stable treatment rate.

Calculations using both the conventional Bond – Roland method and SMCC’s latest method for the power required to reduce the ore from a P80 of 8.5 mm to the target P80 of 75 µm at a rate of 42 t/h indicate a mill capable of drawing 1,300 kW is required.

13.1.8

Gravity Concentration

As discussed above, the mineralogy indicates that inclusion of both flotation and gravimetric devices will maximise recovery of the silver and gold minerals. The expected gravity recoverable gold and silver varies from 2% to 30% for each metal. Flash flotation recovery will vary from 10% to 55% for each of gold and silver. To maximise the gravity recoverable gold a flash flotation machine and a centrifugal gravity separator will treat >75% of the cyclone underflow.

The cyclone underflow has been selected as the feed stream to the gravity circuit as it will be deslimed thereby overcoming viscosity issues in flotation. Gold and silver in the cyclone overflow stream (particles that have not been recovered in the flash flotation machine and gravity device in the cyclone underflow) will report to the leaching circuit.

Test work indicated that increasing the flotation time from 1 minute to 2 minutes improved the recovery of both gold and silver into flash flotation concentrate particularly for high grade ores (Metcon report M2026). Tests were conducted at 50% solids to overcome the viscosity associated with the illite clays

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and at natural pH which ranged from 7 to 8. A scale up factor of 2.5 has been applied to the laboratory test time which gives a required cell volume of 8.5 m³ (allowing for 15% air hold up). An 8 m³ SK240 flash flotation machine has been selected. A manually adjustable froth crowder will enable mass pull variations of 2% to 10% (relative to plant feed).

As per the optimum test work flowsheet, tailing from the flash flotation machine will report directly to a centrifugal concentrator to recover the precious metals minerals that were not collected in the flotation stage due to mineralogy, particle surface chemistry or particle size. Variable cycle times or bowl configurations can be used to change the mass pull to the gravity concentrate. A 20 minute cycle has been selected for design to maximise the mass recovery into concentrate.

13.1.9

Flotation and Gravity Concentrate Treatment

The concentrate from the gravity and flotation stages will be reground and report to the leaching circuit. Intensive cyanidation test work indicated that a regrind P80 size of less than 62 µm was sufficient to extract over 94% of both the gold and silver. Without regrinding, extraction of gold and silver was lower at 92% for gold and 79% for silver. Copper extraction into solution ranged from 17% to 29% and appeared to be affected by copper grade and mineralogy with high cyanide consumptions when extraction of copper was higher.

The consumption of sodium cyanide was generally higher at 13.5 kg/t for the reground concentrate – unground concentrate consumed 12.6 kg/t of sodium cyanide. The effect of regrind sizes finer than 30 µm on leach time and reagent consumption has not been assessed and indications are that reagent consumption is driven by the mineralogy.

The leach kinetic curves shown in Figure 41 indicate that silver reaches an equilibrium after 8 to 16 hours while gold continues to leach, although at a low rate for a further 8 hours at least. These test were done under intensive cyanidation conditions similar to those that occur in an Acacia reactor – 10% solids, 0.2% caustic (NaOH), 1% to 2% sodium cyanide and 12 kg/t to 25 kg/t leach aid. Dissolved oxygen levels were 7 ppm to 8 ppm. The intensive cyanidation tests were done using 200 g of sample. Leach kinetics tended to be faster when the cyanide concentration was maintained at a higher value, but 24 hour extractions were not affected. The process design provides for batch leaching the concentrate under intensive cyanidation conditions. The concentrate leach residue will return to the main leaching circuit to ensure gold and silver extraction is maximised and the pregnant solution will be sent to the Merrill Crowe circuit for precious metal recovery.

No test work to generate a signature plot of grind size versus power has been done to accurately size the concentrate regrind mill. A specific power of 40 kWh/t has been assumed based on similar duties for sulphide flotation concentrates. A 100 litre IsaMill which is powered by a 75 kW motor has been selected for the duty. This mill will have additional capacity to handle higher concentrate mass pulls, higher hardness concentrate or a greater size reduction.

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Figure 41 Concentrate Leach Kinetics – Effect of Regrind

13.1.10

Leaching

A summary of test work results showing overall extractions for gold and silver using the flash flotation, gravity and cyanide leach flowsheet is given in Table 20. The results show that overall extractions are

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generally high for both gold and silver. Test results (summarised in Table 21) also indicated that compared to flotation of the gravity tail, leaching of the gravity tail resulted in an overall higher extraction of 10% for gold and 5% for silver.

The leaching test work indicated typical slower leach kinetics for the silver minerals compared to the gold. The leach decay curves indicate that dissolution of silver continues 48 hours of leaching as indicated in Figure 42, 13.9 and 13.10. Several tests indicate that re-adsorption of gold onto solids occurred after 32 hours of leaching when the available free cyanide concentration decreased below 500 ppm. Therefore, stage addition of cyanide to maintain a suitable free cyanide concentration will be necessary. A cyanide concentration of 0.1% has been selected in the design. High residual cyanide levels in the tailings will therefore require a cyanide destruction circuit to minimise possible impacts on the environment from tailings disposal.

A leach pH of 10.5 has been selected to minimise generation of hydrogen cyanide gas.

Most tests were conducted without the injection of oxygen. Tests using oxygen sparged for up to 4 hours showed higher initial dissolution rates but extractions over 48 hours were similar. Nevertheless, as the reground flash flotation and gravity concentrates will be included in the leach stream and to minimise losses when high grade feed is treated, the design includes provision to add oxygen into all leach tanks.

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Report	Sample Description	Test	Calculated Feed			Flash	Gravity	Gold Extraction					Silver Extraction					Comment
Report	Sample Description	Test	g/t Au	g/t Ag	%S	% Wt	% Wt	Flash	Gravity	Conc Leach	Leach	Total	Flash	Gravity	Conc Leach	Leach	Total	Comment
M1668	Escondida Central - Fresh	CMF3,CML1	19.5	565	2.21	2.8	3.4	31.6	56.6	99	11.5	98.8	43.0	30.4	96	18.6	89.0	Series 1
		CMF21,CML5	19.5	390	2.25	3.8	1.0	49.6	29.5	99	20.5	98.8	64.7	7.3	96	17.5	86.6	Series 2
	Escondida Central - Oxide	CMF7,CML2	19.1	692	2.19	3.1	3.5	32.5	44.3	99	22.7	98.7	37.0	27.7	96	25.2	87.3	Series 1
		CMF22,CML6	18.9	536	2.10	4.0	0.9	43.2	11.3	99	44.6	98.5	55.1	7.0	96	25.8	85.4	Series 2
	Escondida East	CMF11,CML3	9.7	132	1.16	2.0	2.9	37.3	47.5	99	14.7	98.6	51.6	8.0	96	31.3	88.5	Series 1
		CMF23,CML7	8.08	118	1.07	2.4	0.5	50.9	20.7	99	27.8	98.7	55.1	2.3	96	30.0	85.1	Series 2
	Gabriela	CMF13,CML4	0.95	134	0.51	1.5	3.1	35.2	8.9	99	53.3	97.0	51.2	7.4	96	34.4	90.6	Series 1
		CMF24,CML8	0.85	120	0.53	1.5	0.8	43.7	1.6	99	48.9	93.7	48.8	1.9	96	38.1	86.8	Series 2
	Escondida West	CMF25,CML9	4.44	146	0.93	2.3	0.9	48.7	4.7	99	44.1	97.0	47.8	3.1	96	41.0	89.8	Series 2
	Esperanza	CMF26,CML10	1.91	186	1.11	2.1	1.0	42.1	2.0	99	50.6	94.2	41.3	4.2	96	45.1	88.8	Series 2
	Esperanza South East	CMF27,CML11	3.16	83	0.21	1.2	1.2	61.1	1.5	99	35.5	97.5	49.7	3.6	96	37.1	88.2	Series 2
	Loma Escondida	CMF28,CML12	9.11	443	1.17	2.6	1.4	44.2	6.4	99	46.0	96.1	57.1	4.0	96	33.6	92.2	Series 2
M2026	Escondida Far West	Oxide	36.7	2,475		1.4	2.0	18.5	29.6	99	48.1	95.7	11.7	22.9	96	34.6	67.8
		Low Grade 1	4.44	254	2.12	2.3	2.6	32.9	19.7	99	52.6	104.6	29.3	8.0	96	37.3	73.1
		Low Grade 2	3.82	235		2.6	2.6	45.6	15.8	99	36.0	96.8	43.5	8.6	96	42.3	92.3	Inc ti, xan
		High Grade 1	33.6	2,245	4.55	3.7	2.5	17.8	29.6	99	47.4	94.3	16.6	14.9	96	31.5	61.7
		High Grade 2	31.6	2,129		8.1	2.5	45.7	19.6	99	33.2	97.8	47.9	12.7	96	34.8	93.0	Inc ti, xan
M2014	Escondida Composite		11.9	405	1.57	1.4	1.9	16.7	47.4	99	34.9	98.3	19.6	32.0	96	38.3	87.8
A12671	MD292A	HS22238/85/99	1.46	25	1.8	7.1	2.1	22.7	31.5	99	42.7	96.3	25.9	10.5	96	45.7	80.6
	MD293	HS22239/86/300	9.27	443	1.11	3.8	1.7	52.2	15.0	99	32.1	98.6	61.4	4.1	96	22.9	85.7
	MD294	HS22240/87/301	75.1	1032	1.86	5.2	2.5	47.7	38.7	99	13.4	98.9	60.7	13.5	96	12.5	83.7
	MD295	HS22241/88/302	1.97	158	0.29	2.9	2.2	25.7	11.5	99	59.0	95.8	36.7	10.0	96	42.5	87.3
	MD297	HS22242/82/96	2.92	19.4	0.03	8.0	2.2	26.7	23.4	99	48.9	98.5	13.6	15.5	96	61.6	89.5
	MD298	HS22243/83/97	8.79	130	0.03	8.5	2.2	22.0	9.9	99	66.6	98.2	20.0	16.9	96	53.4	88.8
	MD299	HS22244/84/98	8.24	592	0.84	4.7	2	28.6	21.0	99	47.0	96.1	39.0	6.8	96	41.5	85.5
A12791	MD293		9.11	445	1.09	3.7	2.2	58.4	9.7	99	31.3	98.7	62.5	3.1	96	22.6	85.6
		Average				3.6	2.0	37.7	21.4	99	39.0	97.6	42.0	11.0	96	34.6	85.4

Table 19 Test work results showing overall extractions for gold and silver using the flash flotation, gravity and cyanide leach flow sheet

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	Flash - Gravity - Flotation		Flash	Gravity	Flotn	Conc Leach	Total	Flash	Gravity	Flotn	Conc Leach	Total
M1668	Escondida Central - Fresh	CMF17	31.6	56.6	10.6	99	97.8	43.0	30.4	21.8	96	91.4	Series 1
	Escondida Central - Oxide	CMF18	32.5	44.3	19.1	99	94.9	37.0	27.7	28.9	96	89.9	Series 1
	Escondida East	CMF19	37.3	47.5	12.7	99	96.5	51.6	8.0	27.4	96	83.5	Series 1
	Gabriela	CMF20	35.2	8.9	19.2	99	62.7	51.2	7.4	22.7	96	78.0	Series 1
	Flash - Gravity - Flotation	Average	34.2	39.3	15.4	99	88.0	45.7	18.4	25.2	96	85.7	Series 1
	Flash - Gravity - Leach	Average	34.2	39.3		99	98.1	45.7	18.4		96	88.8	Series 1
	Escondida Central - Fresh	CMF29	49.6	29.5	19.8	99	97.9	64.7	7.3	22.5	96	90.7
	Escondida Central - Oxide	CMF30	43.2	11.3	41.5	99	95.0	55.1	7.0	32.2	96	90.5
	Escondida East	CMF31	50.9	20.7	25.7	99	96.3	55.1	2.3	29.3	96	83.2
	Gabriela	CMF32	43.7	1.6	25.9	99	70.5	48.8	1.9	24.8	96	72.5
	Escondida West	CMF33	48.7	4.7	25.5	99	78.1	47.8	3.1	30.5	96	78.1
	Esperanza	CMF34	42.1	2.0	32.9	99	76.2	41.3	4.2	34.7	96	77.0
	Esperanza South East	CMF35	61.1	1.5	19.5	99	81.3	49.7	3.6	23.7	96	73.9
	Loma Escondida	CMF36	44.2	6.4	23.4	99	73.3	57.1	4.0	24	96	81.7
	Flash G____ - Flotation	Average	47.9	9.7	26.8		83.6	52.5	4.2	27.7		81.0	Series 2
	Flash G_____ - Leach	Average	47.9	9.7			96.8	52.5	4.2			87.9	Series 2

Table 20 Test work results showing overall extractions for gold and silver using the flash flotation, gravity and flotation flow sheet

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Figure 42 Effect of Primary Grind Size on Flotation-Gravity Tail Leach – Escondida Central

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Figure 43 Effect of Primary Grind Size on Gravity Tail Leach – Escondida Central

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Figure 44 Effect of Primary Grind Size on Flotation - Gravity Tail Leach – Escondida Far West

Due to the inclusion of flash flotation and gravity machines in the grinding circuit and the consequential highwater additions required for their effective operation, the cyclone overflow density will be low

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(<35% solids). A leach feed thickener has been included in the design to enable the density for the leach circuit to be increased – this reduces the size of the leach circuit, and will recover non-cyanide contaminated waterfor use in the grinding, flotation and gravity circuits.

The leach feed density will beapproximately 40% solids to minimise any viscosity issues in the leach circuit. The leaching circuit has been designed to provide a minimum residence time of 48 hours. Five stages of leaching have been selected to minimise short circuiting, and to provide flexibility to maximise control of the leaching conditions.

Figure 42 shows the leach curves for different primary grind sizes – 45 µm to 150 µm. The 75 µm grind for test HS22599 gave a higher extraction but was from a different sample. The differences are within the variation shown by other tests and no clear trend is apparent although the 45 µm grind resulted in the highest extractions for both gold and silver. Leach kinetics were similar for silver. The initial 4 hour leach kinetics for gold was faster as the grind size decreased. The results indicate little difference between the 75 µm and 45 µm grind sizes in overall extraction, therefore a 75 µm grind size was selected.

Based on the rheological issues discussed in Part 13.1.11 for thickening and filtering, there is scope to coarsen the grind with minimal impact on overall extractions.

13.1.11

Merrill Crowe

The high silver content of the leach solution has dictated the selection of cementation of the precious metals onto zinc powder rather than adsorption onto carbon.

The Merrill Crowe circuit has been designed on the basis of information from existing operations and laboratory test work as reported in AMMTEC Report no A12791(August 2010). The critical design parameters and values used in the design are listed in Table 21. After the zinc addition the oxygen levels were less than 1 ppm. In the test work a zinc dose rate of 80 times stoichiometric was selected although preliminary tests showed 40 times stoichiometric gave recoveries from solution of >99% for both gold and silver. It is considered that the 80 times stoichiometric dose is excessive and based on operating data from other operations a dosage rate of 15 times stoichiometric is more realistic. This figure (15 times stoichiometric) has been used in the design criteria and to determine operating costs.

Addition of lead nitrate was found not to affect recovery once sufficient zinc was added and the solution was de-aerated.

The design envisions recycling the barren solution from the Merrill Crowe circuit to the back of the counter current decantation (CCD) circuit as wash water and also using this solution to provide the flocculant dilution waterfor the CCD thickeners. Excess barren solution will be diverted to the cyanide destruction circuit to ensure cyanide is not recycled to the process water circuit.

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Design Parameter	Value	Source
Zinc Dose	15 times stoichiometric	Operating Data
Lead Nitrate	No addition required	Test Work
Gold Precipitation	99% (solutions were 1.68 ppm gold)	Test Work
Silver Precipitation	99% (solutions were 59 ppm silver)	Test Work

Table 21 Critical Design Data for Merrill Crowe Circuit

13.1.12

Thickening and Filtration

Separation of the leach solution and the leach tailing solids is required to provide the pregnant solution for the Merrill Crowe circuit and to determine options for disposal of tailings (dry stacked or wet conventional). The relatively small proportion of illite and kaolinite clays in the ore has adversely impacted on the rheological characteristics of the slurry. Metcon reported that at a shear rate of 2.5 sˉ¹ the viscosity of the leach slurry exceeded 1 Pas at densities >48% solids when lime was used to achieve a pH of 10.5. The fine nature, sheet-like grain geometry and electrical charge of the illite (micaceous) clay has resulted in low settling rates, low ultimate settled densities and low filtration rates. The particle size distribution for the tailings is shown in Figure 45 with more than 30% in the minus 10 µm size fraction. Consequently, the rheology of the leach slurry is a major aspect for option, design and equipment selection to recover the pregnant solution for the Merrill Crowe.

Options include:

	●	CCD circuit;
	●	Filtration and cake washing; and
	●	Combination of thickening and filtration with cake washing.

Figure 45 Leach Tailing Particle Size Distribution

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Initial settling tests by Metcon and BASF investigated combinations of coagulants and flocculants to optimise thickening parameters. Outotec conducted their standard dynamic thickening test work. Samples of the thickener tailing were then despatched to Delkor and GBL for filtration test work. Table 23 and 13.8 summarise the results of the test work.

Design Parameter	Metcon	Outotec	Outotec	Outotec
Coagulant Dose g/t	1500 (lime)	250 (Magnafloc 368)	250	300
Flocculant Dose g/t Magnafloc 5250	70	20	20	25
Underflow/Settled Density % solids	39 – 43	43	46	45
Specific Settling Rate t/m²h	0.11 – 0.16	0.43	0.30	0.21
Overflow Clarity ppm solids	NA	390	270	510

Table 22 Summary of Initial Thickening Test Work

The coagulant and flocculant dosage requirements will increase the operating costs. The specific settling rates are up to 10 times lower than typical tailings settling rates.

Design Parameter	Delkor	GBL
Coagulant/Flocculant Dose g/t		Residual
Feed Density % solids	46.6	35
Cake Moisture %	16 – 29	20 – 27
Specific Filtration Rate kg/m²h	17.5 – 30	86 – 166

Table 23 Summary of Initial Filtration Test Work

The filter cake, although having a high moisture content (27%), was compact and soft with a consistency of potter’s clay. Nevertheless, the cake tended to be friable when rubbed. The filtration rates were up to 10 times lower than typical for tailings. Typical plasticine-like filtration cakes from test work are shown in Figure 46.

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Figure 46 Filter cakes from tests showing plasticine nature.

Further thickening and filtration tests were undertaken to optimise the separation conditions. Thickening test work was conducted by using a higher feed pH (Outotec Report S1474TB) than the 10.3 pH used previously by Outotec – cylinder settling tests indicated that lime addition with a pH >11 resulted in clear supernatant. Using lime to adjust the feed pH above 10.67, an improvement in settling performance was achieved. An underflow density of 56% solids at a specific settling rate of 1.00 t/m2/h was achieved with a flocculant dosage of 20 g/t of Magnafloc 5250. The overflow clarity was less than 100 mg/L. The yield stress of the underflow was 93 Pa and the slurry exhibited shear thinning behaviour. This work is reported in Outotec Report S1474TB (February 2011). It is considered that the improvement in performance was due to the surface charge modification properties of the lime as this test work used the same composite as the previous test work. Therefore the thickener design has been based on this latest test work. A specific settling rate of 1.00 t/m2/h has been used for the leach feed and tailings duties and a conservative specific settling rate of 0.50 t/m2/h for the CCD duty. The lower rate in the CCD duty will allow for process and ore type variations to ensure a high clarity in the pregnant solution directed to the Merrill Crowe circuit.

The design has incorporated lime addition to the leach feed and tailings thickeners to provide the high pH in the thickener feed (the CCD thickener is already at an elevated pH courtesy of the leaching circuit). If the leach feed and tailings thickener overflows are untreated this will result in an elevated pH in the process water circuit which may adversely affect the flotation of sulphides in the grinding circuit. The design provides for the addition of sulphuric acid to the settling pond to ensure the pH of the process water remains approximately neutral.

Vendor filtration test work was also conducted by Ishigaki Oceania. The test work demonstrated that a specific filtration rate of 100 kg/m2/h with a filter cake moisture content of 18% to 21% can be achieved with a filter feed density of 37% to 43% solids in a vertical chamber filter press with a cake squeeze of 15 to 20 bar and no air blow of the filter cake. The filter cake was very compressible as evidenced by the effectiveness of the filter cake squeeze. While there was no overall improvement in filtration performance over the previous test work the ability to achieve acceptable cake moistures without an air blow of the filter cake would prolong filter cloth life in a full scale operation.

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The plastic nature of the cake indicated that to effectively wash the cake in order to recover the residual pregnant liquor (which based on a cake moisture content of 21% represents approximately 16% of the pregnant solution in the leach tailing), the cake would need to be repulped with barren solution from the Merrill Crowe circuit and refiltered to effect a wash. Several filter washing stages would be required to provide an acceptable recovery of the gold and silver in solution. It was therefore considered that filtration did not offer any process advantage and would incur higher capital costs than an equivalent CCD circuit.

The flow point and transportable moisture limit of the tailings filter cake were 18.2% and 16.4% respectively. It is considered likely that the tailings would liquefy in transport and therefore that co-disposal of the tailings using truck transport will require leak proof containers and would present difficulties in unloading the contents. The tailings disposal design has therefore been based on thickening and pumping to either a conventional paddock or integrated landform tailings storage facility.

13.1.13

Cyanide Destruction

The leach tailings were tested for cyanide destruction using the Inco process. This work is reported in ALS AMMTEC Report no A13274 (February 2011). The critical parameters and the values used in the test work are summarised in Table 24. The design has been based on a sulphur dioxide dose of 6.0 g SO2 per g of WAD cyanide, a lime dose of 1.0 g of lime per g of SO2 and a copper concentration of 150 ppm. WAD cyanide levels of approximately 7 ppm were achieved without the addition of copper sulphate and of less than 0.5 ppm with copper sulphate present.

The tests were conducted for up to 106 minutes of retention time. A total time of 120 minutes has been selected for design. Due to the local and environmental conditions, the cyanide detoxification circuit has been designed for 100% redundancy – two cyanide destruction tanks will be installed, each having 120 minutes of residence time.

Design Parameter	Unit	Value
Initial WAD Cyanide Concentration	ppm	306
Sulphur Dioxide Dose	g SO2 / g WAD Cyanide	5 - 7
Lime Dose	g Lime / g SO2	0.7 – 1.3
Copper Concentration	ppm	30 – 150
Residual WAD Cyanide Concentration	ppm	0.2 - 8

Table 24 Summary of Cyanide Destruction Test Work

13.1.14

Metallurgical Recoveries

The metallurgical recoveries selected for financial evaluation of the Escondida ore are 95% for gold and 87.4% for silver. This has been calculated from the leach extraction data for individual ore sources and proposed mining schedule on a weighted average basis. It assumes a CCD wash efficiency of 99% and Merrill Crowe precipitation efficiency of 99%.

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13.1.15

Recommendations

The following additional test work is recommended:

	●	Test work to optimise the consumption of the cyanide, copper sulphate in flotation and zinc because of the significant impact these reagents have on operating cost;
	●	Test work to quantify the benefits of a dedicated Merrill Crowe circuit to recover precious metals from the pregnant solution generated by intensive cyanidation;
	●	Test work to quantify the benefits of electrowinning of precious metals from pregnant solution generated by intensive cyanidation;
	●	Test work on any new potential ore bodies to verify metallurgical performance;
	●	Test work on individual ore bodies to be repeated using a P80 of 500 µm for the flash flotation test work as per vendor recommendations and with intensive cyanidation of the resultant concentrates. The previous test work used a P80 of 250 µm and did not test the intensive cyanidation of the resultant concentrates.

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14.	MINERAL RESOURCE ESTIMATE

14.1

Summary

A summary of the Cerro Moro Indicated and Inferred mineral resources above a cut-off of 1 ppm gold equivalent are shown in Table 25 and Table 26 respectively.

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
Escondida	620,000	18.8	829.2	35.4	374,000	16,530,000	705,000
Loma Escondida	44,000	18.4	919.5	36.8	26,000	1,297,000	52,000
Gabriela	537,000	2.4	371.0	9.9	42,000	6,411,000	170,000
Total	1,201,000	11.5	627.5	24.0	443,000	24,238,000	927,000

Table 25 Cerro Moro Indicated Mineral Resources above 1 ppm Gold Equivalent

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
Escondida	508,000	4.3	164.8	7.6	70,000	2,689,000	123,000
Loma Escondida	13,000	9.7	595.4	21.6	4,000	256,000	9,000
Gabriela	390,000	2.3	394.8	10.2	29,000	4,948,000	128,000
Esperanza	371,000	2.6	175.0	6.1	31,000	2,090,000	72,000
Deborah	579,000	2.4	48.1	3.4	45,000	896,000	63,000
Total	1,861,000	3.0	181.8	6.6	178,000	10,879,000	396,000

Table 26 Cerro Moro Inferred Mineral Resources above 1 ppm Gold Equivalent

*Gold equivalent values have been calculated by Extorre on the basis of the following parameters:

	●	Long term gold price US$1,320/oz;
	●	Long term silver price US$26/oz;
	●	Metallurgical recovery gold 100%;
	●	Metallurgical recovery silver 100%.

The gold equivalent value is calculated by dividing the silver grade (ppm) by 50 (approximate ratio of gold/silver US$ value) and adding it to the gold grade (ppm).

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Key Assumptions, Parameters, and Methods Used to Estimate Mineral Resources

It is the opinion of the Qualified Person that the Indicated and Inferred mineral resources summarised in Table 25 and Table 26 respectively have reasonable prospects for economic extraction. A number of factors have been taken into consideration when quantifying a subset of the Cerro Moro mineralised system with reasonable prospects for economic extraction. These include:

Potentially economic mineralisation at Cerro Moro occurs at relatively shallow depths, with the majority being within 200 meters of the surface. The shallow depth of occurrence indicates that exploitation may be achieved by well-established open pit or shallow underground mining methods.

	●	The gold and silver grade of the Cerro Moro mineralisation is high by world standards. High grade mineralisation at Cerro Moro occurs in well developed “shoot-like” zones surrounded by minor occurrences of low grade material within a distinct geological structure;
	●	Extorrereleased a PEA on December 2, 2010 that estimated open pit mining costs at U.S.$3.40/tonne (including tailings co-disposal) and milling and processing costs at U.S.$38/tonne. These costs provide a cut-off grade of 1.8 Au ppm gold equivalent;
	●	Based on preliminary metallurgical testwork gold and silver recoveries have been estimated at 95% and 90%, respectively;
	●	Base assumptions for long-term gold and silver prices are U.S.$1,320/oz and U.S.$26/oz, respectively;
	●	Less than 2% of the Cerro Moro gold equivalent mineral resources reported in the above Tables fall below a nominal economic cut-off grade of 1.8 Au ppm.

The Qualified Person believes that the cost and commodity price parameters set out above are consistent with those adopted by companies with similar sized projects elsewhere in the world. While the Qualified Person believes that the long-term gold and silver price assumptions of U.S.$1,320/oz and U.S.$26/oz, respectively, provide an appropriate basis for project evaluation, recent commodity spot prices have reached significantly higher levels. Since January 2011, gold and silver spot prices have consistently traded in excess of U.S.$1,320/oz and U.S.$26/oz respectively. It is the Qualified Person’s opinion that at current commodity spot prices there is a reasonable expectation that material below the 1.0 Au ppm gold equivalent cut-off used for reporting current mineral resources may become available for economic extraction.

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opportunities to exploit lower grade resources in the Cerro Moro project area. These options could be particularly attractive given the relatively shallow nature of mineralisation readily amenable to open-pit mining. Such options could result in processing costs being as low as U.S.$5.50/tonne resulting in potentially economic cut-off grades of 0.3 Au ppm gold equivalent.

14.2	Coordinate Systems

The grid system used at Cerro Moro is Zone 2 of the Argentinean Gauss-Krüger (GK) projection which is based on the Campo Inchauspe datum. Several local grids have been established over the project area which are orientated parallel to the strike of each prospect. Escondida and Loma Escondida block models have been provided in Gauss-Krüger coordinates whereas Gabriela, Esperanza and Deborah have been provided in local grid coordinates. Table 27 summarises the coordinate systems together with the relevant linear grid transformations for each prospect.

Prospect	Grid System		GK X	GK Y	Local X	Local Y
Escondida	Gauss-Krüger
Loma Escondida	Gauss-Krüger
Gabriela	Local Grid	Point 1	2673281.911	4671503.988	20000	5000
Gabriela	Local Grid	Point 2	2677425.282	4668096.813	25352	5000
Esperanza	Local Grid	Point 1	2673281.911	4671503.988	20000	5000
Esperanza	Local Grid	Point 2	2677425.282	4668096.813	25352	5000
Deborah	Local Grid	Point 1	2860338.7	4669092.7	40000	5000
Deborah	Local Grid	Point 2	2680802.8	4669465.6	40594	5000

Table 27 Cerro Moro Coordinate Systems

14.3	Bulk Density

At each deposit a statistical analysis of the available density data was undertaken and the results used to derive unique values for the assignment of bulk density to waste and variously oxidised mineralised material. Cube has not undertaken any independent analysis of density for the purposes of this estimate.

14.4

Oxidation

No oxidation surfaces were interpreted. Oxidation of mineralised domains is observed to be of insignificant depth and therefore not material to this stage of resource estimation. The effect of oxidation on metallurgical recoveries may be considered more significant in future studies.

14.5	Drilling Database

Extorre provided Cube with a complete MS Access drilling database for the Cerro Moro project area (CM_GeoData.mdb) on February 28th 2011. This database was suitable for direct connectivity to

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Surpac mining software and was subsequently renamed by cube to CM_GeoData_28022011.mdb. A description of the database and the relevant tables and fields are shown in Table 28.

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Table

Field

Description

DH_Collar	Propect	Escondida, Gabriela, etc
	HoleID	Hole Id
	Easting	Gauss-Krüger Easting
	Northing	Gauss-Krüger Northing
	LG_E	Local Grid Easting
	LG_N	Local Grid Northing
	RL	Elevation
	Depth	Total Depth of Hole
	RC	Depth of RC
	DDH	Depth of Diamond
	Type	Drillhole or Trench
	DateStart	Start Date of Hole
	DateEnd	Completion Date of Hole
	Company	Estelar (Exeter) or Mincorp
DH_survey	HoleID	Hole Id
	Depth	Downhole Depth of Survey
	DIP	Dip of Hole trace
	TrueAzimuth_GK	Gauss-Krüger Azimuth of Hole Trace
	TrueAzimuth_LG	Local Grid Azimuth of Hole Trace
DH_Assay	HoleID	Hole Id
	From	Interval Depth From
	To	Interval Depth To
	SampleID	Sample Id
	SampleType	RC, DDH, RockChip, STD, DUP etc
	QAQC	QAQC Identifier
	AuPlot_ppm	Preferred Gold Grade ppm - Numerical
	AgPlot_ppm	Preferred Silver Grade ppm - Numerical
03_StratigraphyUnit	HoleID	Hole Id
	From	Interval Depth From
	To	Interval Depth To
	StratUnit	Summarised Stratigraphy
04_Oxidatiion	HoleID	Hole Id
	From	Interval Depth From
	To	Interval Depth To
	Oxidation	SAG, Complete, Partial, Fresh
10 Vein	HoleID	Hole Id
	From	Interval Depth From
	To	Interval Depth To
	Qv% per m	Quartz Vein Percentage
zonecode	hole_id	Hole Id
	depth_from	Interval Depth From
	depth_to	Interval Depth To
	zonecode	Mineralised Intercept Code

Table 28 Drill Hole Database Structure

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All compositing, descriptive statistics, geostatistical analysis and estimation for gold and silver was based on the AuPlot_ppm and AgPlot_ppm fields respectively.

The AuPlot_ppm and AgPlot_ppm fields are a preferred value for gold and silver based on a hierarchical system of analysis techniques. Gold values below 10 ppm are based on a 50 gm fire assay whereas gold values reporting greater than 10 ppm are re-assayed by fire assay with a gravimetric finish. Similarly, silver values below 1,000 ppm are based on a four acid digest and AAS finish whereas silver values reporting greater than 1,000 ppm are re-assayed by fire assay with a gravimetric finish.

All gold grades greater than 1 ppm are routinely re-assayed and the average grade is recorded in the AuPlot_ppm database field. Cube compared the original and repeated assay values and notes that there is excellent correspondence between the two values with no apparent biases. In addition, the averaging of the two values results in little, if any, reduction in sample variability (Table 29) confirming that the averaged value can be confidently used for resource estimation.

Figure 47 Original vs Repeat Assays – >1 Au ppm

	Au1	Au2	AuAve
# samples	1309	1309	1309
Mean	8.97	8.92	8.95
Std Dev	25.82	25.89	25.85
CoV	2.88	2.90	2.89

Table 29 Database - Original vs Repeat Assays – >1 Au ppm

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Table 30 shows the numerical equivalent values assigned to below detection character values stored in the assay fields.

Au	AgPlot_ppm	Ag	AgPlot_ppm
-0.01	0.005	-1	0.5
-0.005	0 0025	-0.5	0.5

Table 30 Database - Numerical Equivalent Values for Below Detection Values

14.6	Resource Estimation

14.6.1

Escondida

Data Types

The Escondida estimate was primarily based on Extorre diamond drilling and small number of selected reverse circulation drill holes. Trenching was used to aid geological interpretation but trench grades have not been used in the estimation. The Escondida dataset comprises 622 drill holes for 90,796.50 m (Extorre 90,691.50 m, Mincorp 105 m).

Domaining and Volume Modelling

Typically, the main economic mineralised zones within the Escondida prospect area consist of a series of steeply dipping E-W to ESE vertically extensive narrow (0.1-4 m) epithermal quartz vein structures. Economic gold and silver mineralisation primarily occurs in multi-phase banded quartz-adularia and quartz-chlorite-sulphide veins, hydrothermal breccias and peripheral stockworks. The domain outlines used to control volume and estimations have been predominantly based on geological attributes and observations from drill core, particularly epithermal vein textures rather than grade criteria. In most cases, a clear distinction between a main epithermal quartz vein structure and surrounding stockwork mineralisation can be determined based on detailed geological logging and core photography (Figure 48). In addition, a number of surface trenches and sub-outcrop (Figure 49) have been used to guide the interpretation. Surrounding stockwork zones are characterised by stockwork veining of varying intensity with minor breccia zones.

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Figure 48 Escondida Drillhole MD633 Showing Styles of Mineralisation

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Figure 49 Escondida Central – Quartz Float and Sub-Outcrop

The Escondida main epithermal vein structures are characterised by distinct shoot like ‘bonanza’ zones occurring wholly within the plane of the main epithermal vein structure, effectively creating two mineralisation styles:

	●	Main Zone (MZ) – continuous material characterised by classic epithermal vein textures including crustiform/colloform chalky white quartz-adularia sulphide banded and brecciated veining - typically grading 0.5 to 10 Au ppm;
	●	Bonanza Zone (BZ) – semi-continuous identifiable zones within the Main Zone. Characterised by brecciated quartz-adularia sulphide banded vein ‘ginguro’ material and often associated with base metal sulphides and black silica - typically grading 10 to 200 Au ppm.

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It is the Author's experience with similar projects that the main epithermal vein structures (MZ) can be manually (deterministically) interpreted and modelled in 3D with a high degree of confidence directly from drill hole data. By contrast, the BZ shoots within the MZ often exhibit somewhat more complex architecture making it problematic to accurately generate 3D models from a deterministic interpretation approach. The Author's experience also confirms the necessity to separate the MZ and BZ mineralisation events during estimation due to the distinct geological and statistical differences between these domains.

The domaining and volume modelling approach is threefold:

Main Zone (MZ): Cross-sections were generated in accordance with the drill hole spacing which typically ranges between 10m-20m spacing. Sectional interpretations of the main epithermal vein structure using geological logging and core photography was completed over the full strike extent of the Escondida prospect. The resulting MZ mineralisation interpretation was graphically digitised and a 3D wireframe model produced. The interpretation of the MZ is based on geological attributes such as epithermal vein textures ensuring that the final 3D mineralisation model reflects an in-situ geological model whereby no cut-off grade or minimum mining width criteria has been applied. Figure 50 and Figure 51 show the Escondida main epithermal vein zones together with drilling locations and surface topography.

Bonanza Zone (BZ): The BZ domains are ‘shoot-like’ zones wholly contained within the MZ. As mentioned earlier, the BZ shoots exhibit a somewhat more complex architecture than the encompassing MZ zones. The BZ domains require separate treatment for estimation and the approach to defining the BZ estimation domains is discussed in more detail in Section 14.6.1 “Domaining of Bonanza Shoots”.

Figure 50 Escondida Main Epithermal Vein Zones – Plan View

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Figure 51 Escondida Main Epithermal Vein Zones – Looking North-East

Stockwork: A surrounding stockwork interpretation was generated in a similar manner to the MZ domains and a 3D wireframe model produced. Whilst the stockwork domain was predominantly based on geological logging and core photography, anomalous gold and silver grades outside the MZ were also used as a guide. The stockwork domain typically extends 2-5m from the MZ hangingwall and footwall contacts and encompasses stockwork veining of varying intensity and minor breccia zones.

Database Coding and Compositing

Samples within the MZ zones and surrounding stockworks are assigned a unique database code which is used to control the compositing process. A four digit numbering system (zonecode) was created to define the various domains within the Escondida prospect (Table 31).

Zonecode	Domain
1101	Escondida Central-East Main Zone
1201	Escondida West 2 Main Zone
1301	Escondida Far West Main Zone
1401	Escondida West 1 Main Zone
1901	Escondida Footwall Stockwork
1902	Escondida Hangingwall Stockwork

Table 31 Escondida Domain Numbering

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Main Zone (MZ+BZ)

There are several key physical features of the Escondida epithermal vein zones that need to be considered and accommodated by the selected resource modelling technique. These features are:

	●	Mineralised epithermal vein zone true thickness is variable typically ranging from 0.1m to 4m in width.
	●	Undulating or variable zone geometry (dip and strike) and possible grade/thickness trends within this variable geometry;
	●	Sampling has been taken over geological intervals creating samples of unequal length or variable support;
	●	The continuity and geometry of the ‘Bonanza Zone’ (BZ) within the ‘Main Zone’ (MZ) is quite variable making it difficult to apply separate deterministic domaining to the BZ;
	●	Mining selectivity across the epithermal vein zones is unlikely due the narrow nature of the mineralised structures;
	●	Drill spacing is variable but typically around 20mN by 20mE within the plane of mineralisation.

A method that allows estimation of metal content on a projected plane is considered to be an appropriate approach to addressing the features outlined above. This resource modelling approach is best achieved using geological intercept composites and accumulation estimation. MZ grades are composited across the entire coded interval resulting in a single geological intercept composite at each intercept location. The geological composites are projected onto a vertical 2D plane approximately parallel with the vein structure. The mid-point of each geological composite is assigned the horizontal width of the vein structure and used to compute a ‘metal accumulation’ variable.

Geological intercept composites are not of equal support as the lode thickness varies. Given the variable thickness, an additive variable must be created as the product of grade and thickness or ‘metal accumulation’ variable. The accumulation a(x) is defined as the product of thickness t(x) and grade z(x) assuming a constant density:

a(x) = t(x) . z(x)

Stockwork

Traditional downhole composites of 1m were generated for the surrounding stockwork domains whereby the zonecode flagging was used to control compositing. The downhole composites were extracted using a ‘best fit’ method which optimises the composite length over the coded interval to eliminate the problem of residuals.

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Descriptive Statistics and High Grade Capping

Main Zone (MZ)

An examination of gold and silver raw sample statistics within the coded intervals was undertaken prior to compositing. It is the Author’s observations from similar deposits that a small number of statistical outliers can occur at sample support leading to statistical instability of geological composites. Raw sample statistics were examined for the individual and combined main zone domains for Escondida. It was decided to apply a modest high grade assay cut of 600 Au ppm and 15,000 ppm to gold and silver raw samples respectively prior to geological compositing. Both high grade assays cuts are above the 99th percentile of the raw sample distribution. A high grade assay cut of 600 Au ppm applies to 5 individual gold samples out of 1,736 and has the effect of reducing the raw sample gold grade by around 3.3%. Similarly, a high grade assay cut of 15,000 Ag ppm applies to 14 individual silver samples out of 1,734 and has the effect of reducing the raw sample silver grade by around 5.9%. Whilst the application of high grade capping is somewhat arbitrary it is considered by the Author to be a prudent and necessary risk adjustment strategy at this stage of the project evaluation. Figure 52 and Figure 53 show log-probability plots of raw gold and silver samples respectively for the individual and combined (cyan) Escondida MZ domains.

Figure 52 Escondida Main Zones – Raw Samples – Au ppm – Log-Probability Plot

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Figure 53 Escondida Main Zones–Raw Samples – Ag ppm – Log Probability Plot

Geological intercept composites were generated after a high grade assay cut of 600 Au ppm and 15,000 Ag ppm had been applied to the raw samples. It is important to consider the raw and declustered grade statistics together with the spatial distribution of grade when analysing a mineral deposit. Furthermore, when dealing with unequal length composites such as geological intercept composites it is necessary to incorporate width and the accumulation in the analysis. Figure 54 to Figure 58 show the spatial distribution of raw intercept grades, horizontal widths and accumulations for the Escondida MZ domains. It can be clearly seen from Figure 54 and Figure 65 that high grade gold and silver values form semi-continuous shoot-like occurrences within the plane of the main vein. In addition, Figure 56 indicates that areas of elevated gold and silver grades appear to be coincident with increased vein width. This positive association between grade and width is common in epithermal vein systems and often reflects structural features such as dilation zones or flexures in zone geometry. Not surprisingly, the association between grade and width is particularly evident when visualising the accumulation values as shown in Figure 57 and Figure 58.

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Figure 54 Escondida Main Zones – Geological Composites – Au ppm - Long Section

Figure 55 Escondida Main Zones – Geological Composites – Ag ppm – Long Section

Figure 56 Escondida Main Zones – Geological Composites – Horiz. Width - Long Section

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Figure 57 Escondida Main Zones–Geological Composites – Au*HW (gxm) – Long Section

Figure 58 Escondida Main Zones–Geological Composites – Ag*HW (gxm) – Long Section

Figure 59 to Figure 63 show log-probability plots of raw intercept grades, horizontal widths and accumulations for the individual and combined (cyan) Escondida MZ domains. Whilst individual zones exhibit differing global statistics it could be reasonably argued that all zones share similar distribution characteristics which is consistent with the similar geological, structural and mineralisation framework between zones.

No additional grade capping was considered necessary other than that described in Section 14.6.1 “Descriptive Statistics and High Grade Capping”

A list of Escondida geological intercept composites is included in Section 14.8.

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Figure 59 Escondida Main Zones – Geological Composites – Au ppm – Log-Probability Plot

Figure 60 Escondida Main Zones – Geological Composites – Ag ppm – Log-Probability

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Figure 61 Escondida Main Zones – Geological Composites – Horiz. Width – Log-Probability Plot

Figure 62 Escondida Main Zones – Geological Composites – Au*HW (gxm) – Log-Probability Plot

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Figure 63 Escondida Main Zones – Geological Composites – Ag*HW (gxm) – Log-Probability Plot

A notable feature of the drilling configuration is that there is greater drilling density (by design) in areas of elevated grade compared to the lower grade peripheral areas. As such, declustering will be required in order to compute robust global statistics. By way of illustrating this point, Figure 54 shows a comparison between raw intercept gold grade and declustered intercept gold grade using various declustering cell dimensions for the combined Escondida MZ data set. It is clearly shown that the grade decreases as the declustering cell size becomes larger reflecting a weighting toward wider drill spacing in lower grade areas. It could be reasonably expected that an unbiased global grade estimate of the Escondida MZ domains would be roughly equivalent to a 60m x 60m declustered intercept composite grade based on the overall drill spacing. A comparison such as this forms a key component of model validation at a later stage.

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Figure 64 Escondida Main Zones – Declustered Grade Comparison – Au ppm

Table 32 and Table 33 summarises combined Escondida MZ raw geological intercept grades (cut and uncut) and accumulation (cut only) statistics respectively. It can be seen from Table 32 that declustering has a significant impact on the average intercept grade reducing the global gold and silver grade by around 39%. It is also important to note that the width weighted declustered mean grades (highlighted in red) in Table 33 are significantly higher grade than the raw declustered mean grades (highlighted in blue) in Table 32. This confirms statistically that there is a positive association between elevated grade and zone width within the Escondida MZ domains. Furthermore, the association between grade and width necessitates the application of an accumulation style method of estimation for the Escondida epithermal main vein domains.

	Raw Au ppm	Raw Ag ppm	Cut Au ppm	Cut Ag ppm
Number	465	465	465	465
Minimum	0.00	0.40	0.00	0.40
Maximum	290.06	8668.31	223.05	7130.15
Raw Mean	15.61	588.81	15.13	563.23
Declust. Mean 60x60	9.29	348.44	9.27	344.57
Std Dev	33.61	1197.77	31.17	1089.06
Coeff Var	2.15	2.03	2.06	1.93

Table 32 Escondida Main Zones - Summary Statistics – Geological Composites - Raw Grades

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	Horizontal Width	Cut Au X HW	Cut Ag X HW	HWidth Weighted Cut Au ppm	HWidth Weighted Cut Ag ppm
Number	465	465	465	437	437
Minimum	0.13	0.00	0.25
Maximum	5.51	494.30	20113.89
Raw Mean	1.43	25.69	1024.65	18.01	718.05
Declust. Mean 60x60	1.37	15.53	608.45	11.34	444.12
Std Dev	0.87	57.21	2356.03
Coeff Var	0.61	2.23	2.30

Table 33 Escondida Main Zones - Summary Statistics – Geological Composites - Accumulations

Stockwork

Log-probability plots of gold and silver 1m composites for the Escondida combined stockwork domains are shown in Figure 65 and Figure 66 respectively. An assay top cut of 3 Au ppm and 200 Ag ppm was applied to gold and silver respectively as a model risk adjustment strategy. Table 34 shows summary statistics of raw and cut gold and silver 1m composites for Escondida stockwork mineralisation.

	Raw Au ppm	Raw Ag ppm	Cut Au ppm	Cut Ag ppm
Number	2628	2628	2628	2628
Minimum	0.00	0.12	0.00	0.12
Maximum	12.76	702.00	3.00	200.00
Raw Mean	0.27	15.03	0.25	13.84
Std Dev	0.57	39.39	0.40	28.32
Coeff Var	2.14	2.62	1.57	2.05

Table 34 Escondida Stockwork Zones - Summary Statistics – 1m Downhole Composites

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Figure 65 Escondida Stockwork Domains – 1m Downhole Composites – Au ppm – Log-Probability Plot

Figure 66 Escondida Stockwork Domains – 1m Downhole Composites – Ag ppm – Log-Probability Plot

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Domaining of Bonanza Shoots

As mentioned in Section 14.6.1 “Domaining and Volume Modelling” the main epithermal vein zones host distinct shoot like ‘bonanza’ zones occurring wholly within the plane of the main epithermal vein structure. While the BZ zones appear to exhibit sharp grade boundaries, the continuity and geometry of these zones is quite variable making it difficult to apply separate deterministic domaining. Nevertheless, it is the Author’s experience from similar deposits that the BZ zones require separate domaining for the purposes of grade estimation to avoid potential over-smoothing of the bonanza grades.

Cube adopted an Indicator Simulation (Sequential Indicator Simulation - SIS) approach to objectively define the BZ and MZ domains. Exploratory data analysis showed that a gold grade indicator of 10 Au ppm resulted in excellent definition of the BZ shoots effectively creating a grade based separation of BZ and surrounding MZ. The MZ domain was further split into a low grade and moderate grade domain using a gold grade indicator of 2 Au ppm. Figure 67 and Figure 68 show a statistical and spatial representation of the indicator thresholds applied for the Indicator Simulation.

Figure 67 Escondida Main Zones - Gold Indicator Thresholds – Log-Probability Plot

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Figure 68 Escondida Main Zones –Geological Composites – Au ppm – Indicators - Long Section

Figure 68 clearly shows that the spatial distribution of BZ is not evenly distributed throughout the project area. It was determined that the results of the SIS could be improved by estimating a localised BZ and MZ probability at each location. This was achieved by interpolating the BZ and MZ indicators using Ordinary Kriging into each simulation grid node. This process effectively creates a 2D probability map of each indicator similar to that shown for BZ in the Central and East Zone (Figure 69). An omni-directional variogram model based on the BZ indicator threshold was fitted for the combined Escondida main zone dataset and used as the basis for the SIS (Figure 70).

Figure 69 Escondida Central-East Zone – BZ Indicator Probability Map

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Figure 70 Escondida Main Zones – BZ Indicator Variogram Model

The SIS process incorporates a Simple Kriging step which uses these localised probabilities as a local mean during simulation. A sequential neighbourhood was defined whereby a maximum of 16 initial data points and 10 already simulated nodes were used to condition the simulation. A set of 100 simulation realisations of the BZ and MZ indicators was performed into 2D 1m x 1m grid nodes. The use of SIS allows the generation of a series of equally probable geological models whereby a statistically derived ‘most probable’ geological/mineralisation model can be determined. Figure 71 shows examples of several simulated domain models together with the statistically derived ‘most probable’ geological model. The SIS process was repeated separately for the four Escondida main zone domains (1101 - 1401). The ‘most probable’ geological model was re-grouped into X=10m x Z=10m blocks whereby each larger block has an estimated proportion of each mineralisation style.

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Figure 71 Escondida Central-East Zone – Simulated Domains

Variography

Main Zone (MZ+BZ)

Geological composites for individual main zone domains were separated on the basis of the gold grade indicator. Variography was used to characterise the spatial continuity of the horizontal width and accumulation variables for each grade indicator within the plane of each vein structure. It was decided to combine the low grade MZ and medium grade MZ data to improve the quality of variography for these indicators. Variography demonstrated similar spatial behaviour for each of the four main zone domains (1101-1401) and it was also decided to combine these into a single dataset in order to generate robust global variogram models for the combined Escondida main zones. No obvious direction of preferred continuity was determined and omni-directional variogram models were fitted for all variables. Variograms of horizontal width and accumulations demonstrated similar spatial structure and variance proportions. Given the similarities between these variograms a simplification was made whereby a single variogram model was used to estimate both the accumulation and thickness variables. This approach has the advantage that it is simpler and ensures no instability in the back-calculated block grades due to unstable or unexpected width estimations which may occur when a different variogram model is used. Experimental variography was undertaken on Gaussian transformed data. Variogram models were fitted to the Gaussian data and subsequently back-transformed into real variances to obtain the appropriate parameters for interpolation of the raw accumulation variables by Ordinary Kriging. Fitted Gaussian based variogram models and back-transformed variogram model parameters for the Escondida ‘Bonanza Zones’ (BZ) are shown in Figure 72 and Table 35 respectively.

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Figure 72 Escondida Bonanza Zones (BZ) – Variogram Models – Gaussian Au*HW and Ag*HW

		Sill (Relative Variance)	Range	Azimuth	Plunge	Dip	Major/ Semi Major Ratio	Major/ Minor Ratio
Au*HW	Nugget Co	0.38
	Structure 1	0.33	30	0	0	0	1	1
	Structure 2	0.29	100	0	0	0	1	1
Ag*HW	Nugget Co	0.44
	Structure 1	0.30	30	0	0	0	1	1
	Structure 2	0.26	100	0	0	0	1	1

Table 35 Escondida Bonanza Zones (BZ) – Variogram Models – Back-Transformed Raw Au*HW and Ag*HW

Fitted Gaussian based variogram models and back-transformed variogram model parameters for the Escondida ‘Main Zones’ (MZ) are shown in Figure 73 and Table 36 respectively.

Figure 73 Escondida Main Zones (MZ) – Variogram Models – Gaussian Au*HW and Ag*HW

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		Sill (Relative Variance)	Range	Azimuth	Plunge	Dip	Major/ Semi Major Ratio	Major/ Minor Ratio
Au*HW	Nugget Co	0.34
	Structure 1	0.43	34	0	0	0	1	1
	Structure 2	0.23	100	0	0	0	1	1
Ag*HW	Nugget Co	0.28
	Structure 1	0.36	25	0	0	0	1	1
	Structure 2	0.37	100	0	0	0	1	1

Table 36 Escondida Main Zones (MZ) – Variogram Models – Raw Au*HW and Ag*HW

The separation of the main zone data on the basis of the indicators described in 1.6.1.5 results in well structured experimental variography for both the BZ and MZ domains. The BZ gold and silver accumulations exhibit moderate relative nugget effects of 38% and 44% respectively with maximum ranges in the order of 100m. Similarly, the surrounding MZ (low grade + medium grade) domains exhibit moderate relative nugget effects of 34% and 28% respectively with maximum ranges also in the order of 100m.

Stockwork

The Escondida stockwork domains (1901 and 1902) were combined for the purpose of variography. 3D experimental variography was undertaken on cut Gaussian transformed 1m down hole composites. Variogram models were fitted to the Gaussian data and subsequently back-transformed into real variances to obtain the appropriate parameters for interpolation of the raw accumulation variables by Ordinary Kriging. Back-transformed variogram model parameters for the Escondida stockwork zones are shown in Table 37.

		Sill (Relative Variance)	Range	Azimuth	Plunge	Dip	Major/ Semi Major Ratio	Major/ Minor Ratio
Au	Nugget Co	0.23
	Structure 1	0.27	3.8	306	0	85	1	3.8
	Structure 2	0.23	20	306	0	85	1	4.0
	Structure 3	0.27	60	306	0	85	1	6.0
Ag	Nugget Co	0.29
	Structure 1	0.22	3.8	306	0	85	1	3.8
	Structure 2	0.39	20	306	0	85	1	4.0
	Structure 3	0.10	60	306	0	85	1	6.0

Table 37 Escondida Stockwork Zones – Variogram Models – Raw Au ppm and Ag ppm

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Grade Estimation

Main Zone (MZ+BZ)

Estimation of block grades within the main zone domains used an accumulation methodology whereby the gold and silver accumulations and horizontal width were estimated on a 2D vertical plane into X=10m x Z=10m blocks using Ordinary Kriging for each grade indicator. Final block grade is calculated by dividing the estimated accumulation by the estimated thickness and weighted by the proportion of each mineralisation style (BZ and MZ) computed previously by the SIS approach. The estimation of horizontal width used for block grade back-calculation used the same variogram models as for the accumulations. This simplification ensures no instability in the back calculated grades due to unstable or unexpected width estimations which may occur when a different variogram is used. Separate estimation was undertaken for each of the four main zone domains (1101-1401) using the uniquely coded composite data. The adopted estimation methodology resulted in very good reproduction of the ‘bonanza’ shoots within the main vein zones.

Kriging neighbourhood analysis was undertaken to determine appropriate inputs to the Ordinary Kriging process. Estimation parameters for the BZ and MZ domains are summarised in Table 38 and Table 39 respectively.

Parameter	CutAuXHW	CutAgXHW	HWidth
Minimum number of Comps	2	2	2
Maximum number of Comps	16	16	16
Search Major Distance	120	120	120
Search Orientation	000	000	000
Plunge of Major Axis	0	0	0
Dip of Major Axis	0	0	0
Anisotropy major/semi-major	1.0	1.0	1.0
Anisotropy major/minor	1.0	1.0	1.0
Block Discretization	10 x 10	10 x 10	10 x 10

Table 38 Escondida Bonanza Zones (BZ) – Estimation Parameters

Parameter	CutAuXHW	CutAgXHW	HWidth
Minimum number of Comps	2	2	2
Maximum number of Comps	16	16	16
Search Major Distance	120	120	120
Search Orientation	000	000	000
Plunge of Major Axis	0	0	0
Dip of Major Axis	0	0	0
Anisotropy major/semi-major	1.0	1.0	1.0
Anisotropy major/minor	1.0	1.0	1.0
Block Discretization	10 x 10	10 x 10	10 x 10

Table 39 Escondida Main Zones (Low Grade MZ + Medium Grade MZ) – Estimation Parameters

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Stockwork

The stockwork mineralisation surrounding the main epithermal vein zones is estimated using traditional 3D Ordinary Kriging of the uniquely coded 1.0m downhole composite data.

All block estimates were based on grade interpolation into X=10m x Z=10m x Y=2m parent cells. Estimation parameters for the Stockwork domains are summarised in Table 40.

Parameter	Cut Au	Cut Ag
Minimum number of Comps	2	2
Maximum number of Comps	16	16
Search Major Distance	120	120
Search Orientation	000	000
Plunge of Major Axis	0	0
Dip of Major Axis	0	0
Anisotropy major/semi-major	1.0	1.0
Anisotropy major/minor	1.0	1.0
Block Discretization	10 x 10	10 x 10

Table 40 Escondida Stockwork Zones – Estimation Parameters

3DBlock Model Definition

At the completion of 2D block grade back-calculation each of the X=10m x Z=10m parent cells was re-located into real-world space and imported into a 3D block model. The primary consideration of the 3D models was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each mineralised zone. All individual mineralised zones were ultimately combined to create a single block model in the Gauss-Krüger coordinate system. The Escondida block model definition is shown in Table 41. A standard list of field names and descriptions used in the block model are shown in Table 42.

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Origin	Minimum	Maximum	Model Extent
Y	4,667,500	4,669,500	2,000
X	2,673,000	2,676,500	3,500
Z	-500	200	700
Parent Cell Y	2m	Min Sub-Cell Y m	0.25
Parent Cell X	10m	Min Sub-Cell X m	1.25
Parent Cell Z	10m	Min Sub-Cell Z m	1.25

Table 41 Escondida 3D Block Model Definition

Field Name	Description
x	X Block Centroid
y	Y Block Centroid
z	Z Block Centroid
au	Au ppm – Interpolated
ag	Ag ppm – Interpolated
aueq60	Gold Equivalent ppm = Au + Ag/60
density	Density g/cm3 – Direct Assignment – Default 2.5
rescode	Resource Classification - 1=Measured 2=Indicated 3=Inferred
zone	Domain Code Eg. 1101, 1201, 1301, 1401, 1901 , 1902

Table 42 Escondida Block Model Attribute Names

Bulk Density

Global bulk densities were assigned as follows:

●

Main Epithermal Vein Zones - 2.65 g/cm3 based on 58 on-site density determinations.

●

Stockwork Zones - 2.5 g/cm3 based on 237 on-site density determinations.

Oxidation

Oxidation of mineralised domains is observed to be of insignificant depth and does not appear to materially influence the bulk density or characteristics of the main mineralised zones. Whilst trenching and sub-outcrop indicates that the Escondida main zone is very close to the surface in places, Cube has assigned a zero grade to all material within 2m below the topographical surface.

Final Model

The final Surpac block model was clipped by the topography to remove volume above current surface and exported into CSV format. The following Escondida models in the Gauss-Krüger coordinate system were provided to Extorre:

esc_mar2011_min_zones.mdl – Surpac format block model with sub-blocking to blocks of Y=0.25m, X=1.25m and Z=1.25m

esc_mar2011_min_zones_07042011.csv – CSV export of above Surpac model

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esc_mar2011_min_zones_reg_07042011.csv – CSV export of above Surpac model regularised to blocks of Y=1m, X=1.25m and Z=2.5m

esc_mar2011_ppct_min_zones_29052011.csv – CSV export of above Surpac model zone partial percentages within blocks of Y=2m, X=10m and Z=10m.

Model Validation

Model validation was carried out visually and by comparing the modelled outcomes against de-clustered composite grades. Figure 74 and Figure 75 show back-calculated model grades for gold and silver together with geological intercept composites for the combined Escondida epithermal main zone (MZ+BZ). The model outcomes show very good correspondence to the informing composite, in particular, providing very good definition of the bonanza shoots. Statistical comparisons of composite grades (raw and declustered width weighted) with the modelled outcomes are shown in Table 43 for the combined Escondida epithermal main zones. Cube concludes that both visual and statistical comparisons confirm that a robust and unbiased model outcome has been achieved.

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Figure 74 Escondida Main Zones (MZ+BZ) – Composites vs Model Grades – Au ppm

Figure 75 Escondida Main Zones (MZ+BZ) – Composites vs Model Grades – Ag ppm

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Zone	De-clustering Cell Size	Raw Composite Grades Cut Au ppm	De-clustered Composite Grade Au ppm (HW wt)	Model Grade Au ppm	Raw Composite Grades Ag ppm	De-clustered Composite Grade Ag ppm (HW wt)	Model Grade Ag ppm
1101-1401	X=60 x Z=60	15.1	11.3	11.6	563.2	444.1	501

Table 43 Escondida Combined Main Zones - De-Clustered Composite vs Model Grade

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Resource Classification and Reporting

The mineral resource estimate was prepared by Mr Ted Coupland, MAusIMM, CPGeo, Director and Principal Geostatistician of Cube. The mineral resource estimates have been classified and reported in accordance with the CIM guidelines (CIM 2005) National Instrument NI 43-101. Mr Ted Coupland is ‘independent’ and a ‘qualified person’ as defined by NI 43-101.

Cube has classified a substantial proportion of the Escondida mineral resources as Indicated where drilling is 20m x 20m or closer. Cube believes that this level of drill hole spacing is sufficient to demonstrate a high level of confidence in the geometry, continuity and grade of the Escondida deposit. Surrounding areas of Escondida have been classified as Inferred where drill spacing is wider spaced or where unresolved geological complexity exists. Figure 76 shows the resource classification applied to the Escondida epithermal main zones. A summary of the Escondida Indicated and Inferred mineral resources above a cut-off of 1.0 ppm gold equivalent are shown in Table 44 and Table 45 respectively.

Figure 76 Escondida Main Zones – Resource Classification

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
1101 – Central-East	167,000	21.9	495.1	31.8	118,000	2,662,000	171,000
1201- West 2	96,000	16.7	686.8	30.4	52,000	2,118,000	94,000
1301- Far West	344,000	17.3	1018.4	37.7	192,000	11,270,000	417,000
1401- West 1	13,000	33.1	1177.5	56.7	14,000	480,000	23,000
Total Esc Ind.	620,000	18.8	829.2	35.4	374,000	16,530,000	705,000

Table 44 Indicated Escondida Resources above 1 ppm Gold Equivalent

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Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
1101 – Central-East	114,000	4.1	96.8	6.0	15,000	356,000	22,000
1201- West 2	69,000	7.7	400.5	15.7	17,000	887,000	35,000
1301- Far West	318,000	3.4	128.0	6.0	35,000	1,309,000	61,000
1401- West 1	6,000	14.0	679.5	27.6	3,000	137,000	6,000
Total Esc Inf.	508,000	4.3	164.8	7.6	70,000	2,689,000	123,000

Table 45 Inferred Escondida Resources above 1 ppm Gold Equivalent

*Gold equivalent values have been calculated by Extorre on the basis of the following parameters:

	●	Long term gold price US$1,320/oz
	●	Long term silver price US$26/oz
	●	Metallurgical recovery gold 100%
	●	Metallurgical recovery silver 100%

The gold equivalent value is calculated by dividing the silver grade (ppm) by 50 (approximate ratio of gold/silver US$ value) and adding it to the gold grade (ppm).

14.6.2

Loma Escondida

Data Types

The Loma Escondida estimate was based on Extorre diamond holes and trenches. Trench sampling at Loma Escondida is considered to be of high quality and trench grades have been incorporated in the estimation. The total Loma Escondida dataset comprises 57 diamond drill holes (57 Extorre 3588.55m), 8 RC drill holes (6 Extorre 432m and 2 Mincorp 124m) and 42 Extorre trenches. No RC holes intersected the Loma Escondida mineralised structure.

Domaining and Volume Modelling

The main potentially economic mineralised zone within the Loma Escondida prospect area consists of a single steeply dipping E-W vertically extensive narrow (0.2-2.3m) epithermal quartz vein structure. High grade gold and silver mineralisation primarily occurs in multi-phase banded quartz-adularia and quartz-chlorite-sulphide veins, hydrothermal breccias and peripheral stockworks. The domain outlines used to control volume and estimations have been predominantly based on geological attributes and observations from drill core, particularly epithermal vein textures rather than grade criteria. In most cases, a clear distinction between a main epithermal quartz vein structure and surrounding host rock can be determined based on detailed geological logging and core photography (Figure 77). In addition, a number of surface trenches and sub-outcrop (Figure 78) have been used to guide the interpretation. There does not appear to be a well developed stockwork envelope surrounding the

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Loma Escondida main vein however a zone of slightly increased alteration around the main vein has been interpreted to provide a dilution estimate.

Figure 77 Loma Escondida Drillhole MD195 Showing Styles of Mineralisation

Figure 78 Loma Escondida Central – Quartz Float and Sub-Outcrop

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The Loma Escondida main epithermal vein structure is characterised by distinct shoot like ‘bonanza’ zones occurring wholly within the plane of the main epithermal vein structure, effectively creating two mineralisation styles:

●

Main Zone (MZ) – continuous material characterised by classic epithermal vein textures including crustiform/colloform chalky white quartz-adularia sulphide banded and brecciated veining - typically grading 0.5 to 10 Au ppm;

●

Bonanza Zone (BZ) – semi-continuous identifiable zones within the Main Zone. Characterised by brecciated quartz-adularia sulphide banded vein ‘ginguro’ material and often associated with base metal sulphides and black silica - typically grading 10 to 80 Au ppm.

The Loma Escondida BZ shoots appear to have some similarities to those identified within the Escondida prospect.

The domaining and volume modelling approach is twofold:

Main Zone (MZ): Cross-sections were generated in accordance with the drill hole spacing which typically ranges between 20m-40m spacing. Sectional interpretations of the main epithermal vein structure using geological logging and core photography was completed over the full strike extent of the Loma Escondida prospect. The resulting MZ mineralisation interpretation was graphically digitised and a 3D wireframe model produced. The interpretation of the MZ is based on geological attributes such as epithermal vein textures ensuring that the final 3D mineralisation model reflects an in-situ geological model whereby no cut-off grade or minimum mining width criteria has been applied. Figure 79 and Figure 80 show the Loma Escondida main epithermal vein zone together with drilling locations and surface topography.

Alteration Zone: A surrounding alteration interpretation was generated in a similar manner to the MZ domains and a 3D wireframe model produced. Whilst the alteration domain was predominantly based on geological logging and core photography, anomalous gold and silver grades outside the MZ were also used as a guide. The alteration domain typically extends 1-3m from the MZ hangingwall and footwall contacts.

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Figure 79 Loma Escondida Main Epithermal Vein Zone – Plan View

Figure 80 Loma Escondida Main Epithermal Vein Zone – Looking North

Database Coding and Compositing

Samples within the MZ zones and surrounding alteration zone are assigned a unique database code which is used to control the compositing process. A four digit numbering system (zonecode) was created to define the various domains within the Loma Escondida prospect (Table 46).

Zonecode	Domain
2101	Loma Escondida Main Zone
2901	Loma Escondida Alteration Zone

Table 46 Loma Escondida Domain Numbering

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Main Zone (MZ+BZ)

As with Escondida, there are several key physical features of the Loma Escondida epithermal vein zone that need to be considered and accommodated by the selected resource modelling technique. These features are:

	●	Mineralised epithermal vein zone true thickness is variable typically ranging from 0.2m to 2.3m in width.
	●	Undulating or variable zone geometry (dip and strike) and possible grade/thickness trends within this variable geometry;
	●	Sampling has been taken over geological intervals creating samples of unequal length or variable support;
	●	The continuity and geometry of the ‘Bonanza Zone’ (BZ) within the ‘Main Zone’ (MZ) is quite variable making it difficult to apply separate deterministic domaining to the BZ;
	●	Mining selectivity across the epithermal vein zones is unlikely due the narrow nature of the mineralised structures;
	●	Drill spacing is variable but typically around 20mN by 20mE to 40mN by 40mE within the plane of mineralisation.

Alteration Zone

Descriptive Statistics and High Grade Capping

Main Zone (MZ)

Figure 81 to Figure 85 show the spatial distribution of raw intercept grades, horizontal widths and accumulations for the Loma Escondida MZ domains. It can be clearly seen from Figure 81 and

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Figure 82 that high grade gold and silver values form semi-continuous shoot-like occurrences within the plane of the main vein. In addition, Figure 83 indicates that areas of elevated gold and silver grades appear to be coincident with increased vein width albeit this relationship is more subtle than for Escondida. This positive association between grade and width is common in epithermal vein systems and often reflects structural features such as dilation zones or flexures in zone geometry. Not surprisingly, the association between grade and width is particularly evident when visualising the accumulation values as shown in Figure 84 and Figure 85.

Figure 81 Loma Escondida Main Zone – Geological Composites – Au ppm - Long Section

Figure 82 Loma Escondida Main Zone– Geological Composites – Ag ppm – Long Section

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Figure 83 Loma Escondida Main Zone – Geological Composites – Horiz. Width - Long Section

Figure 84 Loma Escondida Main Zone – Geological Composites – Au*HW (gxm) – Long Section

Figure 85 Loma Escondida Main Zone – Geological Composites – Ag*HW (gxm) – Long Section

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An examination of gold and silver geological composite statistics over the coded intervals was undertaken prior to calculating the accumulation to identify any outlier composites that may require high grade capping. It was decided to apply a modest high grade assay cut of 100 Au ppm and 4000 Ag ppm to gold and silver respectively. Whilst the application of high grade capping is somewhat arbitrary it is considered by the Author to be a prudent and necessary risk adjustment strategy at this stage of project evaluation. Figure 86 to Figure 90 show log-probability plots of raw intercept grades, horizontal width and accumulationsfor the Loma Escondida MZ domain.

Figure 86 Loma Escondida Main Zone – Geological Composites – Au ppm – Log-Probability Plot

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Figure 87 Loma Escondida Main Zone – Geological Composites – Ag ppm – Log-Probability Plot

Figure 88 Loma Escondida Main Zone – Geological Composites – Horiz. Width – Log-Probability Plot

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Figure 89 Loma Escondida Main Zone – Geological Composites – Au*HW (gxm) – Log-Probability Plot

Figure 90 Loma Escondida Main Zone – Geological Composites – Ag*HW (gxm) – Log-Probability Plot

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A list of Loma Escondida geological intercept composites is included in Section 14.8.

Table 47 details the high grade caps applied to geological composite data prior to calculating the accumulation variables for the Loma Escondida prospect.

Domain	Zonecode	Gold Cap ppm	Number Capped	Percentile	Silver Cap ppm	Number Capped	Percentile
Loma Escondida Main Zone	2101	100	3	97th	4000	6	93th

Table 47 Loma Escondida High Grade Caps

As with Escondida, there is greater drilling density (by design) in areas of elevated grade compared to the lower grade peripheral areas. As such, declustering will be required in order to compute robust global statistics. By way of illustrating this point, Figure 91 shows a comparison between raw intercept gold grade and declustered intercept gold grade using various declustering cell dimensions for the Loma Escondida MZ data set. It is clearly shown that the grade decreases as the declustering cell size becomes larger reflecting a weighting toward wider drill spacing in lower grade areas. It could be reasonably expected that an unbiased global grade estimate of the Loma Escondida MZ domains would be roughly equivalent to a 45m x 45m declustered intercept composite grade based on the overall drill spacing. A comparison such as this forms a key component of model validation at a later stage.

Figure 91 Loma Escondida Main Zone – Declustered Grade Comparison – Au ppm

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Table 48 and Table 49 summarises Loma Escondida MZ raw geological intercept grades (cut and uncut) and accumulation (cut only) statistics respectively. It can be seen from Table 48 that declustering has a significant impact on the average intercept grade reducing the global gold and silver grade by around 18%. It is also important to note that the width weighted declustered mean grades (highlighted in red) in Table 49 are higher grade (particularly for gold) than the raw declustered mean grades (highlighted in blue) in Table 48. This confirms statistically that there is a positive association between elevated grade and zone width within the Loma Escondida MZ domains. Furthermore, the association between grade and width necessitates the application of an accumulation style method of estimation for the Loma Escondida epithermal main vein domains.

	Raw Au ppm	Raw Ag ppm	Cut Au ppm	Cut Ag ppm
Number	75	75	75	75
Minimum	0.08	0.53	0.08	0.53
Maximum	145.50	5963.00	100.00	4000.00
Raw Mean	18.02	887.95	17.06	804.72
Declust. Mean 45x45	14.73	738.67	14.02	652.25
Std Dev	30.12	1456.49	26.72	1214.08
Coeff Var	1.672	1.64	1.566	1.509

Table 48 Loma Escondida Main Zone - Summary Statistics – Geological Composites - Raw Grades

	Horizontal Width	Cut Au X HW	Cut Ag X HW	HWidth Weighted Cut Au ppm	HWidth Weighted Cut Ag ppm
Number	75	75	75
Minimum	0.17	0.02	0.09
Maximum	2.60	100.00	4000.00
Raw Mean	0.70	11.644	527.503	16.59	751.43
Declust. Mean 45x45	0.66	9.84	430.50	15.00	656.25
Std Dev	0.46	20.58	881.90
Coeff Var	0.65	1.77	1.67

Table 49 Loma Escondida Main Zone - Summary Statistics – Geological Composites - Accumulations

Alteration Zone

Log-probability plots of gold and silver 1m composites for the Loma Escondida alteration domains are shown in Figure 92 and Figure 93 respectively. An assay top cut of 3 Au ppm and 200 Ag ppm was applied to gold and silver respectively as a model risk adjustment strategy. Table 50 shows summary statistics of raw and cut gold and silver 1m composites for the Loma Escondida alteration zone.

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	Raw Au ppm	Raw Ag ppm	Cut Au ppm	Cut Ag ppm
Number	284	284	284	284
Minimum	0.01	0.15	0.01	0.15
Maximum	4.66	166.06	3.00	100.00
Raw Mean	0.18	10.96	0.17	10.37
Std Dev	0.45	21.44	0.38	17.93
Coeff Var	2.55	1.96	2.26	1.73

Table 50 Loma Escondida Alteration Zones - Summary Statistics – 1m Downhole Composites

Figure 92 Loma Escondida Alteration Zones – 1m Downhole Composites – Au ppm – Log-Probability Plot

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Figure 93 Loma Escondida Alteration Zones – 1m Downhole Composites – Ag ppm – Log-Probability Plot

Domaining of Bonanza Shoots

As mentioned in Section 14.6.2 “Domaining and Volume Modelling” the Loma Escondida main epithermal vein zone hosts distinct shoot like ‘bonanza’ zones occurring wholly within the plane of the main epithermal vein structure. While the BZ zones appear to exhibit sharp grade boundaries, the continuity and geometry of these zones is quite variable making it difficult to apply separate deterministic domaining. Nevertheless, it is the Author’s opinion that the BZ zones require separate domaining for the purposes of grade estimation to avoid potential over-smoothing of the bonanza grades.

Cube adopted a similar Indicator Simulation (Sequential Indicator Simulation - SIS) approach as described for the Escondida prospect for the purposes of defining the BZ shoots within the main vein zone. Exploratory data analysis showed that a gold grade indicator of 16 Au ppm resulted in excellent definition of the BZ shoots effectively creating a grade based separation of BZ and surrounding MZ. The MZ domain was further split into a low grade and moderate grade domain using a gold grade indicator of 2 Au ppm. Figure 94 and Figure 95 and show a statistical and spatial representation of the indicator thresholds applied for the Indicator Simulation.

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Figure 94 Loma Escondida Main Zone – Gold Indicator Threshold – Log-Probability Plot

Figure 95 Loma Escondida Main Zone –Geological Composites – Au ppm – Indicators - Long Section

Figure 95 clearly shows that the spatial distribution of BZ is not evenly distributed throughout the project area. It was determined that the results of the SIS could be improved by estimating a localised BZ and MZ probability at each location. This was achieved by interpolating the BZ and MZ indicators using Ordinary Kriging into each simulation grid node. This process effectively creates a 2D probability map of each indicator similar to that shown for the Loma Escondida BZ in Figure 96. An omni-directional variogram model based on the BZ indicator threshold was fitted for the Loma Escondida main zone dataset and used as the basis for the SIS (Figure 97).

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Figure 96 Loma Escondida Main Zone – BZ Indicator Probability Map

Figure 97 Loma Escondida Main Zone – BZ Indicator Variogram Model

The SIS process incorporates a Simple Kriging step which uses these localised probabilities as a local mean during simulation. A sequential neighbourhood was defined whereby a maximum of 16 initial data points and 10 already simulated nodes were used to condition the simulation. A set of 100 simulation realisations of the BZ and MZ indicators was performed into 2D 1m x 1m grid nodes. The use of SIS allows the generation of a series of equally probable geological models whereby a statistically derived ‘most probable’ geological/mineralisation model can be determined. Figure 98 shows examples of several simulated domain models together with the statistically derived ‘most probable’ geological model for the Loma Escondida Main Zone. The ‘most probable’ geological model was re-grouped into X=10m x Z=10m blocks whereby each larger block has an estimated proportion of each mineralisation style.

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Figure 98 Loma Escondida Main Zone – Simulated Domains

Variography

Geological composites for individual main zone domains were separated on the basis of the gold grade indicator. Variography was used to characterise the spatial continuity of the horizontal width and accumulation variables for each grade indicator within the plane of each vein structure. Experimental variography for the BZ was poorly structured due to the relatively few informing data points informing. It was decided to combine all BZ and MZ data to improve the quality of variography for these indicators. No obvious direction of preferred continuity was determined and omni-directional variogram models were fitted for all variables. Variograms of horizontal width and accumulations demonstrated similar spatial structure and variance proportions. Given the similarities between these variograms a simplification was made whereby a single variogram model was used to estimate both the accumulation and thickness variables. This approach has the advantage that it is simpler and ensures no instability in the back-calculated block grades due to unstable or unexpected width estimations which may occur when a different variogram model is used. Experimental variography was undertaken on Gaussian transformed data. Variogram models were fitted to the Gaussian data and subsequently back-transformed into real variances to obtain the appropriate parameters for interpolation of the raw accumulation variables by Ordinary Kriging. Fitted Gaussian based variogram models and back-transformed variogram model parameters for the combined Loma Escondida ‘Bonanza Zones’ (BZ) and ‘Main Zones’ (MZ) are shown in Figure 99 and Table 51 respectively.

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Figure 99 Loma Escondida Combined BZ and MZ –Variogram Models – Gaussian Au*HW and Ag*HW

Sill
(Relative
Variance)

Range

Azimuth

Plunge

Dip

Major/ Semi
Major Ratio

Major/

Minor Ratio

Au*HW	Nugget Co	0.44
	Structure 1	0.14	30	0	0	0	1	1
	Structure 2	0.42	85	0	0	0	1	1
Ag*HW	Nugget Co	0.39
	Structure 1	0.17	30	0	0	0	1	1
	Structure 2	0.44	85	0	0	0	1	1

Table 51 Loma Escondida Combined BZ and MZ – Back-Transformed Raw Au*HW and Ag*HW

The combined BZ and MZ data resulted in well structured experimental variography for gold and silver accumulations with both exhibiting moderate relative nugget effects of 44% and 39% respectively with maximum ranges in the order of 85m.

Alteration Zone

Experimental variography on the Loma Escondida alteration zone resulted in very poorly structured variograms. It was decided to apply the Escondida stockwork variography for the purpose of the Loma Escondida alteration estimates. Variogram model parameters for the Loma Escondida alteration zones are shown in Table 52.

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		Sill (Relative Variance)	Range	Azimuth	Plunge	Dip	Major/ Semi Major Ratio	Major/ Minor Ratio
Au	Nugget Co	0.23
	Structure 1	0.27	3.8	306	0	85	1	3.8
	Structure 2	0.23	20	306	0	85	1	4.0
	Structure 3	0.27	60	306	0	85	1	6.0
Ag	Nugget Co	0.29
	Structure 1	0.22	3.8	306	0	85	1	3.8
	Structure 2	0.39	20	306	0	85	1	4.0
	Structure 3	0.10	60	306	0	85	1	6.0

Table 52 Loma Escondida Alteration Zone – Variogram Models – Raw Au ppm and Ag ppm

Grade Estimation

Main Zone (MZ+BZ)

Estimation of block grades within the main zone domains used an accumulation methodology whereby the gold and silver accumulations and horizontal width were estimated on a 2D vertical plane into X=10m x Z=10m blocks using Ordinary Kriging for each grade indicator. Final block grade is calculated by dividing the estimated accumulation by the estimated thickness and weighted by the proportion of each mineralisation style (BZ and MZ) computed previously by the SIS approach. The estimation of horizontal width used for block grade back-calculation used the same variogram models as for the accumulations. This simplification ensures no instability in the back calculated grades due to unstable or unexpected width estimations which may occur when a different variogram is used. The adopted estimation methodology resulted in very good reproduction of the ‘bonanza’ shoots within the main vein zones.

Kriging neighbourhood analysis was undertaken to determine appropriate inputs to the Ordinary Kriging process. Estimation parameters for the BZ and MZ domains are summarised in Table 53.

Parameter	CutAuXHW	CutAgXHW	HWidth
Minimum number of Comps	2	2	2
Maximum number of Comps	20	20	20
Search Major Distance	150	150	150
Search Orientation	000	000	000
Plunge of Major Axis	0	0	0
Dip of Major Axis	0	0	0
Anisotropy major/semi-major	1.0	1.0	1.0
Anisotropy major/minor	1.0	1.0	1.0
Block Discretization	10 x 10	10 x 10	10 x 10

Table 53 Loma Escondida (BZ + MZ) – Estimation Parameters

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3DBlock Model Definition

At the completion of 2D block grade back-calculation each of the X=10m x Z=10m parent cells was re-located into real-world space and imported into a 3D block model. The primary consideration of the 3D models was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each mineralised zone. All individual mineralised zones were ultimately combined to create a single block model in the Gauss-Krüger coordinate system. The Loma Escondida block model definition is shown in Table 54. A standard list of field names and descriptions used in the block model are shown in Table 55.

Origin	Minimum	Maximum	Model Extent
Y	4,667,500	4,669,500	2,000
X	2,673,000	2,676,500	3,500
Z	-500	200	700
Parent Cell Y	2m	Min Sub-Cell Y m	0.25
Parent Cell X	10m	Min Sub-Cell X m	1.25
Parent Cell Z	10m	Min Sub-Cell Z m	1.25

Table 54 Loma Escondida 3D Block Model Definition

Field Name	Description
x	X Block Centroid
y	Y Block Centroid
z	Z Block Centroid
au	Au ppm – Interpolated
ag	Ag ppm – Interpolated
aueq60	Gold Equivalent ppm = Au + Ag/60
density	Density g/cm3 – Direct Assignment – Default 2.5
rescode	Resource Classification - 1=Measured 2=Indicated 3=Inferred
zone	Domain Code Eg. 2101, 2901

Table 55 Loma Escondida Block Model Attribute Names

Bulk Density

Global bulk densities were assigned as follows:

Main Epithermal Vein Zone - 2.65 g/cm3 assumed - based on 58 on-site density determinations for the Escondida main epithermal vein zone.

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Oxidation

Oxidation of mineralised domains is observed to be of insignificant depth and does not appear to materially influence the bulk density or characteristics of the main mineralised zones. Whilst trenching and sub-outcrop indicates that the Loma Escondida main zone is very close to the surface in places, Cube has assigned a zero grade to all material within 1m below the topographical surface.

Final Model

The final Surpac block model was clipped by the topography to remove volume above the current surface and exported into CSV format. Due to its proximity to the Escondida prospect, the Loma Escondida model was incorporated into the same model as provided for Escondida. The following Loma Escondida model in the Gauss-Krüger coordinate system was provided to Extorre:

esc_mar2011_min_zones.mdl – Surpac format block model with sub-blocking to blocks of Y=0.25m, X=1.25m and Z=1.25m

esc_mar2011_min_zones_07042011.csv – CSV export of above Surpac model

esc_mar2011_min_zones_reg_07042011.csv – CSV export of above Surpac model regularised to blocks of Y=1m, X=1.25m and Z=2.5m

esc_mar2011_ppct_min_zones_29052011.csv – CSV export of above Surpac model zone partial percentages within blocks of Y=2m, X=10m and Z=10m.

Model Validation

Model validation was carried out visually and by comparing the modelled outcomes against de-clustered composite grades. Figure 100 and Figure 101 show back-calculated model grades for gold and silver together with geological intercept composites for the Loma Escondida epithermal main zone (MZ+BZ). The model outcomes show very good correspondence to the informing composite, in particular, providing very good definition of the bonanza shoots. Statistical comparisons of composite grades (raw and declustered width weighted) with the modelled outcomes are shown in Table 56 for the Loma Escondida epithermal main zone. Whilst the statistical comparisons indicate that a slightly higher modelled grade could be expected, the visual comparison confirms that a robust estimate has been achieved. The statistical comparisons should only be considered as an 'order of magnitude' guide given the paucity of available data.

Cube concludes that both visual and statistical comparisons confirm that a robust and unbiased model outcome has been achieved.

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Figure 100 Loma Escondida Main Zone (MZ+BZ) – Composites vs Model Grades – Au ppm

Figure 101 Loma Escondida Main Zone (MZ+BZ) – Composites vs Model Grades – Ag ppm

Zone	De-clustering Cell Size	Raw Composite Grades Au ppm	De-clustered Composite Grade Au ppm (HW wt)	Model Grade Au ppm	Raw Composite Grades Ag ppm	De-clustered Composite Grade Ag ppm (HW wt)	Model Grade Ag ppm
2101	X=45 x Z=45	18.02	15.00	14.29	804.25	656.25	735.12

Table 56 Loma Escondida Combined Main Zone - De-Clustered Composite vs Model Grade

Resource Classification and Reporting

Cube has classified a substantial proportion of the Loma Escondida mineral resources as Indicated where drilling is 20m x 20m or closer. Cube believes that this level of drill hole spacing is sufficient to demonstrate a high level of confidence in the geometry, continuity and grade of the Loma Escondida deposit. Surrounding areas of Loma Escondida have been classified as Inferred where drill spacing is wider spaced or where unresolved geological complexity exists. Figure 102 shows the resource classification applied to the Loma Escondida epithermal main zone. A summary of the Loma Escondida Indicated and Inferred mineral resources above a cut-off of 1.0 ppm gold equivalent are shown in Table 57 and Table 58 respectively.

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Figure 102 Loma Escondida Main Zone – Resource Classification

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
2101	44,000	18.4	919.5	36.8	26,000	1,297,000	52,000
Total Loma Esc Ind.	44,000	18.4	919.5	36.8	26,000	1,297,000	52,000

Table 57 Indicated Loma Escondida Resources above 1 ppm Gold Equivalent

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
2101	13,000	9.7	595.4	21.6	4,000	256,000	9,000
Total Loma Esc Inf.	13,000	9.7	595.4	21.6	4,000	256,000	9,000

Table 58 Inferred Loma Escondida Resources above 1 ppm Gold Equivalent

* Gold equivalent values have been calculated by Extorre on the basis of the following parameters:

	●	Long term gold price US$1,320/oz
	●	Long term silver price US$26/oz
	●	Metallurgical recovery gold 100%
	●	Metallurgical recovery silver 100%

The gold equivalent value is calculated by dividing the silver grade (ppm) by 50 (approximate ratio of gold/silver US$ value) and adding it to the gold grade (ppm).

14.6.3

Gabriela

Data Types

The Gabriela estimate was based on a mixture of Extorre diamond drilling and reverse circulation drill holes. The total Gabriela dataset comprises 127 diamond drill holes (Extorre 19,603.65m), 14 RC drill holes (Extorre 1,239m) and 6 trenches (Extorre Dec 2010).

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Domaining and Volume Modelling

The main economic mineralised zones within the Gabriela prospect area consist of a series of steeply dipping E-W to ESE vertically extensive narrow (0.1-4m) epithermal quartz vein structures. Economic gold and silver mineralisation primarily occurs in multi-phase banded quartz-adularia and quartz-chlorite-sulphide veins, hydrothermal breccias and peripheral stockworks. The domain outlines used to control volume and estimations have been predominantly based on geological attributes and observations from drill core, particularly epithermal vein textures rather than grade criteria. In most cases, a clear distinction between a main epithermal quartz vein structure and surrounding stockwork mineralisation can be determined based on detailed geological logging and core photography (Figure 103). In addition, a number of surface trenches and sub-outcrop have been used to guide the interpretation (Figure 104). Surrounding stockwork are characterised by stockwork veining of varying intensity with minor breccia zones, accompanied by elevated silver grades.

Figure 103 Gabriela Drillhole MD371 Showing Styles of Mineralisation

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Figure 104 Gabriela – Quartz Float and Sub-Outcrop

The domaining and volume modelling approach was twofold:

Vein Zones: Cross-sections were generated in accordance with the drill hole spacing which typically range between 20m-80m spacing. Sectional interpretations of the main epithermal vein structure (MZ) using geological logging and core photography was completed over the full strike extent of each prospect. The resulting mineralisation interpretations were graphically digitised and a 3D wireframe model produced. The interpretation of each MZ was based on geological attributes such epithermal vein textures ensuring that the final 3D mineralisation model reflects an in-situ geological model whereby no cut-off grade or minimum mining width criteria has been applied.

Stockwork Zones: A surrounding stockwork interpretation was generated in a similar manner to the vein domains and a 3D wireframe model produced. Whilst the stockwork domain was predominantly based on geological logging and core photography, anomalous gold and silver grades outside the MZ were also used as a guide. The stockwork domain typically extends 2-5m from the vein hangingwall and footwall contacts and encompasses stockwork veining of varying intensity and minor breccia zones.

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Figure 105 and Figure 106 show the Gabriela main epithermal vein zone together with drilling locations and surface topography.

Figure 105 Gabriela Main Epithermal Vein Zone – Plan View

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Figure 106 Gabriela Main Epithermal Vein Zone – Looking North

Database Coding and Compositing

Zonecode	Domain
3101	Gabriela Main Zone
3901	Gabriela Stockwork Zone

Table 59 Gabriela Domain Numbering

Main Zones (Gabriela 3101)

There are several key physical features of the Gabriela epithermal vein zones that need to be considered and accommodated by the selected resource modelling technique. These features are:

	●	Mineralised epithermal vein zone true thickness is variable typically ranging from 0.1m to 4m in width.
	●	Undulating or variable zone geometry (dip and strike) and possible grade/thickness trends within this variable geometry;
	●	Sampling has been taken over geological intervals creating samples of unequal length or variable support;
	●	Mining selectivity across the epithermal vein zones is unlikely due the narrow nature of the mineralised structures;

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●

Drill spacing is variable but typically around 25m by 80m or larger within the plane of mineralisation.

A method that allows estimation of metal content on a projected plane is considered to be an appropriate approach to addressing the features outlined above. This resource modelling approach is best achieved using geological intercept composites and accumulation estimation. MZ grades are composited across the entire coded interval resulting in a single geological intercept composite at each intercept location. The estimation approach for the Gabriela prospect differs slightly from the Escondida methodology whereby the accumulation and width variables were estimated into 3D blocks orientated parallel to the vein orientation in Local Grid coordinates. In this orientation a projection to a 2D plane was not considered necessary as each MZ is aligned by design, closely to Local Grid East-West (see Figure 105). The mid-point of each geological composite is assigned the horizontal width of the vein structure and used to compute a ‘metal accumulation’ variable.

a(x) = t(x) . z(x)

Stockworks (Gabriela 3901)

Traditional downhole composites of 1.5m were generated for the surrounding stockwork domains whereby the zonecode flagging was used to control compositing. The downhole composites were extracted using a ‘best fit’ method which optimises the composite length over the coded interval to eliminate the problem of residuals.

Descriptive Statistics and High Grade Capping

Main Zone (MZ)

Figure 107 to Figure 111 show the spatial distribution of raw intercept grades, horizontal widths and accumulations for the Gabriela MZ domains. It can be clearly seen from Figure 107 and Figure 108 that high grade gold and silver values form semi-continuous zones within the plane of the main vein. In addition, Figure 109 indicates that areas of elevated gold and silver grades appear to be coincident with increased vein width. This positive association between grade and width is common in epithermal vein systems and often reflects structural features such as dilation zones or flexures in zone geometry. Not surprisingly, the association between grade and width is particularly evident when visualising the accumulation values as shown in Figure 110 and Figure 111.

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Figure 107 Gabriela Main Zone – Geological Composites – Au ppm - Long Section

Figure 108 GabrielaMain Zone – Geological Composites – Ag ppm – Long Section

Figure 109 GabrielaMain Zone – Geological Composites – Horiz. Width - Long Section

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Figure 110 GabrielaMain Zone – Geological Composites – Au*HW (gxm) – Long Section

Figure 111 GabrielaMain Zone – Geological Composites – Ag*HW (gxm) – Long Section

Figure 112 to Figure 116 show log-probability plots of raw intercept grades, horizontal widths and accumulations for the Gabriela MZ domain. Unlike the Escondida MZ, the style of mineralisation at Gabriela does not contain 'bonanza' grade samples. An examination of gold and silver geological composite statistics over the coded intervals was undertaken prior to calculating the accumulation to identify any outlier composites that may require high grade capping. It was decided to apply a high grade assay cut of 7 Au ppm and 1200 Ag ppm to gold and silver respectively.

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Figure 112 Gabriela Main Zone – Geological Composites – Au ppm – Log-Probability Plot

Figure 113 Gabriela Main Zone – Geological Composites – Ag ppm – Log-Probability Plot

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Figure 114 Gabriela Main Zone – Geological Composites – Horiz. Width – Log-Probability Plot

Figure 115 Gabriela Main Zone – Geological Composites – Au*HW (gxm) – Log-Probability Plot

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Figure 116 Gabriela Main Zone – Geological Composites – Ag*HW (gxm) – Log-Probability Plot

A list of Gabriela geological intercept composites is included in Section 14.8.

Table 60 details the high grade caps applied to geological composite data prior to calculating the accumulation variables for the Gabriela prospect.

Domain	Zonecode	Gold Cap ppm	Number Capped	Percentile	Silver Cap ppm	Number Capped	Percentile
Gabriela Main Zone	3101	7	4	97th	1200	4	97th

Table 60 Gabriela High Grade Caps

Drill spacing at Gabriela ranges from 20m x 20m to 80m x 80m. Unlike Escondida, there does not appear to be a bias in drilling density within the higher grade areas of the deposit. The western two thirds of the Gabriela prospect has been systematically drilled on a 20m x 20m or 20m x 40m pattern whilst the eastern third of the prospect is currently drilled on a 40m x 40m to 80m x 80m pattern. Given the variable drill spacing at Gabriela it is difficult to obtain robust declustered statistics. Figure 17 shows a comparison between raw intercept gold grade and declustered intercept gold grade using various declustering cell dimensions for the Gabriela MZ data set. This comparison clearly shows some instability of the declustered results related to the variable drill spacing.

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Figure 117 Gabriela Main Zone – Declustered Grade Comparison – Au ppm

Table 61 and Table 62 summarises Gabriela MZ raw geological intercept grades (cut and uncut) and accumulation (cut only) statistics respectively. Declustering statistics have not been presented in these tables due to the instability in results as highlighted above. It is also important to note that the width weighted mean grades (highlighted in red) in Table 62 are higher grade (particularly for gold) than the raw mean grades (highlighted in blue) in Table 61. This confirms statistically that there is a positive association between elevated grade and zone width within the Gabriela MZ domains. Furthermore, the association between grade and width necessitates the application of an accumulation style method of estimation for the Gabriela epithermal main vein domains.

	Raw Au ppm	Raw Ag ppm	Cut Au ppm	Cut Ag ppm
Number	113	113	113	113
Minimum	0.03	2.62	0.03	2.62
Maximum	17.36	2222.54	7.00	1200.00
Raw Mean	2.25	342.30	2.11	328.81
Std Dev	2.32	330.15	1.70	274.00
Coeff Var	1.03	0.96	0.81	0.83

Table 61 Gabriela Main Zone - Summary Statistics – Geological Composites - Raw Grades

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	Horizontal Width	Cut Au X HW	Cut Ag X HW	HWidth Weighted Cut Au ppm	HWidth Weighted Cut Ag ppm
Number	113	113	113
Minimum	0.26	0.02	2.41
Maximum	8.54	26.88	4608.00
Raw Mean	1.78	4.31	656.25	2.42	368.68
Std Dev	1.39	4.90	793.10
Coeff Var	0.78	1.14	1.21

Table 62 Gabriela Main Zone - Summary Statistics – Geological Composites - Accumulations

Stockwork Zone

Log-probability plots of gold and silver 1.5m composites for the Gabriela stockwork zones are shown in Figure 118 and Figure 119 respectively. An assay top cut of 2 Au ppm and 200 Ag ppm was applied to gold and silver respectively as a model risk adjustment strategy. Table 63 shows summary statistics of raw and cut gold and silver 1.5m composites for the Gabriela stockwork zone.

	Raw Au ppm	Raw Ag ppm	Cut Au ppm	Cut Ag ppm
Number	1273	1273	1273	1273
Minimum	0.00	0.15	0.00	0.15
Maximum	7.99	725.29	2.00	200.00
Raw Mean	0.18	24.83	0.17	23.38
Std Dev	0.40	47.35	0.28	36.20
Coeff Var	2.25	1.91	1.68	1.55

Table 63 Gabriela Stockwork Zones - Summary Statistics – 1m Downhole Composites

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Figure 118 Gabriela Stockwork Zones – 1m Downhole Composites – Au ppm – Log-Probability Plot

Figure 119 Gabriela Stockwork Zones – 1m Downhole Composites – Ag ppm – Log-Probability Plot

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Domaining

The Gabriela epithermal main vein differs from Escondida and Loma Escondida in that there are no well developed 'bonanza' grade shoots evident. Whilst high grade shoots do exist, the absence of 'bonanza' grades reduces the necessity to apply separate domains within the main vein. It was determined that the entire Gabriela main vein could be treated as a single domain for the purpose of grade estimation.

Variography

Variograms of horizontal width and accumulations demonstrated similar spatial structure and variance proportions. Given the similarities between these variograms a simplification was made whereby a single variogram model was used to estimate both the accumulation and thickness variables. This approach has the advantage that it is simpler and ensures no instability in the back-calculated block grades due to unstable or unexpected width estimations which may occur when a different variogram model is used. Experimental variography was undertaken on Gaussian transformed data. Variogram models were fitted to the Gaussian data and subsequently back-transformed into real variances to obtain the appropriate parameters for interpolation of the raw accumulation variables by Ordinary Kriging.

Fitted Gaussian based variogram models and back-transformed variogram model parameters for the Gabriela Main Zone (MZ) are shown in Figure 120 and Table 64 respectively.

Figure 120 Gabriela Main Zone – Variogram Models – Gaussian Au*HW and Ag*HW

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		Sill (Relative Variance)	Range	Azimuth	Plunge	Dip	Major/ Semi Major Ratio	Major/ Minor Ratio
Au*HW	Nugget Co	0.11
	Structure 1	0.46	48	0	0	0	1	1
	Structure 2	0.43	200	0	0	0	1	1
Ag*HW	Nugget Co	0.11
	Structure 1	0.46	48	0	0	0	1	1
	Structure 2	0.43	200	0	0	0	1	1

Table 64 Gabriela Main Zone – Back-Transformed Raw Au*HW and Ag*HW

Stockwork Zone

3D experimental variography was undertaken on cut Gaussian transformed 1.5m down hole composites. Variogram models were fitted to the Gaussian data and subsequently back-transformed into real variances to obtain the appropriate parameters for interpolation of the raw accumulation variables by Ordinary Kriging. Experimental variography on the Gabriela stockwork zone resulted in well structured variograms. Back-transformed variogram model parameters for the Gabriela stockwork zones are shown in Table 65.

		Sill (Relative Variance)	Range	Azimuth	Plunge	Dip	Major/ Semi Major Ratio	Major/ Minor Ratio
Au	Nugget Co	0.55
	Structure 1	0.20	10	0	0	0	1	1
	Structure 2	0.25	160	0	0	0	1	1
Ag	Nugget Co	0.55
	Structure 1	0.20	10	0	0	0	1	1
	Structure 2	0.25	160	0	0	0	1	1

Table 65 Gabriela Stockwork Zone – Variogram Models – Raw Au ppm and Ag ppm

Grade Estimation

Main Zone (MZ)

The estimation approach for the Gabriela differs slightly from the Escondida methodology whereby the accumulation and horizontal width variables were estimated by 3D Ordinary Kriging into X=10m x Y=2m x Z=10m blocks orientated parallel to the vein orientation in Local Grid coordinates. In this orientation a projection to a 2D plane was not considered necessary as each MZ is aligned by design, closely to Local Grid East-West. Final block grade is calculated by dividing the estimated accumulation by the estimated thickness. The estimation of horizontal width used for block grade back-calculation used the same variogram models as for the accumulations. This simplification ensures no instability in the back calculated grades due to unstable or unexpected width estimations which may occur when a different variogram is used.

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Kriging neighbourhood analysis was undertaken to determine appropriate inputs to the Ordinary Kriging process. Estimation parameters for the Gabriela MZ domain are summarised in Table 66.

Parameter	CutAuXHW	CutAgXHW	HWidth
Minimum number of Comps	4	4	4
Maximum number of Comps	16	16	16
Search Major Distance	200	200	200
Search Orientation	090	090	090
Plunge of Major Axis	0	0	0
Dip of Major Axis	0	0	0
Anisotropy major/semi-major	1.0	1.0	1.0
Anisotropy major/minor	1.0	1.0	1.0
Block Discretization X x Y x Z	5 x 1 x 5	5 x 1 x 5	5 x 1 x 5

Table 66 Gabriela Main Zone – Estimation Parameters

Stockwork

The stockwork mineralisation surrounding the Gabriela MZ domain was estimated using traditional 3D Ordinary Kriging of the uniquely coded 1.5m downhole composite data. All block estimates were based on grade interpolation into X=10m x Z=10m x Y=2m parent cells. Estimation parameters for the Gabriela Stockwork domains are summarised in Table 67.

Parameter	Cut Au	Cut Ag
Minimum number of Comps	6	6
Maximum number of Comps	32	32
Search Major Distance	150	150
Search Orientation	090	090
Plunge of Major Axis	0	0
Dip of Major Axis	0	0
Anisotropy major/semi-major	1.0	1.0
Anisotropy major/minor	1.0	1.0
Block Discretization X x Yx Z	5 x 2 x 5	5 x 2 x 5

Table 67 Gabriela Stockwork Zones – Estimation Parameters

3DBlock Model Definition

A 3D block model was defined in local grid co-ordinates as shown in Table 68. The 3D wireframes for each MZ and Stockwork zone were used to create block model volume constraints for each zone. A standard list of field names and descriptions used in the block model are shown in Table 69.

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Origin	Minimum	Maximum	Model Extent
Y	6400	6800	400
X	22700	24200	1500
Z	-200	200	400
Parent Cell Y	2m	Min Sub-Cell Y m	0.25
Parent Cell X	10m	Min Sub-Cell X m	1.25
Parent Cell Z	10m	Min Sub-Cell Z m	1.25

Table 68 Gabriela Local grid 3D Block Model Definition

Field Name	Description
x	X Block Centroid
y	Y Block Centroid
z	Z Block Centroid
au	Au ppm – Interpolated
ag	Ag ppm – Interpolated
aueq60	Gold Equivalent ppm = Au + Ag/60
density	Density g/cm3 – Direct Assignment – Default 2.5
rescode	Resource Classification - 1=Measured 2=Indicated 3=Inferred
zone	Domain Code Eg. 3101, 3901

Table 69 Gabriela Block Model Attribute Names

Bulk Density

Global bulk densities were assigned as follows:

Gabriela Main Epithermal Vein Zones - 2.54 g/cm3 based on 20 on-site density determinations.

Gabriela Stockwork Zones - 2.5 g/cm3 assumed.

Oxidation

Oxidation of mineralised domains is observed to be of insignificant depth and does not appear to materially influence the bulk density or characteristics of the main mineralised zones. Whilst trenching and sub-outcrop indicates that the Gabriela main zone is very close to the surface in places, Cube has assigned a zero grade to all material within 1m below the topographical surface.

Final Model

The final Surpac block model was clipped by the topography to remove volume above the current surface and exported into CSV format. The following Gabriela model in the Local Grid coordinate system was provided to Extorre:

gab_mar2011_min_zones_lg_06042011.mdl – Surpac format block model with sub-blocking to blocks of Y=0.25m, X=1.25m and Z=1.25m

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gab_mar2011_min_zones_lg_06042011.csv – CSV export of above Surpac model

gab_mar2011_min_zones_reg_lg_07042011.csv – CSV export of above Surpac model regularised to blocks of Y=1m, X=1.25m and Z=2.5m

gab_mar2011_ppct_min_zones_29052011.csv – CSV export of above Surpac model zone partial percentages within blocks of Y=2m, X=10m and Z=10m.

Model Validation

Model validation was carried out visually and by comparing the modelled outcomes against de-clustered composite grades. Figure 121 and Figure 122 show back-calculated model grades for gold and silver together with geological intercept composites for the Gabriela epithermal main zone. The model outcomes show very good correspondence to the informing composite, in particular, providing very good definition of the higher grade zones. Statistical comparisons of composite grades (raw and width weighted) with the modelled outcomes are shown in Table 70 for the Gabriela epithermal main zone. Cube concludes that both visual and statistical comparisons confirm that a robust and unbiased model outcome has been achieved.

Figure 121 Gabriela Main Zone – Composites vs Model Grades – Au ppm

Figure 122 Gabriela Main Zone – Composites vs Model Grades – Ag ppm

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Zone	Raw Composite Grades Au ppm	Composite Grade Au ppm (HW wt)	Model Grade Au ppm	Raw Composite Grades Ag ppm	Composite Grade Ag ppm (HW wt)	Model Grade Ag ppm
3101	2.25	2.42	2.38	342.30	368.68	381.01

Table 70 Gabriela Main Zone - Composite vs Model Grade

Resource Classification and Reporting

Cube has classified a substantial proportion of the Gabriela mineral resources as Indicated where drilling is 40m x 40m or closer. Cube believes that this level of drill hole spacing is sufficient to demonstrate a high level of confidence in the geometry, continuity and grade of the Gabriela deposit. Surrounding areas of Gabriela have been classified as Inferred where drill spacing is wider spaced or where unresolved geological complexity exists. Figure 123 shows the resource classification applied to the Gabriela epithermal main zone. A summary of the Gabriela Indicated and Inferred mineral resources above a cut-off of 1.0 ppm gold equivalent are shown in Table 71 and Table 72 respectively.

Figure 123 Gabriela Main Zone – Resource Classification

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
3101	537,000	2.4	371.0	9.9	42,000	6,411,000	170,000
Total Gabriela Ind.	537,000	2.4	371.0	9.9	42,000	6,411,000	170,000

Table 71 Indicated Gabriela Resources above 1 ppm Gold Equivalent

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
3101	390,000	2.3	394.8	10.2	29,000	4,948,000	128,000
Total Gabriela Inf.	390,000	2.3	394.8	10.2	29,000	4,948,000	128,000

Table 72 Inferred Gabriela Resources above 1 ppm Gold Equivalent

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* Gold equivalent values have been calculated by Extorre on the basis of the following parameters:

●

Long term gold price US$1,320/oz

●

Long term silver price US$26/oz

●

Metallurgical recovery gold 100%

●

Metallurgical recovery silver 100%

The gold equivalent value is calculated by dividing the silver grade (ppm) by 50 (approximate ratio of gold/silver US$ value) and adding it to the gold grade (ppm).

14.6.4

Esperanza and Deborah

Data Types

The Esperanza and Deborah estimates were based on a mixture of Extorre diamond drilling and reverse circulation drill holes. A small number of Mincorp diamond holes were used within Inferred areas of the resource estimate. Trenching was used to aid geological interpretation but trench grades have not been used in the estimation. The datasets used are summarised in Table 73.

Prospect	Extorre RCm	Extorre DDHm	Mincorp DDHm	Total holes/m
Esperanza	19 / 2,237.80	24 / 1,875.95	3 / 188.35	46 / 4,113.75
Deborah	15 / 737	4 / 198.45	2 / 102.15	21 / 1,214.2

Table 73 Drilling Types Esperanza and Deborah

Domaining and Volume Modelling

The main economic mineralised zones within the Esperanza and Deborah prospect areas consist of a series of steeply dipping E-W to ESE vertically extensive narrow (0.1-4m) epithermal quartz vein structures. Economic gold and silver mineralisation primarily occurs in multi-phase banded quartz-adularia and quartz-chlorite-sulphide veins, hydrothermal breccias and peripheral stockworks. The domain outlines used to control volume and estimations have been predominantly based on geological attributes and observations from drill core, particularly epithermal vein textures rather than grade criteria. In most cases, a clear distinction between a main epithermal quartz vein structure and surrounding stockwork mineralisation can be determined based on detailed geological logging and core photography (Figure 124 to Figure 125). In addition, a number of surface trenches and sub-outcrop have been used to guide the interpretation (Figure 126 to Figure 127). Surrounding stockwork zones (Esperanza) are characterised by stockwork veining of varying intensity with minor breccia zones, accompanied by elevated silver grades. There does not appear to be a well developed stockwork envelope surrounding the Deborah main vein.

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Figure 124 Esperanza Drillhole MD315 Showing Styles of Mineralisation

Figure 125 Deborah Drillhole MD078 Showing Styles of Mineralisation

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Figure 126 Esperanza – Quartz Float and Sub-Outcrop

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Figure 127 Deborah – Quartz Float and Sub-Outcrop

The domaining and volume modelling approach was twofold:

Vein Zones: Cross-sections were generated in accordance with the drill hole spacing which typically range between 25m-80m spacing. Sectional interpretations of the main epithermal vein structure (MZ) using geological logging and core photography was completed over the full strike extent of each prospect. The resulting mineralisation interpretations were graphically digitised and a 3D wireframe model produced. The interpretation of each MZ was based on geological attributes such as epithermal vein textures ensuring that the final 3D mineralisation model reflects an in-situ geological model whereby no cut-off grade or minimum mining width criteria has been applied.

Figure 128 to Figure 132 show wireframes of the Esperanza and Deborah main epithermal vein zones together with drilling locations.

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Figure 128 Esperanza Main Epithermal Vein Zone – Plan View GK Grid

Figure 129 Esperanza Main Epithermal Vein Zone – Plan View Local Grid

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Figure 130 Deborah Main Epithermal Vein Zone – Plan View GK Grid

Figure 131 Deborah Main Epithermal Vein Zone – Plan View Local Grid

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Figure 132 Esperanza and Deborah Main Epithermal Vein Zone – Long Section View Local Grid

Database Coding and Compositing

Samples within the MZ and surrounding stockworks are assigned a unique database code which is used to control the compositing process. A four digit numbering system (zonecode) was created to define the various domains within the minor prospects (Table 74).

Zonecode	Domain
4101	Esperanza Main Zone
4102	Esperanza Stockwork
5101	Deborah Main Zone

Table 74 Esperanza and Deborah Domain Numbering

Main Zones (Esperanza 4101 and Deborah 5101)

There are several key physical features of the Esperanza and Deborah epithermal vein zones that need to be considered and accommodated by the selected resource modelling technique. These features are:

●

Mineralised epithermal vein zone true thickness is variable typically ranging from 0.1m to 4m in width.

●

Undulating or variable zone geometry (dip and strike) and possible grade/thickness trends within this variable geometry;

●

Sampling has been taken over geological intervals creating samples of unequal length or variable support;

●

Mining selectivity across the epithermal vein zones is unlikely due the narrow nature of the mineralised structures;

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Drill spacing is variable but typically around 25m by 80m or larger within the plane of mineralisation.

A method that allows estimation of metal content on a projected plane is considered to be an appropriate approach to addressing the features outlined above. This resource modelling approach is best achieved using geological intercept composites and accumulation estimation. MZ grades are composited across the entire coded interval resulting in a single geological intercept composite at each intercept location. The estimation approach for the Esperanza and Deborah prospects differs slightly from the Escondida methodology whereby the accumulation and width variables were estimated into 3D blocks orientated parallel to the vein orientation in Local Grid coordinates. In this orientation a projection to a 2D plane was not considered necessary as each MZ is aligned by design, closely to Local Grid East-West. The mid-point of each geological composite is assigned the horizontal width of the vein structure and used to compute a ‘metal accumulation’ variable.

a(x) = t(x) . z(x)

Stockworks (Esperanza 4102)

Descriptive Statistics and High Grade Capping

Main Zones (Esperanza 4101 and Deborah 5101)

Unlike the Escondida MZ, the coded MZ domains for Esperanza and Deborah contain few extreme outlier samples. An examination of gold and silver geological composite statistics over the coded intervals was undertaken prior to calculating the accumulation to identify any outlier composites that may require high grade capping. It was decided to apply a modest high grade assay cut to gold and silver in all MZ zones. Whilst the application of high grade capping is somewhat arbitrary it is considered by the Author to be a prudent and necessary risk adjustment strategy at this stage of project evaluation.

Figure 133 and Figure 134 show log-probability plots of uncut gold and silver geological composites respectively for the Esperanza MZ domain.

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Figure 133 Esperanza Main Zone – Geological Composites – Au ppm – Log-Probability Plot

Figure 134 Esperanza Main Zone –Geological Composites – Ag ppm– Log Probability Plot

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Figure 135 and Figure 136 show log-probability plots of uncut gold and silver geological composites respectively for the Deborah MZ domain.

Figure 135 Deborah Main Zone – Geological Composites – Au ppm – Log-Probability Plot

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Figure 136 Deborah Main Zone –Geological Composites – Ag ppm– Log Probability Plot

A list of Esperanza and Deborah geological intercept composites is included in Section 14.8.

Table 75 details the high grade caps applied to geological composite data prior to calculating the accumulation variables for the minor prospects.

Domain	Zonecode	Gold Cap ppm	Number Capped	Percentile	Silver Cap ppm	Number Capped	Percentile
Esperanza Main Zone	4101	10.0	3	93rd	500	3	93rd
Deborah Main Zone	5101	4.0	1	96th	None

Table 75 Esperanza and Deborah High Grade Caps

Table 76 and Table 77 summarise the Esperanza MZ raw geological intercept grades (cut and uncut) and accumulation (cut only) statistics respectively. It can be seen from Table 77 that declustering has an impact on the average intercept grade reducing the global gold and silver grade by 15% and 2% respectively.

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	Raw Au ppm	Raw Ag ppm	Cut Au ppm	Cut Ag ppm
Number	34	34	34	34
Minimum	0.005	1.1	0.005	1.1
Maximum	17.12	664.0	10	500
Raw Mean	3.32	176.9	2.92	169.1
Declust. Mean 60x60	2.97	174.9	2.57	166.4
Std Dev	4.23	177.7	3.13	158.8
Coeff Var	1.27	1.00	1.07	0.93

Table 76 Esperanza Main Zone - Summary Statistics – Geological Composites - Raw Grades

	Horizontal Width	Cut Au X HW	Cut Ag X HW	HWidth Weighted Cut Au ppm	HWidth Weighted Cut Ag ppm
Number	34	34	34
Minimum	0.27	0.007	1.499
Maximum	4.76	20.2	1647.203
Raw Mean	1.701	4.566	294.741	2.68	173.3
Declust. Mean 60x60	1.67	3.83	286.0	2.29	171.3
Std Dev	1.129	5.07	361.796
Coeff Var	0.664	1.11	1.228

Table 77 Esperanza Main Zone - Summary Statistics – Geological Composites - Accumulations

Table 78 and Table 79 summarise the Deborah MZ raw geological intercept grades (cut and uncut) and accumulation (cut only) statistics respectively. It can be seen from Table 79 that declustering has an impact on the average intercept grade reducing the global gold and silver grade by 11% and 7% respectively.

	Raw Au ppm	Raw Ag ppm	Cut Au ppm	Cut Ag ppm
Number	20	20	20	20
Minimum	0.75	5.3	0.75	5.3
Maximum	7.38	93.0	4.00	93.0
Raw Mean	2.36	46.5	2.19	46.5
Declust. Mean 60x60	2.30	42.74	2.19	42.74
Std Dev	1.51	28.04	1.04	28.04
Coeff Var	0.64	0.60	0.47	0.60

Table 78 Deborah Main Zone - Summary Statistics – Geological Composites - Raw Grades

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	Horizontal Width	Cut Au X HW	Cut Ag X HW	HWidth Weighted Cut Au ppm	HWidth Weighted Cut Ag ppm
Number	20	20	20
Minimum	1.17	1.19	6.2
Maximum	9.65	29.42	662.9
Raw Mean	4.63	11.06	233.8	2.63	50.5
Declust. Mean 60x60	4.55	10.67	215.5	2.35	47.4
Std Dev	2.44	8.34	196.9
Coeff Var	0.53	0.75	0.84

Table 79 Deborah Main Zone - Summary Statistics – Geological Composites - Accumulations

Stockwork

Log-probability plots of gold and silver 1.5m composites for the Esperanza stockwork domain are shown in Figure 137 and Figure 138 respectively. An assay top cut of 1.5 Au ppm and 80 Ag ppm was applied to gold and silver respectively as a model risk adjustment strategy. Table 80 shows summary statistics of raw and cut gold and silver 1.5m composites for Esperanza stockwork mineralisation.

	Raw Au ppm	Raw Ag ppm	Cut Au ppm	Cut Ag ppm
Number	175	175	175	175
Minimum	0.004	0.8	0.004	0.8
Maximum	5.65	484.0	1.50	80.0
Raw Mean	0.24	23.9	0.19	19.2
Std Dev	0.08	52.2	0.29	17.2
Coeff Var	2.63	2.18	1.48	0.90

Table 80 Esperanza Stockwork Zone - Summary Statistics – 1.5m Downhole Composites

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Figure 137 Esperanza Stockwork Domain – 1.5m Downhole Composites – Cut Au ppm – Log-Probability Plot

Figure 138 Esperanza Stockwork Domain – 1.5m Downhole Composites – Cut Ag ppm – Log-Probability Plot

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Variography

Main Zones (Esperanza and Deborah)

Variography was used to characterise the spatial continuity of the horizontal width and accumulation variables within the plane of each vein structure. No obvious direction of preferred continuity was determined and omni-directional variogram models were fitted for all variables. Variograms of horizontal width and accumulations demonstrated similar spatial structure and variance proportions. Given the similarities between these variograms a simplification was made whereby a single variogram model was used to estimate both the accumulation and thickness variables. This approach has the advantage that it is simpler and ensures no instability in the back-calculated block grades due to unstable or unexpected width estimations which may occur when a different variogram model is used.

Experimental variography was undertaken on Gaussian transformed data. Variogram models were fitted to the Gaussian data and subsequently back-transformed into real variances to obtain the appropriate parameters for interpolation of the raw accumulation variables by Ordinary Kriging. Back-transformed variogram model parameters for Esperanza and Deborah are shown in Table 81 and Table 82 respectively.

		Sill (Relative Variance)	Range	Azimuth	Plunge	Dip	Major/ Semi Major Ratio	Major/ Minor Ratio
Au*HW	Nugget Co	0.47
	Structure 1	0.28	60	90	0	0	1	1
	Structure 2	0.26	150	90	0	0	1	1
Ag*HW	Nugget Co	0.49
	Structure 1	0.37	150	90	0	0	1	1
	Structure 2	0.14	200	90	0	0	1	1

Table 81 Esperanza Main Zone – Variogram Models – Back-Transformed Raw Au*HW and Ag*HW

		Sill (Relative Variance)	Range	Azimuth	Plunge	Dip	Major/ Semi Major Ratio	Major/ Minor Ratio
Au*HW	Nugget Co	0.11
	Structure 1	0.89	150	90	0	0	1	1
Ag*HW	Nugget Co	0.11
	Structure 1	0.89	120	90	0	0	1	1

Table 82 Deborah Main Zone – Variogram Models – Back-Transformed Raw Au*HW and Ag*HW

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Stockwork Zones (Esperanza)

Experimental variography was undertaken on cut Gaussian transformed 1.5m down hole composites. Variogram models were fitted to the Gaussian data and subsequently back-transformed into real variances to obtain the appropriate parameters for interpolation of the Cut Au and Cut Ag variables by Ordinary Kriging. Back-transformed variogram model parameters for Esperanza are shown in Table 83.

		Sill (Relative Variance)	Range	Azimuth	Plunge	Dip	Major/ Semi Major Ratio	Major/ Minor Ratio
Cut Au	Nugget Co	0.48
	Structure 1	0.31	8	90	0	0	1	1
	Structure 2	0.20	55	90	0	0	1	1
Cut Ag	Nugget Co	0.51
	Structure 1	0.31	8	90	0	0	1	1
	Structure 2	0.18	130	90	0	0	1	1

Table 83 Esperanza Stockwork – Variogram Models – Cut Au and Cut Ag

Grade Estimation

Main Zones (Esperanza and Deborah)

The estimation approach for the Esperanza and Deborah prospects differs slightly from the Escondida methodology whereby the accumulation and horizontal width variables were estimated by 3D Ordinary Kriging into X=20m x Y=2m x Z=20m blocks orientated parallel to the vein orientation in Local Grid coordinates. In this orientation a projection to a 2D plane was not considered necessary as each MZ is aligned by design, closely to Local Grid East-West. Final block grade is calculated by dividing the estimated accumulation by the estimated thickness. The estimation of horizontal width used for block grade back-calculation used the same variogram models as for the accumulations. This simplification ensures no instability in the back calculated grades due to unstable or unexpected width estimations which may occur when a different variogram is used. Separate estimation was undertaken for Esperanza and Deborah (4101 and 5101) using the uniquely coded composite data.

Kriging neighbourhood analysis was undertaken to determine appropriate inputs to the Ordinary Kriging process. Estimation parameters for the MZ domains are summarised in Table 84 and Table 85.

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Parameter	CutAuXHW	CutAgXHW	HWidth
Minimum number of Comps	2	2	2
Maximum number of Comps	4	4	4
Search Major Distance	100	100	100
Search Orientation	090	090	090
Plunge of Major Axis	0	0	0
Dip of Major Axis	0	0	0
Anisotropy major/semi-major	1.0	1.0	1.0
Anisotropy major/minor	1.0	1.0	1.0
Block Discretization X x Z	5 x 5	5 x 5	5 x 5

Table 84 Esperanza Main Zone – Estimation Parameters

Parameter	CutAuXHW	CutAgXHW	HWidth
Minimum number of Comps	2	2	2
Maximum number of Comps	4	4	4
Search Major Distance	200	200	200
Search Orientation	090	090	090
Plunge of Major Axis	0	0	0
Dip of Major Axis	0	0	0
Anisotropy major/semi-major	1.0	1.0	1.0
Anisotropy major/minor	1.0	1.0	1.0
Block Discretization X x Z	5 x 5	5 x 5	5 x 5

Table 85 Deborah Main Zone – Estimation Parameters

Stockwork

The stockwork mineralisation surrounding the epithermal vein zones was estimated using traditional 3D Ordinary Kriging of the uniquely coded 1.5m downhole composite data. All block estimates were based on grade interpolation into X=20m x Z=20m x Y=2m parent cells. Estimation parameters for the Stockwork domains (Esperanza) are summarised in Table 86.

Parameter	Cut Au	Cut Ag
Minimum number of Comps	6	6
Maximum number of Comps	32	32
Search Major Distance	150	150
Search Orientation	090	090
Plunge of Major Axis	0	0
Dip of Major Axis	0	0
Anisotropy major/semi-major	1.0	1.0
Anisotropy major/minor	1.0	1.0
Block Discretization X x Yx Z	5 x 2 x 5	5 x 2 x 5

Table 86 Esperanza Stockwork Zones – Estimation Parameters

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3D Block Model Definition

Separate 3D models in local grid co-ordinates were defined for the Esperanza and Deborah prospects. The 3D wireframes for each MZ and Stockwork zone were used to create block model volume constraints for each zone. The individual block model definitions are shown in Table 87 to Table 88. A standard list of field names and descriptions used in the block model are shown in Table 89.

Origin	Minimum	Maximum	Model Extent
Y	4850	5250	400
X	22000	23600	1600
Z	-200	200	400
Parent Cell Y	2m	Min Sub-Cell Y m	0.25
Parent Cell X	20m	Min Sub-Cell X m	10
Parent Cell Z	20m	Min Sub-Cell Z m	10

Table 87 Esperanza Local grid 3D Block Model Definition

Origin	Minimum	Maximum	Model Extent
Y	4850	5200	350
X	39800	40800	1000
Z	-200	200	400
Parent Cell Y	2m	Min Sub-Cell Y m	0.25
Parent Cell X	20m	Min Sub-Cell X m	10
Parent Cell Z	20m	Min Sub-Cell Z m	10

Table 88 Deborah Local grid 3D Block Model Definition

Field Name	Description
x	X Block Centroid
y	Y Block Centroid
z	Z Block Centroid
au	Au ppm – Interpolated
ag	Ag ppm – Interpolated
aueq60	Gold Equivalent ppm = Au + Ag/60
density	Density g/cm3 – Direct Assignment – Default 2.5
rescode	Resource Classification - 1=Measured 2=Indicated 3=Inferred
zone	Domain Code Eg. 1101, 1201, 1301, 1401, 1901 , 1902

Table 89 Esperanza and Deborah Block Model Attribute Names

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Bulk Density

Global bulk densities were assigned to each model as follows:

Esperanza Main Epithermal Vein Zones - 2.58 g/cm3 based on 15 on-site density determinations.

Esperanza Stockwork Zones - 2.5 g/cm3 assumed.

Deborah Main Epithermal Vein Zones - 2.6 g/cm3 assumed.

Oxidation

Oxidation of mineralised domains is observed to be of insignificant depth and does not appear to materially influence the bulk density or characteristics of the main mineralised zones.

Final Model

The final Surpac block models were clipped by the topography to remove volume above current surface and exported into CSV format. The following models in the Local Grid coordinate system were provided to Extorre:

Esperanza - esp_mar2010_min_zones.csv

Deborah - deb_mar2010_min_zones.csv.

Model Validation

Model validation was carried out visually and by comparing the modelled outcomes against de-clustered composite grades. Figure 139 to Figure 142 show back-calculated model grades for gold and silver together with geological intercept composites for the Esperanza and Deborah prospects. Statistical comparisons of composite grades (raw and declustered width weighted) with the modelled outcomes are shown in Table 90 for the epithermal main zones. Cube concludes that both visual and statistical comparisons confirm that a robust and unbiased model outcome has been achieved.

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Figure 139 Esperanza Main Zone – Composites vs Model Grades – Au ppm

Figure 140 Esperanza Main Zone – Composites vs Model Grades – Ag ppm

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Figure 141 Deborah Main Zone –Composites vs Model Grades – Au ppm

Figure 142 Deborah Main Zone –Composites vs Model Grades – Ag ppm

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Zone	De-clustering Cell Size	Raw Composite Grades Au ppm	De-clustered Composite Grade Au ppm (HW wt)	Model Grade Au ppm	Raw Composite Grades Ag ppm	De-clustered Composite Grade Ag ppm (HW wt)	Model Grade Ag ppm
4101	Y=60 X=60 Z=30	2.68	2.29	2.60	173.3	171.3	171.1
5101	Y=60 X=60 Z=30	2.63	2.35	2.40	50.5	47.4	48.1

Table 90 Esperanza and Deborah Main Zones - De-Clustered Composite vs Model Grade

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Resource Classification and Reporting

Cube has classified all of the Esperanza and Deborah mineral resources as Inferred. Cube believes that the level of drill hole spacing and other information is sufficient to demonstrate a level of confidence in the geometry, continuity and grade of these deposits consistent with the Inferred resource category. A summary of the Esperanza and Deborah Inferred mineral resources above a cut-off of 1 ppm gold equivalent is shown in Table 91.

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
Esperanza 4101	371,000	2.6	175.0	6.1	31,000	2,090,000	72,000
Deborah 5101	579,000	2.4	48.1	3.4	45,000	896,000	63,000

Total Inferred	950,000	2.5	97.7	4.4	76,000	2,986,000	135,000

Table 91 Inferred Esperanza and Deborah Resources above 1 ppm Gold Equivalent

* Gold equivalent values have been calculated by Extorre on the basis of the following parameters:

●

Long term gold price US$1,320/oz

●

Long term silver price US$26/oz

●

Metallurgical recovery gold 100%

●

Metallurgical recovery silver 100%

The gold equivalent value is calculated by dividing the silver grade (ppm) by 50 (approximate ratio of gold/silver US$ value) and adding it to the gold grade (ppm).

14.7	Cerro Moro Resource Reporting

A summary of the Cerro Moro Indicated and Inferred mineral resources above a cut-off of 1 ppm gold equivalent are shown in Table 92 and Table 93 respectively.

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Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
Escondida	620,000	18.8	829.2	35.4	374,000	16,530,000	705,000
Loma Escondida	44,000	18.4	919.5	36.8	26,000	1,297,000	52,000
Gabriela	537,000	2.4	371.0	9.9	42,000	6,411,000	170,000
Total	1,201,000	11.5	627.5	24.0	443,000	24,238,000	927,000

Table 92 Cerro Moro Indicated Mineral Resources above 1 ppm Gold Equivalent

Zone	Tonnes	Gold (ppm)	Silver (ppm)	Gold Equivalent Grade* (ppm)	Gold (ounces)	Silver (ounces)	Gold Equivalent Ounces*
Escondida	508,000	4.3	164.8	7.6	70,000	2,689,000	123,000
Loma Escondida	13,000	9.7	595.4	21.6	4,000	256,000	9,000
Gabriela	390,000	2.3	394.8	10.2	29,000	4,948,000	128,000
Esperanza	371,000	2.6	175.0	6.1	31,000	2,090,000	72,000
Deborah	579,000	2.4	48.1	3.4	45,000	896,000	63,000
Total	1,861,000	3.0	181.8	6.6	178,000	10,879,000	396,000

Table 93 Cerro Moro Inferred Mineral Resources above 1 ppm Gold Equivalent

* Gold equivalent values have been calculated by Extorre on the basis of the following parameters:

●

Long term gold price US$1,320/oz

●

Long term silver price US$26/oz

●

Metallurgical recovery gold 100%

●

Metallurgical recovery silver 100%

The gold equivalent value is calculated by dividing the silver grade (ppm) by 50 (approximate ratio of gold/silver US$ value) and adding it to the gold grade (ppm).

Key Assumptions, Parameters, and Methods Used to Estimate Mineral Resources

It is the opinion of the Qualified Person that the Indicated and Inferred mineral resources summarised in Table 92 and Table 93 respectively have reasonable prospects for economic extraction. A number of factors have been taken into consideration when quantifying a subset of the Cerro Moro mineralised system with reasonable prospects for economic extraction. These include:

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The gold and silver grade of the Cerro Moro mineralisation is high by world standards. High grade mineralisation at Cerro Moro occurs in well developed “shoot-like” zones surrounded by minor occurrences of low grade material within a distinct geological structure.

●

Extorrereleased a PEA on December 2, 2010 that estimated open pit mining costs at U.S.$3.40/tonne (including tailings co-disposal) and milling and processing costs at U.S.$38/tonne. These costs provide a cut-off grade of 1.8 Au ppm gold equivalent.

●

Based on preliminary metallurgical testwork gold and silver recoveries have been estimated at 95% and 90%, respectively.

●

Base assumptions for long-term gold and silver prices are U.S.$1,320/oz and U.S.$26/oz, respectively.

●

Less than 2% of the Cerro Moro gold equivalent mineral resources reported in the above Tables fall below a nominal economic cut-off grade of 1.8 Au ppm.

The Qualified Person has extensive experience in mining similar deposits in other parts of the world with both open pit and underground methods and expects mining will be focused on extracting the full geological structure including some peripheral sub-grade mineralised material. It is the Qualified Person’s opinion that the small proportion of mineralised material that falls below a nominal economic cut-off grade of 1.8 Au ppm should be included within reportable resources as there is a strong likelihood that the majority of this material will be mined in conjunction with extracting the potentially economic component of the mineralised system. The Qualified Person believes that a range of potentially viable low grade processing options, such as heap leach technology, may provide realistic opportunities to exploit lower grade resources in the Cerro Moro project area. These options could be particularly attractive given the relatively shallow nature of mineralisation readily amenable to open-pit mining. Such options could result in processing costs being as low as U.S.$5.50/tonne resulting in potentially economic cut-off grades of 0.3 Au ppm gold equivalent.

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14.8	List of Geological Intercepts

Prospect	Type	Hole ID	From	To	Length	HW (m)	Au g/t	Ag g/t
Escondida	Drillhole	MD0061	39.26	41.20	1.94	1.35	37.29	327.66
Escondida	Drillhole	MD0062	25.47	27.90	2.43	1.78	2.43	104.82
Escondida	Drillhole	MD0064	36.00	38.60	2.60	1.95	44.97	3056.81
Escondida	Drillhole	MD0065	41.48	43.70	2.22	1.51	0.97	31.69
Escondida	Drillhole	MD0082	131.30	132.89	1.59	1.00	223.05	426.72
Escondida	Drillhole	MD0083	119.22	120.17	0.95	0.57	0.81	19.90
Escondida	Drillhole	MD0085	62.00	63.27	1.27	1.16	3.46	51.70
Escondida	Drillhole	MD0086	9.95	10.26	0.31	0.18	0.64	18.10
Escondida	Drillhole	MD0087	30.06	32.07	2.01	1.46	19.10	409.29
Escondida	Drillhole	MD0088	35.40	38.10	2.70	1.88	6.67	288.38
Escondida	Drillhole	MD0089	65.70	66.04	0.34	0.27	9.19	21.80
Escondida	Drillhole	MD0090	68.00	70.70	2.70	2.17	30.83	222.78
Escondida	Drillhole	MD0091	85.49	88.90	3.41	2.66	47.13	279.57
Escondida	Drillhole	MD0095	43.10	45.95	2.85	2.62	48.70	1773.05
Escondida	Drillhole	MD0096	10.68	12.90	2.22	1.81	22.66	437.63
Escondida	Drillhole	MD0097	67.00	67.72	0.72	0.76	2.02	2.14
Escondida	Drillhole	MD0098	77.49	81.32	3.83	2.89	101.73	3331.77
Escondida	Drillhole	MD0104a	155.00	159.82	4.82	3.55	3.05	84.83
Escondida	Drillhole	MD0105	79.00	80.00	1.00	0.85	0.66	39.10
Escondida	Drillhole	MD0106	29.40	31.00	1.60	1.13	129.62	3939.56
Escondida	Drillhole	MD0107	85.84	86.17	0.33	0.24	2.44	26.60
Escondida	Drillhole	MD0108	23.08	24.06	0.98	0.76	18.19	619.19
Escondida	Drillhole	MD0109	126.50	128.73	2.23	1.68	2.71	85.54
Escondida	Drillhole	MD0110	27.39	29.08	1.69	1.26	43.34	970.92
Escondida	Drillhole	MD0111	26.50	27.78	1.28	0.96	60.73	439.43
Escondida	Drillhole	MD0112	100.40	102.84	2.44	1.71	44.62	3095.01
Escondida	Drillhole	MD0113	51.00	52.24	1.24	1.07	0.27	18.96
Escondida	Drillhole	MD0114	54.37	55.00	0.63	0.46	1.80	11.35
Escondida	Drillhole	MD0115	54.35	55.73	1.38	1.01	45.61	3305.48
Escondida	Drillhole	MD0116	51.18	52.47	1.29	0.99	0.44	32.83
Escondida	Drillhole	MD0117	43.99	45.60	1.61	1.24	19.50	92.87
Escondida	Drillhole	MD0118	50.63	52.60	1.97	1.46	51.84	79.84
Escondida	Drillhole	MD0119	133.70	135.66	1.96	1.39	1.68	61.82
Escondida	Drillhole	MD0120	80.62	82.26	1.64	1.15	0.16	9.84
Escondida	Drillhole	MD0121	42.12	43.28	1.16	1.06	0.39	19.29
Escondida	Drillhole	MD0122	39.70	40.00	0.30	0.29	2.31	100.00
Escondida	Drillhole	MD0123	32.74	34.75	2.01	1.86	5.24	286.93
Escondida	Drillhole	MD0124	31.65	32.20	0.55	0.48	0.07	2.39
Escondida	Drillhole	MD0125	28.48	29.63	1.15	0.99	0.10	2.81
Escondida	Drillhole	MD0135	76.00	78.20	2.20	1.66	9.96	139.54
Escondida	Drillhole	MD0136	70.36	73.00	2.64	1.85	1.08	75.14
Escondida	Drillhole	MD0137	72.28	73.13	0.85	0.65	20.86	199.71

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Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Escondida	Drillhole	MD0139	30.72	33.97	3.25	2.12	33.84	539.83
Escondida	Drillhole	MD0140	81.00	85.25	4.25	2.30	13.79	481.96
Escondida	Drillhole	MD0142	117.90	118.80	0.90	0.62	0.53	70.85
Escondida	Drillhole	MD0145	50.43	51.00	0.57	0.52	2.12	210.46
Escondida	Drillhole	MD0147	22.00	23.28	1.28	1.11	0.41	9.41
Escondida	Drillhole	MD0148	131.40	132.10	0.70	0.53	1.70	34.91
Escondida	Drillhole	MD0149	41.83	42.55	0.72	0.69	13.04	1288.00
Escondida	Drillhole	MD0150	57.50	58.05	0.55	0.50	0.10	6.20
Escondida	Drillhole	MD0151	98.40	99.50	1.10	0.86	8.51	19.84
Escondida	Drillhole	MD0152	99.40	100.47	1.07	0.82	6.86	68.40
Escondida	Drillhole	MD0153	26.89	27.65	0.76	0.57	1.66	71.05
Escondida	Drillhole	MD0167	93.80	94.94	1.14	1.13	2.97	266.16
Escondida	Drillhole	MD0168	84.97	86.00	1.03	0.99	0.91	42.70
Escondida	Drillhole	MD0169	35.00	37.25	2.25	2.06	6.97	237.82
Escondida	Drillhole	MD0170	63.80	68.00	4.20	2.25	74.63	2572.23
Escondida	Drillhole	MD0206	37.85	42.70	4.85	2.62	1.50	121.76
Escondida	Drillhole	MD0214	32.06	32.98	0.92	0.63	2.42	28.73
Escondida	Drillhole	MD0216	97.87	101.53	3.66	2.28	66.05	4550.72
Escondida	Drillhole	MD0218	152.08	155.82	3.74	2.43	48.80	3116.19
Escondida	Drillhole	MD0221	74.00	79.00	5.00	3.35	29.04	2003.10
Escondida	Drillhole	MD0224	90.18	92.80	2.62	1.41	0.42	34.48
Escondida	Drillhole	MD0226	196.45	202.00	5.55	3.92	4.32	252.37
Escondida	Drillhole	MD0228	79.00	81.75	2.75	2.46	23.38	1255.42
Escondida	Drillhole	MD0229	149.25	153.35	4.10	3.28	5.78	462.31
Escondida	Drillhole	MD0233	173.72	175.40	1.68	1.13	4.41	325.83
Escondida	Drillhole	MD0237	331.20	334.17	2.97	2.35	0.27	36.32
Escondida	Drillhole	MD0249	118.00	120.00	2.00	1.44	0.23	10.95
Escondida	Drillhole	MD0258	228.84	235.24	6.40	4.13	1.00	57.40
Escondida	Drillhole	MD0259	208.80	213.07	4.27	2.26	0.67	32.91
Escondida	Drillhole	MD0259A	191.00	195.00	4.00	2.69	3.90	445.44
Escondida	Drillhole	MD0260	130.22	132.00	1.78	1.57	10.45	479.81
Escondida	Drillhole	MD0261	70.30	71.20	0.90	0.86	3.23	132.00
Escondida	Drillhole	MD0262	76.00	78.00	2.00	1.55	0.58	40.13
Escondida	Drillhole	MD0268	247.75	249.10	1.35	0.92	0.14	10.58
Escondida	Drillhole	MD0270	251.67	253.00	1.33	1.08	9.50	354.75
Escondida	Drillhole	MD0271	230.40	230.78	0.38	0.25	0.22	2.81
Escondida	Drillhole	MD0274	252.28	258.20	5.92	5.51	0.36	28.35
Escondida	Drillhole	MD0283	236.63	238.77	2.14	1.96	0.44	18.69
Escondida	Drillhole	MD0288	180.95	183.15	2.20	1.53	0.60	68.56
Escondida	Drillhole	MD0291	161.45	162.72	1.27	1.02	1.52	76.27
Escondida	Drillhole	MD0292A	33.00	36.03	3.03	1.69	1.54	53.58
Escondida	Drillhole	MD0293	81.00	83.00	2.00	1.52	188.81	4706.68
Escondida	Drillhole	MD0296	92.00	93.00	1.00	0.86	0.13	0.40
Escondida	Drillhole	MD0298	29.15	30.30	1.15	1.03	30.33	600.30

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Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Escondida	Drillhole	MD0299	64.03	68.08	4.05	2.42	13.39	1246.87
Escondida	Drillhole	MD0301	237.23	239.11	1.88	1.70	0.56	26.48
Escondida	Drillhole	MD0304	241.45	242.45	1.00	0.83	0.03	6.80
Escondida	Drillhole	MD0307	191.78	192.38	0.60	0.51	0.52	28.95
Escondida	Drillhole	MD0310	285.08	287.51	2.43	1.70	0.75	81.77
Escondida	Drillhole	MD0311	90.15	91.75	1.60	1.78	24.65	2557.50
Escondida	Drillhole	MD0340	124.12	124.70	0.58	0.52	20.95	54.00
Escondida	Drillhole	MD0342	205.95	206.55	0.60	0.48	7.45	3181.00
Escondida	Drillhole	MD0344	143.00	143.87	0.87	0.73	0.75	16.70
Escondida	Drillhole	MD0345	99.05	100.00	0.95	0.84	0.23	22.59
Escondida	Drillhole	MD0346	340.00	342.94	2.94	2.22	0.52	63.48
Escondida	Drillhole	MD0348	270.95	273.12	2.17	1.74	1.55	54.45
Escondida	Drillhole	MD0354	165.00	165.75	0.75	0.64	0.20	7.41
Escondida	Drillhole	MD0389	69.45	72.39	2.94	1.86	82.98	7130.15
Escondida	Drillhole	MD0398	113.03	115.96	2.93	1.91	36.22	2468.34
Escondida	Drillhole	MD0405	188.29	190.50	2.21	1.56	0.65	28.21
Escondida	Drillhole	MD0444	108.50	110.00	1.50	1.70	4.01	171.79
Escondida	Drillhole	MD0445	152.46	154.07	1.61	1.26	19.86	1506.94
Escondida	Drillhole	MD0446	144.65	145.50	0.85	0.70	4.67	490.17
Escondida	Drillhole	MD0448	171.28	171.89	0.61	0.50	0.23	18.15
Escondida	Drillhole	MD0449	110.12	110.70	0.58	0.50	3.45	83.97
Escondida	Drillhole	MD0450	186.95	188.54	1.59	1.14	0.50	24.55
Escondida	Drillhole	MD0451	222.88	224.12	1.24	0.80	0.06	8.02
Escondida	Drillhole	MD0452	120.30	120.60	0.30	0.28	0.13	8.19
Escondida	Drillhole	MD0453	107.60	108.12	0.52	0.41	3.04	223.00
Escondida	Drillhole	MD0454	175.60	178.19	2.59	1.73	79.71	118.39
Escondida	Drillhole	MD0455	198.85	202.51	3.66	2.58	0.76	19.53
Escondida	Drillhole	MD0456	131.55	133.85	2.30	1.71	0.73	12.70
Escondida	Drillhole	MD0457	91.05	92.05	1.00	0.71	1.52	56.40
Escondida	Drillhole	MD0458	94.06	96.06	2.00	1.42	0.56	18.44
Escondida	Drillhole	MD0459	45.75	50.22	4.47	3.06	15.12	246.41
Escondida	Drillhole	MD0465	221.41	222.50	1.09	0.81	2.51	117.00
Escondida	Drillhole	MD0466	207.10	208.10	1.00	0.74	0.37	23.50
Escondida	Drillhole	MD0467	165.60	166.94	1.34	1.26	11.05	383.63
Escondida	Drillhole	MD0468	62.80	66.00	3.20	2.11	10.94	503.89
Escondida	Drillhole	MD0469	122.60	129.25	6.65	3.46	6.52	304.40
Escondida	Drillhole	MD0470	187.40	189.30	1.90	1.13	1.05	104.60
Escondida	Drillhole	MD0471	200.20	207.00	6.80	3.85	1.64	138.20
Escondida	Drillhole	MD0472	27.00	28.62	1.62	1.09	0.78	37.00
Escondida	Drillhole	MD0473	77.54	78.61	1.07	0.61	0.93	63.82
Escondida	Drillhole	MD0474	129.80	132.27	2.47	1.43	4.20	469.30
Escondida	Drillhole	MD0475	137.65	140.25	2.60	1.40	36.75	2966.14
Escondida	Drillhole	MD0476	148.00	151.80	3.80	2.51	2.52	197.20
Escondida	Drillhole	MD0477	141.64	142.68	1.04	0.74	1.79	182.00

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Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Escondida	Drillhole	MD0478	132.90	135.84	2.94	1.98	0.56	16.86
Escondida	Drillhole	MD0479	117.95	121.20	3.25	2.08	51.47	1851.71
Escondida	Drillhole	MD0480	199.33	200.70	1.37	1.08	1.11	32.78
Escondida	Drillhole	MD0481	175.35	178.25	2.90	1.61	6.56	685.54
Escondida	Drillhole	MD0482	206.50	207.00	0.50	0.35	0.09	0.70
Escondida	Drillhole	MD0483	237.15	240.76	3.61	2.08	1.90	9.16
Escondida	Drillhole	MD0484	159.34	160.54	1.20	0.82	3.20	87.50
Escondida	Drillhole	MD0485	211.40	213.00	1.60	1.07	0.77	70.72
Escondida	Drillhole	MD0486	30.10	33.60	3.50	2.37	0.55	39.43
Escondida	Drillhole	MD0487	21.00	24.10	3.10	2.08	1.08	34.87
Escondida	Drillhole	MD0489	10.45	11.80	1.35	0.92	5.58	8.40
Escondida	Drillhole	MD0492	10.00	14.60	4.60	2.79	101.11	2507.13
Escondida	Drillhole	MD0493	44.50	46.60	2.10	1.35	42.94	1247.48
Escondida	Drillhole	MD0494	8.63	12.00	3.37	1.68	2.88	25.00
Escondida	Drillhole	MD0495	10.80	13.00	2.20	1.29	5.32	68.38
Escondida	Drillhole	MD0496	53.76	56.75	2.99	1.91	2.27	158.97
Escondida	Drillhole	MD0497	22.10	23.30	1.20	1.06	0.65	62.63
Escondida	Drillhole	MD0498	47.46	51.62	4.16	2.63	5.93	136.47
Escondida	Drillhole	MD0499	8.65	13.60	4.95	3.00	42.75	730.38
Escondida	Drillhole	MD0500	53.78	55.60	1.82	1.32	0.38	23.73
Escondida	Drillhole	MD0501	48.00	52.85	4.85	2.90	12.72	108.53
Escondida	Drillhole	MD0502	12.25	15.00	2.75	1.95	1.35	105.45
Escondida	Drillhole	MD0503	43.76	46.00	2.24	1.39	2.23	166.95
Escondida	Drillhole	MD0504	15.00	17.20	2.20	1.42	2.84	304.61
Escondida	Drillhole	MD0505	50.70	53.52	2.82	1.97	2.84	97.00
Escondida	Drillhole	MD0506	54.00	55.96	1.96	1.78	133.21	4471.21
Escondida	Drillhole	MD0507	61.00	62.58	1.58	1.43	2.11	265.78
Escondida	Drillhole	MD0508	39.22	39.69	0.47	0.40	11.80	424.00
Escondida	Drillhole	MD0509	53.43	53.88	0.45	0.35	3.09	59.00
Escondida	Drillhole	MD0510	15.34	16.00	0.66	0.57	0.28	35.00
Escondida	Drillhole	MD0511	24.97	26.07	1.10	1.07	62.10	35.36
Escondida	Drillhole	MD0512	9.17	9.80	0.63	0.57	0.10	7.00
Escondida	Drillhole	MD0513	7.00	8.20	1.20	1.43	0.12	10.50
Escondida	Drillhole	MD0514	14.10	14.65	0.55	0.49	2.35	108.00
Escondida	Drillhole	MD0515	55.58	55.94	0.36	0.31	0.16	14.00
Escondida	Drillhole	MD0516	34.00	35.05	1.05	0.98	5.08	477.09
Escondida	Drillhole	MD0517	29.65	29.95	0.30	0.24	2.00	91.30
Escondida	Drillhole	MD0518	12.80	14.48	1.68	1.47	0.41	57.18
Escondida	Drillhole	MD0519	44.50	45.50	1.00	0.85	1.62	46.00
Escondida	Drillhole	MD0520	63.80	64.25	0.45	0.47	0.47	50.00
Escondida	Drillhole	MD0521	69.25	70.41	1.16	1.31	11.27	517.83
Escondida	Drillhole	MD0522	68.49	69.24	0.75	0.71	3.89	283.00
Escondida	Drillhole	MD0524	10.72	13.38	2.66	2.27	20.23	272.08
Escondida	Drillhole	MD0525	16.92	17.67	0.75	0.69	0.52	54.00

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Second Preliminary Economic Assessment

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Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Escondida	Drillhole	MD0526	50.00	52.70	2.70	1.74	6.68	106.49
Escondida	Drillhole	MD0527	12.23	13.70	1.47	1.00	2.18	98.08
Escondida	Drillhole	MD0528	25.30	27.50	2.20	1.49	2.23	126.95
Escondida	Drillhole	MD0529	13.76	16.95	3.19	2.24	16.28	359.02
Escondida	Drillhole	MD0530	17.50	20.27	2.77	2.05	28.78	419.63
Escondida	Drillhole	MD0531	46.00	49.10	3.10	2.13	0.81	34.34
Escondida	Drillhole	MD0532	44.83	47.12	2.29	1.36	171.90	417.44
Escondida	Drillhole	MD0533	26.70	28.50	1.80	1.26	12.72	341.36
Escondida	Drillhole	MD0534	11.33	15.00	3.67	2.46	22.53	1103.04
Escondida	Drillhole	MD0535	41.95	43.25	1.30	0.93	46.94	914.58
Escondida	Drillhole	MD0536	6.40	7.55	1.15	0.79	9.03	51.65
Escondida	Drillhole	MD0537	23.40	24.72	1.32	0.94	1.85	42.19
Escondida	Drillhole	MD0538	48.66	49.67	1.01	0.73	190.36	4622.04
Escondida	Drillhole	MD0539	7.00	8.90	1.90	1.21	17.17	180.26
Escondida	Drillhole	MD0540	12.07	13.15	1.08	0.86	1.44	40.41
Escondida	Drillhole	MD0541	53.00	54.26	1.26	0.80	4.54	118.04
Escondida	Drillhole	MD0542	20.45	21.60	1.15	1.11	6.10	350.35
Escondida	Drillhole	MD0543	9.24	11.20	1.96	1.19	2.88	177.59
Escondida	Drillhole	MD0544	22.75	23.93	1.18	0.78	13.82	564.20
Escondida	Drillhole	MD0545	54.90	55.70	0.80	0.57	6.49	41.50
Escondida	Drillhole	MD0546	7.66	8.88	1.22	0.92	7.19	33.01
Escondida	Drillhole	MD0547	8.51	10.00	1.49	0.99	0.60	10.15
Escondida	Drillhole	MD0548	21.30	23.80	2.50	1.75	25.60	584.40
Escondida	Drillhole	MD0549	53.40	56.04	2.64	1.84	39.49	458.14
Escondida	Drillhole	MD0550	12.05	13.30	1.25	0.84	4.42	24.07
Escondida	Drillhole	MD0551	12.06	13.00	0.94	0.67	0.80	15.95
Escondida	Drillhole	MD0552	23.41	24.48	1.07	0.74	3.70	140.54
Escondida	Drillhole	MD0553	38.08	40.77	2.69	1.88	64.23	1546.82
Escondida	Drillhole	MD0554	21.08	22.16	1.08	0.73	11.23	298.43
Escondida	Drillhole	MD0555	60.60	64.03	3.43	2.53	3.65	53.71
Escondida	Drillhole	MD0556	18.26	19.85	1.59	1.06	14.52	639.03
Escondida	Drillhole	MD0557	10.97	12.80	1.83	1.26	14.16	496.31
Escondida	Drillhole	MD0558	51.54	53.20	1.66	1.54	0.16	9.36
Escondida	Drillhole	MD0559	54.25	55.71	1.46	1.35	0.04	3.28
Escondida	Drillhole	MD0560	23.15	25.00	1.85	1.72	1.50	131.37
Escondida	Drillhole	MD0561	40.06	40.79	0.73	0.46	4.24	452.00
Escondida	Drillhole	MD0562	53.12	55.65	2.53	2.13	0.57	63.94
Escondida	Drillhole	MD0563	36.86	42.00	5.14	4.04	6.24	679.99
Escondida	Drillhole	MD0563	52.84	54.22	1.38	1.22	12.45	1067.68
Escondida	Drillhole	MD0564	39.12	39.79	0.67	1.31	0.06	24.00
Escondida	Drillhole	MD0564	39.79	40.48	0.69	0.57	34.40	336.00
Escondida	Drillhole	MD0564	40.48	40.70	0.22	1.31	0.20	11.00
Escondida	Drillhole	MD0565	35.83	36.20	0.37	0.30	1.93	101.00
Escondida	Drillhole	MD0567	7.12	8.06	0.94	0.69	12.78	64.94
Escondida	Drillhole	MD0567	8.06	8.46	0.40	1.15	0.27	34.00

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

Page 242
Part 14

Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Escondida	Drillhole	MD0568	10.03	13.72	3.69	3.01	41.34	296.46
Escondida	Drillhole	MD0569	7.84	10.75	2.91	2.49	53.77	1319.36
Escondida	Drillhole	MD0570	3.30	3.77	0.47	0.40	46.10	618.00
Escondida	Drillhole	MD0571	13.60	15.00	1.40	1.14	0.28	12.84
Escondida	Drillhole	MD0572	141.70	142.67	0.97	0.66	0.23	10.05
Escondida	Drillhole	MD0573	146.00	147.00	1.00	0.62	0.14	12.00
Escondida	Drillhole	MD0574	152.65	154.74	2.09	1.50	0.87	1.70
Escondida	Drillhole	MD0575	232.36	233.57	1.21	0.93	0.45	97.44
Escondida	Drillhole	MD0576	25.00	26.78	1.78	1.55	8.57	450.62
Escondida	Drillhole	MD0577	22.00	23.00	1.00	0.98	0.67	14.60
Escondida	Drillhole	MD0577	33.00	36.00	3.00	2.68	79.32	5197.32
Escondida	Drillhole	MD0578	258.63	259.00	0.37	0.27	2.95	328.00
Escondida	Drillhole	MD0579	197.64	199.19	1.55	1.11	1.49	155.64
Escondida	Drillhole	MD0580	63.03	64.87	1.84	1.23	46.86	1309.30
Escondida	Drillhole	MD0581	15.90	18.90	3.00	1.86	3.52	278.30
Escondida	Drillhole	MD0582	32.50	33.88	1.38	0.85	147.24	1528.19
Escondida	Drillhole	MD0583	47.50	49.30	1.80	1.11	6.32	112.00
Escondida	Drillhole	MD0584	23.25	26.18	2.93	1.95	54.60	1853.56
Escondida	Drillhole	MD0585	18.38	20.94	2.56	1.67	11.86	1213.84
Escondida	Drillhole	MD0586	82.42	84.07	1.65	1.07	8.24	114.32
Escondida	Drillhole	MD0589	38.87	43.80	4.93	3.29	117.80	4578.20
Escondida	Drillhole	MD0590	89.00	90.48	1.48	0.94	5.59	76.15
Escondida	Drillhole	MD0591	144.15	145.54	1.39	1.05	12.63	830.64
Escondida	Drillhole	MD0592	168.79	170.40	1.61	1.22	23.44	1442.21
Escondida	Drillhole	MD0593	214.06	216.26	2.20	1.69	1.55	84.42
Escondida	Drillhole	MD0594	48.30	53.36	5.06	3.21	9.85	1415.44
Escondida	Drillhole	MD0595	94.56	97.84	3.28	2.20	16.88	1092.41
Escondida	Drillhole	MD0596	124.88	128.00	3.12	2.39	43.14	2948.21
Escondida	Drillhole	MD0597	141.45	143.41	1.96	1.23	1.78	121.59
Escondida	Drillhole	MD0598	91.92	94.77	2.85	1.68	38.74	3492.82
Escondida	Drillhole	MD0599	48.80	51.10	2.30	1.50	6.15	847.80
Escondida	Drillhole	MD0601	126.00	130.36	4.36	2.24	86.27	3915.03
Escondida	Drillhole	MD0602	172.20	178.20	6.00	3.73	2.78	112.74
Escondida	Drillhole	MD0603	217.90	223.08	5.18	4.06	5.14	57.33
Escondida	Drillhole	MD0604	208.10	211.00	2.90	1.48	5.72	115.34
Escondida	Drillhole	MD0605	156.12	158.30	2.18	1.24	35.54	1398.77
Escondida	Drillhole	MD0607	105.87	109.15	3.28	2.05	55.27	4013.53
Escondida	Drillhole	MD0608	108.83	111.50	2.67	1.47	12.96	1129.66
Escondida	Drillhole	MD0609	73.48	75.00	1.52	0.97	4.59	322.05
Escondida	Drillhole	MD0611	164.55	166.18	1.63	0.93	6.41	180.12
Escondida	Drillhole	MD0612	215.37	219.08	3.71	2.13	1.17	79.46
Escondida	Drillhole	MD0614	115.00	115.70	0.70	0.40	0.00	21.00
Escondida	Drillhole	MD0615	161.00	163.00	2.00	1.40	0.57	56.50

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

Page 243
Part 13

Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Escondida	Drillhole	MD0617	103.80	108.70	4.90	3.08	119.22	3959.56
Escondida	Drillhole	MD0618	130.45	135.30	4.85	3.01	10.55	1073.84
Escondida	Drillhole	MD0620	179.07	179.97	0.90	0.63	8.81	596.82
Escondida	Drillhole	MD0621	227.56	232.02	4.46	2.90	0.93	24.55
Escondida	Drillhole	MD0623	242.00	245.93	3.93	2.58	0.43	57.29
Escondida	Drillhole	MD0626	210.70	212.00	1.30	0.83	0.27	11.69
Escondida	Drillhole	MD0629	233.30	236.25	2.95	2.20	0.90	11.86
Escondida	Drillhole	MD0630	49.55	50.34	0.79	0.52	0.09	5.01
Escondida	Drillhole	MD0632	95.53	95.83	0.30	0.19	1.07	117.00
Escondida	Drillhole	MD0633	157.97	159.92	1.95	1.43	87.22	5493.37
Escondida	Drillhole	MD0635	109.00	113.12	4.12	2.76	5.85	545.04
Escondida	Drillhole	MD0637	60.73	63.57	2.84	1.80	8.31	445.93
Escondida	Drillhole	MD0638	178.28	180.33	2.05	1.36	1.38	53.35
Escondida	Drillhole	MD0640	51.50	56.00	4.50	2.88	13.02	1230.65
Escondida	Drillhole	MD0642	102.83	105.86	3.03	1.87	3.85	279.23
Escondida	Drillhole	MD0644	70.30	74.10	3.80	2.34	91.01	5648.61
Escondida	Drillhole	MD0646	112.77	119.00	6.23	3.24	54.44	2410.22
Escondida	Drillhole	MD0647	163.00	166.38	3.38	2.13	7.45	525.77
Escondida	Drillhole	MD0650	71.35	72.10	0.75	0.43	36.62	1378.78
Escondida	Drillhole	MD0652	197.53	200.64	3.11	2.34	3.94	229.39
Escondida	Drillhole	MD0653	208.89	210.40	1.51	1.16	7.17	198.21
Escondida	Drillhole	MD0655	231.20	233.30	2.10	1.44	0.61	73.06
Escondida	Drillhole	MD0656	126.70	130.10	3.40	1.81	23.18	1780.46
Escondida	Drillhole	MD0658	254.93	260.40	5.47	3.89	0.71	44.51
Escondida	Drillhole	MD0661	166.16	169.19	3.03	1.66	5.45	384.22
Escondida	Drillhole	MD0662	109.83	112.29	2.46	1.41	16.07	955.18
Escondida	Drillhole	MD0664	197.71	198.53	0.82	0.53	0.07	3.00
Escondida	Drillhole	MD0666	297.64	300.97	3.33	2.08	1.09	34.07
Escondida	Drillhole	MD0667	305.40	308.16	2.76	1.81	0.45	26.15
Escondida	Drillhole	MD0670	160.00	162.13	2.13	1.15	4.48	201.78
Escondida	Drillhole	MD0671	94.66	95.00	0.34	0.20	1.53	954.00
Escondida	Drillhole	MD0672	145.22	146.90	1.68	0.91	9.51	583.50
Escondida	Drillhole	MD0673	140.62	146.60	5.98	3.26	0.77	50.67
Escondida	Drillhole	MD0675	197.00	199.74	2.74	1.58	1.57	140.04
Escondida	Drillhole	MD0676	175.20	182.15	6.95	4.64	16.76	275.97
Escondida	Drillhole	MD0678	131.11	137.65	6.54	3.85	19.43	709.99
Escondida	Drillhole	MD0680	176.93	177.77	0.84	0.57	1.28	98.72
Escondida	Drillhole	MD0681	128.62	130.00	1.38	0.91	1.16	118.97
Escondida	Drillhole	MD0683	80.11	81.30	1.19	0.80	0.95	75.46
Escondida	Drillhole	MD0685	48.19	50.68	2.49	1.52	2.24	280.41
Escondida	Drillhole	MD0687	100.15	103.25	3.10	2.05	0.08	7.01
Escondida	Drillhole	MD0689	254.80	255.10	0.30	0.13	0.18	7.30
Escondida	Drillhole	MD0691	241.10	243.00	1.90	1.35	2.68	76.15
Escondida	Drillhole	MD0694	207.15	209.00	1.85	1.57	0.06	4.97

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

Page 244
Part 14

Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Escondida	Drillhole	MD0698	162.52	164.28	1.76	1.28	82.81	4660.34
Escondida	Drillhole	MD0699	117.85	118.85	1.00	0.87	184.42	1414.08
Escondida	Drillhole	MD0700	124.00	125.45	1.45	1.55	11.43	435.31
Escondida	Drillhole	MD0702	182.75	184.90	2.15	1.52	3.72	673.72
Escondida	Drillhole	MD0703	208.90	210.00	1.10	0.86	2.11	195.18
Escondida	Drillhole	MD0704	131.05	136.25	5.20	3.54	14.96	1121.28
Escondida	Drillhole	MD0706	174.00	175.00	1.00	0.77	0.80	79.40
Escondida	Drillhole	MD0708	77.80	82.00	4.20	2.60	57.02	2953.43
Escondida	Drillhole	MD0709	181.00	182.65	1.65	1.33	8.37	724.48
Escondida	Drillhole	MD0710	27.20	29.20	2.00	1.49	13.17	20.70
Escondida	Drillhole	MD0711	125.95	128.37	2.42	1.55	12.09	335.74
Escondida	Drillhole	MD0713	224.55	224.85	0.30	0.21	1.61	145.00
Escondida	Drillhole	MD0714	199.00	200.20	1.20	0.93	6.91	351.47
Escondida	Drillhole	MD0715	69.10	69.85	0.75	0.64	2.21	97.60
Escondida	Drillhole	MD0717	98.11	103.74	5.63	2.84	14.21	1390.61
Escondida	Drillhole	MD0718	201.05	201.48	0.43	0.29	0.17	16.00
Escondida	Drillhole	MD0720	126.55	128.26	1.71	0.94	4.99	725.63
Escondida	Drillhole	MD0721	154.70	158.50	3.80	2.04	22.58	1049.56
Escondida	Drillhole	MD0722	147.85	149.00	1.15	1.17	0.42	27.31
Escondida	Drillhole	MD0723	130.90	133.90	3.00	1.78	6.52	555.69
Escondida	Drillhole	MD0724	85.00	87.40	2.40	1.70	0.71	91.76
Escondida	Drillhole	MD0725	126.00	128.00	2.00	1.78	5.34	822.50
Escondida	Drillhole	MD0727	182.45	184.20	1.75	1.00	11.92	1471.29
Escondida	Drillhole	MD0728	114.00	114.70	0.70	0.48	30.99	3408.58
Escondida	Drillhole	MD0730	174.69	176.28	1.59	1.27	3.87	485.55
Escondida	Drillhole	MD0732	119.20	119.71	0.51	0.27	2.97	18.00
Escondida	Drillhole	MD0733	207.85	209.00	1.15	0.75	4.43	332.83
Escondida	Drillhole	MD0734	102.13	102.73	0.60	0.56	2.61	211.50
Escondida	Drillhole	MD0735	160.00	161.00	1.00	0.84	0.15	7.10
Escondida	Drillhole	MD0736	169.12	171.25	2.13	1.30	3.23	207.97
Escondida	Drillhole	MD0737	181.40	182.85	1.45	1.03	1.74	5.07
Escondida	Drillhole	MD0738	98.26	98.64	0.38	0.35	0.62	46.00
Escondida	Drillhole	MD0739	211.60	217.50	5.90	3.51	0.83	97.61
Escondida	Drillhole	MD0740	149.54	149.91	0.37	0.28	1.43	10.00
Escondida	Drillhole	MD0741	149.87	150.16	0.29	0.18	0.02	3.00
Escondida	Drillhole	MD0742	249.28	250.08	0.80	0.39	0.62	21.70
Escondida	Drillhole	MD0743	146.32	148.50	2.18	1.87	1.99	265.81
Escondida	Drillhole	MD0744	106.12	107.70	1.58	0.98	0.36	8.51
Escondida	Drillhole	MD0745	121.00	122.46	1.46	1.19	1.11	135.44
Escondida	Drillhole	MD0746	211.95	213.08	1.13	0.58	0.46	8.66
Escondida	Drillhole	MD0748	65.80	67.76	1.96	1.39	1.76	88.92
Escondida	Drillhole	MD0750	90.00	91.00	1.00	0.89	4.79	835.92
Escondida	Drillhole	MD0751	62.42	64.00	1.58	1.15	32.32	376.08
Escondida	Drillhole	MD0752	94.35	97.03	2.68	1.96	38.88	172.73

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

Page 245
Part 14

Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Escondida	Drillhole	MD0753	90.33	95.90	5.57	3.12	158.43	6446.76
Escondida	Drillhole	MD0754	81.37	82.02	0.65	0.55	0.35	30.00
Escondida	Drillhole	MD0755	71.45	72.22	0.77	0.44	0.29	8.44
Escondida	Drillhole	MD0756	140.53	143.52	2.99	1.78	64.81	3485.99
Escondida	Drillhole	MD0757	80.80	82.00	1.20	1.10	45.29	805.50
Escondida	Drillhole	MD0758	131.16	131.76	0.60	0.50	3.15	237.49
Escondida	Drillhole	MD0759	91.30	92.23	0.93	0.63	10.24	27.84
Escondida	Drillhole	MD0760	46.40	51.50	5.10	2.86	68.41	3913.33
Escondida	Drillhole	MD0761	142.35	143.11	0.76	0.64	2.46	135.92
Escondida	Drillhole	MD0762	87.97	89.17	1.20	0.76	55.16	3504.09
Escondida	Drillhole	MD0763	84.14	84.65	0.51	0.43	0.18	13.00
Escondida	Drillhole	MD0764	113.30	117.00	3.70	2.87	16.94	615.80
Escondida	Drillhole	MD0765	125.50	126.50	1.00	0.64	2.20	361.80
Escondida	Drillhole	MD0766	32.60	33.86	1.26	0.92	0.43	21.13
Escondida	Drillhole	MD0767	54.12	55.00	0.88	0.72	0.06	4.79
Escondida	Drillhole	MD0768	207.00	210.58	3.58	2.39	31.28	652.57
Escondida	Drillhole	MD0769	151.00	153.70	2.70	2.29	16.71	1208.54
Escondida	Drillhole	MD0770	9.73	10.35	0.62	0.45	1.10	61.00
Escondida	Drillhole	MD0771	57.00	58.70	1.70	1.17	2.00	32.52
Escondida	Drillhole	MD0772	155.57	162.84	7.27	4.55	14.49	744.95
Escondida	Drillhole	MD0773	92.00	93.25	1.25	1.13	10.40	1081.40
Escondida	Drillhole	MD0775	67.83	69.47	1.64	1.17	4.82	38.99
Escondida	Drillhole	MD0776	115.42	116.00	0.58	0.58	2.15	322.93
Escondida	Drillhole	MD0777	231.19	233.82	2.63	1.87	70.00	2180.83
Escondida	Drillhole	MD0778	153.38	156.00	2.62	1.63	0.55	43.00
Escondida	Drillhole	MD0779	353.87	356.15	2.28	1.63	0.28	14.57
Escondida	Drillhole	MD0780	324.84	327.23	2.39	1.52	0.33	33.95
Escondida	Drillhole	MD0781	330.95	334.10	3.15	2.12	0.29	17.02
Escondida	Drillhole	MD0782	124.73	125.70	0.97	0.61	0.40	40.18
Escondida	Drillhole	MD0783	95.73	96.06	0.33	0.21	1.21	38.00
Escondida	Drillhole	MD0784	264.80	267.55	2.75	1.79	8.37	298.76
Escondida	Drillhole	MD0785	99.05	100.55	1.50	0.93	1.61	361.33
Escondida	Drillhole	MD0786	251.92	254.55	2.63	1.80	9.54	404.92
Escondida	Drillhole	MD0787	113.75	115.35	1.60	1.12	0.46	26.78
Escondida	Drillhole	MD0788	251.25	253.90	2.65	1.70	1.74	75.08
Escondida	Drillhole	MD0789	95.20	96.35	1.15	0.77	1.45	25.70
Escondida	Drillhole	MD0790	21.00	21.62	0.62	0.53	76.60	292.00
Escondida	Drillhole	MD0791	75.26	77.12	1.86	1.27	0.75	12.77
Escondida	Drillhole	MD0792	73.79	74.57	0.78	0.51	27.80	29.50
Escondida	Drillhole	MD0793	58.55	60.20	1.65	1.38	2.86	49.55
Escondida	Drillhole	MD0794	44.60	45.00	0.40	0.19	1.02	74.00
Escondida	Drillhole	MD0795	107.45	110.00	2.55	1.57	3.77	352.90
Escondida	Drillhole	MD0797	78.42	79.14	0.72	0.57	84.67	1790.15
Escondida	Drillhole	MD0798	138.00	139.60	1.60	1.12	0.28	9.78

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

Page 246
Part 14

Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Escondida	Drillhole	MD0799	73.00	75.00	2.00	1.63	17.31	2224.66
Escondida	Drillhole	MD0800	165.25	166.85	1.60	1.12	0.12	13.00
Escondida	Drillhole	MD0801	134.65	137.00	2.35	1.65	11.58	54.15
Escondida	Drillhole	MD0802	135.58	136.00	0.42	0.39	5.92	851.00
Escondida	Drillhole	MD0803	188.25	188.90	0.65	0.47	11.60	850.00
Escondida	Drillhole	MD0804	113.60	115.00	1.40	0.95	2.04	510.17
Escondida	Drillhole	MD0805	205.30	207.60	2.30	1.53	0.65	35.70
Escondida	Drillhole	MD0806	63.45	65.50	2.05	1.64	0.96	31.00
Escondida	Drillhole	MD0807	41.28	43.00	1.72	1.43	0.47	15.73
Escondida	Drillhole	MD0808	42.31	44.27	1.96	1.73	0.63	35.47
Escondida	Drillhole	MD0809A	15.15	17.90	2.75	2.25	0.18	14.47
Escondida	Drillhole	MD0810	82.80	87.20	4.40	3.12	10.33	431.97
Escondida	Drillhole	MD0811	65.64	66.06	0.42	0.36	0.17	16.00
Escondida	Drillhole	MD0812	115.30	116.40	1.10	0.78	9.92	377.82
Escondida	Drillhole	MD0813	101.30	105.45	4.15	2.40	0.32	28.52
Escondida	Drillhole	MD0814	66.09	66.65	0.56	0.46	0.14	3.98
Escondida	Drillhole	MD0815	107.68	108.00	0.32	0.27	0.48	8.47
Escondida	Drillhole	MD0816	120.32	124.70	4.38	2.67	1.31	123.66
Escondida	Drillhole	MD0817	174.00	177.03	3.03	2.30	12.48	467.34
Escondida	Drillhole	MD0818	94.00	94.42	0.42	0.34	4.91	305.00
Escondida	Drillhole	MD0819	152.30	154.23	1.93	1.64	8.72	1075.86
Escondida	Drillhole	MD0820	105.20	107.00	1.80	1.71	7.62	981.29
Escondida	Drillhole	MD0821	171.68	173.73	2.05	1.15	0.19	29.68
Escondida	Drillhole	MD0822	82.52	83.20	0.68	0.69	4.02	118.90
Escondida	Drillhole	MD0823	205.92	206.57	0.65	0.49	2.40	93.08
Escondida	Drillhole	MD0824	83.21	86.96	3.75	1.91	2.56	127.17
Escondida	Drillhole	MD0826	88.10	89.00	0.90	0.71	1.80	29.00
Escondida	Drillhole	MD0827	105.81	107.40	1.59	1.40	193.15	5314.42
Escondida	Drillhole	MD0828	193.70	197.55	3.85	2.41	4.19	217.85
Escondida	Drillhole	MD0829	50.81	51.31	0.50	0.27	0.48	15.74
Escondida	Drillhole	MD0830	185.30	189.75	4.45	3.65	5.08	402.04
Escondida	Drillhole	MD0831	201.80	202.60	0.80	0.63	0.59	11.25
Escondida	Drillhole	MD0832	221.20	226.48	5.28	2.70	0.28	42.86
Escondida	Drillhole	MD0833	222.00	223.00	1.00	0.66	0.13	4.92
Escondida	Drillhole	MD0890	29.00	29.50	0.50	0.28	0.32	15.00
Escondida	Drillhole	MD0892	45.28	46.69	1.41	0.87	4.26	396.10
Escondida	Drillhole	MD0893	50.26	51.07	0.81	0.46	0.49	15.98
Escondida	Drillhole	MD0894	86.00	87.05	1.05	0.58	0.58	46.48
Escondida	Drillhole	MD0897	64.15	67.35	3.20	2.31	39.05	2145.02
Escondida	Drillhole	MD0906	53.55	54.45	0.90	0.59	0.35	31.33
Escondida	Drillhole	MD0907	102.40	104.67	2.27	1.53	0.47	34.54
Escondida	Drillhole	MD0911	81.30	83.20	1.90	1.26	0.30	18.71
Escondida	Drillhole	MD0930	155.90	157.00	1.10	0.64	0.27	8.45
Escondida	Drillhole	MD0934	158.25	159.05	0.80	0.54	0.12	19.38

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Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

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Part 14

Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Escondida	Drillhole	MD0936	101.82	105.00	3.18	2.56	3.82	23.70
Escondida	Drillhole	MD0937	112.00	113.00	1.00	0.83	23.11	696.19
Escondida	Drillhole	MD0940	178.60	180.00	1.40	1.09	15.71	2045.43
Escondida	Drillhole	MD0942	162.45	163.48	1.03	1.07	0.82	19.92
Escondida	Drillhole	MD0945	174.90	176.72	1.82	1.67	0.85	65.57
Escondida	Drillhole	MD0947	206.35	209.60	3.25	2.25	10.44	788.20
Escondida	Drillhole	MD0951	201.00	203.80	2.80	2.28	4.15	417.57
Escondida	Drillhole	MD0953	304.49	306.00	1.51	1.09	98.41	26.48
Escondida	Drillhole	MD0956	292.70	295.93	3.23	2.32	1.15	15.48
Escondida	Drillhole	MD0961	305.85	307.75	1.90	0.99	0.26	36.99
Escondida	Drillhole	MD0962	395.70	398.60	2.90	1.87	0.16	8.25
Escondida	Drillhole	MD0966	233.90	238.40	4.50	2.95	0.31	14.88
Escondida	Drillhole	MD0967	291.00	293.50	2.50	1.66	2.19	8.64
Escondida	Drillhole	MD0970	309.23	310.64	1.41	0.85	0.89	27.60
Escondida	Drillhole	MD0972	310.60	312.85	2.25	1.59	0.29	18.71
Escondida	Drillhole	MD0975	174.30	176.30	2.00	1.33	0.76	35.00
Escondida	Drillhole	MD1005	339.10	340.28	1.18	1.11	0.87	59.37
Escondida	Drillhole	MD1017	315.75	317.02	1.27	0.85	2.04	10.68
Escondida	Drillhole	MD1023	333.45	336.00	2.55	1.49	0.22	24.09
Escondida	Drillhole	MRC0162	52.00	53.00	1.00	0.67	0.10	11.25
Escondida	Drillhole	MRC0176	13.00	16.00	3.00	1.64	0.41	6.47
Escondida	Drillhole	MRC0178	126.66	128.84	2.18	1.32	5.11	247.48
Escondida	Drillhole	MRC0253	72.00	73.00	1.00	0.78	0.83	8.63
Escondida	Drillhole	MRC0254	94.00	96.00	2.00	1.14	15.73	400.50
Escondida	Drillhole	MRC034	75.95	77.01	1.07	0.87	12.85	104.00
Loma Esc.	Drillhole	MD0092	18.90	20.30	1.40	1.02	51.23	1678.52
Loma Esc.	Drillhole	MD0093	15.25	16.39	1.14	0.77	1.08	30.24
Loma Esc.	Drillhole	MD0094	12.49	12.79	0.30	0.25	4.82	348.00
Loma Esc.	Drillhole	MD0130	27.41	28.72	1.31	0.95	77.86	4152.55
Loma Esc.	Drillhole	MD0131	21.00	23.00	2.00	1.51	4.32	304.92
Loma Esc.	Drillhole	MD0132	29.26	29.56	0.30	0.19	0.13	14.80
Loma Esc.	Drillhole	MD0133	58.00	59.00	1.00	0.84	1.74	107.00
Loma Esc.	Drillhole	MD0189	53.00	54.54	1.54	1.13	15.83	989.05
Loma Esc.	Drillhole	MD0191	54.40	54.70	0.30	0.22	36.20	1270.00
Loma Esc.	Drillhole	MD0195	40.54	41.28	0.74	0.55	57.46	5226.29
Loma Esc.	Drillhole	MD0197	21.00	22.02	1.02	0.82	11.65	911.37
Loma Esc.	Drillhole	MD0199	36.50	37.60	1.10	0.95	0.08	6.28
Loma Esc.	Drillhole	MD0239	85.78	86.43	0.65	0.65	0.46	41.10
Loma Esc.	Drillhole	MD0243	95.72	96.40	0.68	0.55	35.11	3252.35
Loma Esc.	Drillhole	MD0245	90.57	90.87	0.30	0.28	15.98	652.00
Loma Esc.	Drillhole	MD0300	40.60	41.55	0.95	0.77	8.78	931.84
Loma Esc.	Drillhole	MD0305	144.90	145.48	0.58	0.47	0.93	31.20
Loma Esc.	Drillhole	MD0306	89.50	89.80	0.30	0.21	1.80	103.00
Loma Esc.	Drillhole	MD0308	83.90	84.20	0.30	0.17	0.13	0.53

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

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Part 14

Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Loma Esc.	Drillhole	MD1029	18.65	19.18	0.53	0.42	2.32	255.00
Loma Esc.	Drillhole	MD1030	38.56	38.90	0.34	0.24	74.50	4685.00
Loma Esc.	Drillhole	MD1032	18.43	19.50	1.07	0.79	9.46	711.00
Loma Esc.	Drillhole	MD1034	37.45	37.76	0.31	0.22	13.70	1291.00
Loma Esc.	Drillhole	MD1036	63.70	64.12	0.42	0.32	145.50	4770.00
Loma Esc.	Drillhole	MD1038	21.15	22.25	1.10	0.72	45.16	3219.63
Loma Esc.	Drillhole	MD1040	42.42	42.72	0.30	0.22	20.30	1325.00
Loma Esc.	Drillhole	MD1042	65.10	65.47	0.37	0.29	41.70	1340.00
Loma Esc.	Drillhole	MD1044	5.90	6.25	0.35	0.20	3.15	157.00
Loma Esc.	Drillhole	MD1046	37.23	37.63	0.40	0.25	74.20	5963.00
Loma Esc.	Drillhole	MD1048	61.00	62.00	1.00	0.85	15.11	663.20
Loma Esc.	Drillhole	MD1054	1.14	2.02	0.88	0.52	0.17	4.37
Loma Esc.	Drillhole	MD1055	60.00	61.00	1.00	0.89	0.89	69.20
Loma Esc.	Drillhole	MD1057	59.00	59.53	0.53	0.46	0.39	23.00
Loma Esc.	Drillhole	MD1061	37.60	38.44	0.84	0.56	3.25	320.83
Loma Esc.	Drillhole	MD1063	58.80	59.73	0.93	0.76	1.30	165.57
Loma Esc.	Drillhole	MD1065	37.00	37.90	0.90	0.60	19.68	1165.11
Loma Esc.	Drillhole	MD1068	0.00	0.70	0.70	0.35	0.82	22.90
Loma Esc.	Drillhole	MD1069	5.00	6.00	1.00	0.60	15.50	556.71
Loma Esc.	Drillhole	MD1071	59.70	60.06	0.36	0.28	0.90	29.00
Loma Esc.	Drillhole	MD1073	16.47	18.46	1.99	1.24	19.51	1556.69
Loma Esc.	Drillhole	MD1076	57.60	58.25	0.65	0.56	0.10	10.46
Loma Esc.	Drillhole	MD1079	39.78	40.55	0.77	0.62	4.68	294.00
Loma Esc.	Drillhole	MD1082	13.00	16.50	3.50	2.06	9.56	538.86
Loma Esc.	Drillhole	MD1084	39.17	39.95	0.78	0.66	26.12	1420.25
Loma Esc.	Drillhole	MD1085	36.67	37.40	0.73	0.61	0.34	6.45
Loma Esc.	Drillhole	MD1087	31.73	32.50	0.77	0.68	105.26	1591.64
Loma Esc.	Drillhole	MD1089	41.80	42.20	0.40	0.32	9.65	905.00
Loma Esc.	Drillhole	MD1091	18.50	18.85	0.35	0.26	17.30	1184.00
Loma Esc.	Drillhole	MD1094	16.90	17.50	0.60	0.53	0.15	48.00
Loma Esc.	Drillhole	MD1095	40.00	41.00	1.00	0.77	0.44	5.50
Loma Esc.	Drillhole	MD1097	13.95	14.45	0.50	0.35	3.21	41.00
Loma Esc.	Drillhole	MD1141	60.21	60.51	0.30	0.23	0.16	15.00
Loma Esc.	Drillhole	MD1142	53.45	53.80	0.35	0.29	1.03	85.00
Loma Esc.	Drillhole	MD1144	23.82	24.62	0.80	0.65	0.59	17.90
Loma Esc.	Trench	RxLE_03	1.05	2.70	1.65	1.65	2.88	182.09
Loma Esc.	Trench	TLE_07	5.80	6.10	0.30	0.30	1.50	41.20
Loma Esc.	Trench	TLE_08	14.60	16.00	1.40	1.40	0.92	10.96
Loma Esc.	Trench	TLE_09	6.50	6.90	0.40	0.40	0.42	61.50
Loma Esc.	Trench	TLE0011	6.40	7.00	0.60	0.60	0.50	8.00
Loma Esc.	Trench	TLE0012	10.50	13.10	2.60	2.60	3.64	22.69
Loma Esc.	Trench	TLE0013	9.60	10.50	0.90	0.90	0.85	90.00
Loma Esc.	Trench	TLE0014	4.30	5.10	0.80	0.80	0.86	24.00
Loma Esc.	Trench	TLE0015	9.00	10.40	1.40	1.40	7.59	169.00

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

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Part 14

Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Loma Esc.	Trench	TLE0016	2.30	3.10	0.80	0.80	1.04	25.00
Loma Esc.	Trench	TLE0019	2.00	3.00	1.00	1.00	96.60	3710.00
Loma Esc.	Trench	TLE0020	8.50	9.00	0.50	0.50	8.63	246.00
Loma Esc.	Trench	TLE0021	5.70	6.70	1.00	1.00	15.57	140.00
Loma Esc.	Trench	TLE0022	10.40	11.90	1.50	1.50	120.79	5445.67
Loma Esc.	Trench	TLE0023	11.00	12.50	1.50	1.50	23.58	1003.00
Loma Esc.	Trench	TLE0024	5.60	6.60	1.00	1.00	30.10	553.50
Loma Esc.	Trench	TLE0025	6.40	6.90	0.50	0.50	0.31	4.00
Loma Esc.	Trench	TLE0026	5.60	6.20	0.60	0.60	4.81	11.00
Loma Esc.	Trench	TLE0027	8.50	9.00	0.50	0.50	10.60	35.70
Loma Esc.	Trench	TLE0028	3.50	4.50	1.00	1.00	4.07	109.50
Loma Esc.	Trench	TLE0029	14.15	15.15	1.00	1.00	4.46	196.00
Gabriela	Drillhole	MD0163	32.00	35.92	3.92	3.63	2.24	379.05
Gabriela	Drillhole	MD0164	41.70	42.60	0.90	0.81	1.26	242.31
Gabriela	Drillhole	MD0179	29.58	32.00	2.42	1.58	2.23	286.34
Gabriela	Drillhole	MD0181	82.00	84.83	2.83	1.83	2.59	480.46
Gabriela	Drillhole	MD0183	21.00	24.10	3.10	2.88	3.28	582.02
Gabriela	Drillhole	MD0184	23.00	24.60	1.60	1.58	1.20	175.94
Gabriela	Drillhole	MD0295	31.00	34.85	3.85	3.93	1.77	142.08
Gabriela	Drillhole	MD0318	33.77	36.74	2.97	1.70	2.74	360.40
Gabriela	Drillhole	MD0321	55.50	59.78	4.28	2.43	3.88	507.25
Gabriela	Drillhole	MD0322	252.05	254.55	2.50	1.40	1.71	342.72
Gabriela	Drillhole	MD0323	55.17	57.78	2.61	1.68	3.43	419.74
Gabriela	Drillhole	MD0324	166.64	168.72	2.08	1.23	1.78	305.17
Gabriela	Drillhole	MD0327	185.58	187.87	2.29	1.41	1.69	267.51
Gabriela	Drillhole	MD0328	162.25	163.75	1.50	1.17	1.13	272.53
Gabriela	Drillhole	MD0337	176.30	177.24	0.94	0.53	2.23	410.89
Gabriela	Drillhole	MD0338A	204.15	206.25	2.10	1.51	0.35	84.54
Gabriela	Drillhole	MD0357	141.13	143.00	1.87	1.43	0.47	218.99
Gabriela	Drillhole	MD0362	129.73	133.00	3.27	2.24	2.71	305.83
Gabriela	Drillhole	MD0364	110.95	113.79	2.84	1.89	2.96	473.51
Gabriela	Drillhole	MD0366	115.44	117.34	1.90	1.08	1.34	203.27
Gabriela	Drillhole	MD0368	158.56	162.90	4.34	2.45	3.10	502.36
Gabriela	Drillhole	MD0371	74.85	79.95	5.10	3.84	9.25	1446.12
Gabriela	Drillhole	MD0416	48.00	48.70	0.70	0.63	0.75	97.10
Gabriela	Drillhole	MD0420	160.20	162.00	1.80	0.93	0.15	32.34
Gabriela	Drillhole	MD0856	75.70	77.00	1.30	0.94	0.64	106.91
Gabriela	Drillhole	MD0858	77.30	78.10	0.80	0.61	2.85	489.00
Gabriela	Drillhole	MD0859	184.55	188.70	4.15	2.91	1.37	205.64
Gabriela	Drillhole	MD0860	117.25	117.65	0.40	0.35	0.64	115.00
Gabriela	Drillhole	MD0863	94.66	95.60	0.94	0.74	0.71	96.57
Gabriela	Drillhole	MD0866	85.35	87.00	1.65	1.34	0.99	195.62
Gabriela	Drillhole	MD0867	38.89	39.70	0.81	0.65	0.11	27.37
Gabriela	Drillhole	MD0868	54.89	58.00	3.11	3.17	5.63	1015.25

Extorre Gold Mines Limited
Cerro Moro Gold-Silver Project
Second Preliminary Economic Assessment

Page 250
Part 14

Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Gabriela	Drillhole	MD0871	121.00	126.67	5.67	3.69	1.27	220.60
Gabriela	Drillhole	MD0874	72.94	75.72	2.78	1.97	3.39	492.83
Gabriela	Drillhole	MD0875	126.45	128.80	2.35	1.50	6.88	986.20
Gabriela	Drillhole	MD0876	133.70	135.83	2.13	1.30	1.27	200.24
Gabriela	Drillhole	MD0878	127.75	132.00	4.25	2.86	1.97	373.45
Gabriela	Drillhole	MD0882	188.78	190.63	1.85	1.20	2.06	279.08
Gabriela	Drillhole	MD0884	120.08	121.33	1.25	0.92	1.44	229.74
Gabriela	Drillhole	MD0916	120.48	125.80	5.32	3.32	1.91	391.29
Gabriela	Drillhole	MD0922	113.55	118.00	4.45	3.54	2.07	300.85
Gabriela	Drillhole	MD0928	232.70	235.59	2.89	1.39	7.44	1216.01
Gabriela	Drillhole	MD0933	203.07	206.13	3.06	1.80	2.45	441.14
Gabriela	Drillhole	MD0938	163.20	165.90	2.70	1.54	2.65	426.83
Gabriela	Drillhole	MD0944	175.90	185.05	9.15	5.24	2.14	346.09
Gabriela	Drillhole	MD0948	202.00	206.00	4.00	1.94	2.14	309.40
Gabriela	Drillhole	MD0952A	287.00	287.50	0.50	0.26	3.07	624.00
Gabriela	Drillhole	MD0992	200.15	201.75	1.60	1.09	1.54	348.75
Gabriela	Drillhole	MD0996	204.65	207.30	2.65	1.65	2.35	375.74
Gabriela	Drillhole	MD1002	175.31	176.33	1.02	0.76	0.45	79.00
Gabriela	Drillhole	MD1007	140.61	141.50	0.89	0.78	1.04	179.05
Gabriela	Drillhole	MD1019	31.88	32.34	0.46	0.30	1.10	131.38
Gabriela	Drillhole	MD1020	50.47	51.20	0.73	0.57	1.13	172.00
Gabriela	Drillhole	MD1021	20.90	21.76	0.86	0.48	0.80	119.00
Gabriela	Drillhole	MD1022	28.15	31.90	3.75	2.26	4.98	582.49
Gabriela	Drillhole	MD1026	26.79	27.43	0.64	0.38	2.55	366.00
Gabriela	Drillhole	MD1028	47.15	48.15	1.00	0.64	3.66	457.50
Gabriela	Drillhole	MD1031	52.70	53.50	0.80	0.49	5.12	771.00
Gabriela	Drillhole	MD1033	26.00	27.26	1.26	0.76	0.58	68.51
Gabriela	Drillhole	MD1035	50.20	52.85	2.65	1.61	1.07	115.02
Gabriela	Drillhole	MD1037	33.78	35.40	1.62	1.00	4.58	605.91
Gabriela	Drillhole	MD1039	64.00	65.13	1.13	0.69	0.34	61.83
Gabriela	Drillhole	MD1041	87.00	88.23	1.23	0.84	0.82	136.81
Gabriela	Drillhole	MD1043	30.70	32.73	2.03	1.34	2.61	397.50
Gabriela	Drillhole	MD1047	84.30	86.65	2.35	1.62	5.10	764.60
Gabriela	Drillhole	MD1050	38.90	42.00	3.10	1.81	3.65	613.77
Gabriela	Drillhole	MD1051	66.45	71.19	4.74	2.77	1.93	318.93
Gabriela	Drillhole	MD1052	174.00	178.46	4.46	2.67	10.26	1439.70
Gabriela	Drillhole	MD1053	90.65	97.45	6.80	4.37	1.89	218.40
Gabriela	Drillhole	MD1056	38.67	48.00	9.33	4.67	1.61	234.50
Gabriela	Drillhole	MD1058	64.50	69.10	4.60	3.02	1.27	194.30
Gabriela	Drillhole	MD1059	20.00	23.20	3.20	1.87	1.79	141.63
Gabriela	Drillhole	MD1060	218.30	221.40	3.10	1.88	2.57	434.36
Gabriela	Drillhole	MD1062	46.37	49.40	3.03	1.65	5.87	510.86
Gabriela	Drillhole	MD1064	70.41	72.21	1.80	1.08	3.02	391.00
Gabriela	Drillhole	MD1066	61.00	64.40	3.40	1.88	17.36	2222.54

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Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Gabriela	Drillhole	MD1067	153.15	155.20	2.05	1.18	1.69	290.69
Gabriela	Drillhole	MD1070	116.00	118.00	2.00	1.54	0.92	120.50
Gabriela	Drillhole	MD1072	69.38	72.00	2.62	1.57	3.66	533.53
Gabriela	Drillhole	MD1074	47.00	49.07	2.07	1.23	0.71	104.74
Gabriela	Drillhole	MD1075	146.10	147.30	1.20	0.62	1.35	183.00
Gabriela	Drillhole	MD1077	9.00	20.90	11.90	5.66	1.78	232.68
Gabriela	Drillhole	MD1078	46.00	52.20	6.20	5.94	2.28	256.68
Gabriela	Drillhole	MD1080	333.00	334.00	1.00	0.51	1.22	285.00
Gabriela	Drillhole	MD1081	24.00	34.50	10.50	8.54	2.11	258.12
Gabriela	Drillhole	MD1083	54.50	58.92	4.42	3.24	2.26	496.15
Gabriela	Drillhole	MD1086	47.60	52.00	4.40	3.91	5.10	950.92
Gabriela	Drillhole	MD1088	75.00	75.60	0.60	0.45	0.04	7.00
Gabriela	Drillhole	MD1090	53.00	53.60	0.60	0.47	0.52	75.00
Gabriela	Drillhole	MD1092	28.00	28.90	0.90	0.67	0.22	53.00
Gabriela	Drillhole	MD1093	272.60	273.50	0.90	0.51	0.72	210.00
Gabriela	Drillhole	MD1096	90.25	92.05	1.80	1.23	0.20	26.67
Gabriela	Drillhole	MD1098	102.00	103.80	1.80	1.24	1.95	274.89
Gabriela	Drillhole	MD1100	95.00	97.55	2.55	1.77	1.79	189.92
Gabriela	Drillhole	MD1102	150.40	151.64	1.24	0.78	0.87	155.03
Gabriela	Drillhole	MD1103	151.00	153.06	2.06	1.19	0.88	284.16
Gabriela	Drillhole	MD1106	112.80	114.58	1.78	1.26	0.63	78.13
Gabriela	Drillhole	MD1107	156.29	158.29	2.00	1.37	0.60	101.96
Gabriela	Drillhole	MD1110	150.85	152.00	1.15	0.76	0.22	36.48
Gabriela	Drillhole	MD1114	245.20	247.50	2.30	1.36	4.73	760.26
Gabriela	Drillhole	MD1119	266.00	270.00	4.00	2.13	0.61	103.34
Gabriela	Drillhole	MD1124	224.90	226.00	1.10	0.65	1.34	342.37
Gabriela	Drillhole	MD1130	218.00	219.00	1.00	0.52	0.03	5.50
Gabriela	Drillhole	MD1135	121.18	122.10	0.92	0.71	0.82	196.78
Gabriela	Drillhole	MD1137	170.10	171.00	0.90	0.60	0.92	139.53
Gabriela	Drillhole	MD1154	226.75	229.65	2.90	1.91	0.85	156.68
Gabriela	Drillhole	MRC0264	39.00	46.00	7.00	4.44	1.77	200.09
Gabriela	Drillhole	MRC0266	54.00	56.00	2.00	1.65	2.39	317.00
Gabriela	Drillhole	MRC0267	48.00	50.00	2.00	1.57	0.58	79.55
Gabriela	Drillhole	MRC0329	71.00	72.00	1.00	0.92	0.05	2.62
Gabriela	Drillhole	MRC0372	79.00	81.00	2.00	2.10	5.19	948.00
Gabriela	Trench	TG0001	12.30	16.93	4.63	4.69	2.16	90.55
Gabriela	Trench	TG0004	6.90	7.80	0.90	0.32	0.85	54.00
Esperanza	Drillhole	DDH11	50.50	58.20	7.70	4.54	3.05	362.82
Esperanza	Drillhole	DDH12	47.80	48.50	0.70	0.56	0.15	35.00
Esperanza	Drillhole	MD076	42.58	45.60	3.02	2.25	5.62	500.00
Esperanza	Drillhole	MD077	33.23	34.89	1.66	1.44	1.09	122.77
Esperanza	Drillhole	MD099	35.33	38.87	3.54	2.49	3.34	190.29
Esperanza	Drillhole	MD100	38.00	40.25	2.25	1.29	9.86	283.08
Esperanza	Drillhole	MD101	35.00	38.46	3.46	2.02	10.00	228.80

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Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Esperanza	Drillhole	MD102	39.77	40.77	1.00	0.53	1.46	121.00
Esperanza	Drillhole	MD103	9.61	10.55	0.94	0.54	10.00	500.00
Esperanza	Drillhole	MD126	104.00	106.67	2.67	1.86	1.01	69.63
Esperanza	Drillhole	MD128	53.69	56.53	2.84	1.79	0.29	149.91
Esperanza	Drillhole	MD129	41.10	45.00	3.90	2.44	2.24	52.05
Esperanza	Drillhole	MD134	107.30	108.50	1.20	0.85	1.21	48.80
Esperanza	Drillhole	MD138	54.57	55.70	1.13	0.81	0.82	102.05
Esperanza	Drillhole	MD141	52.85	54.83	1.98	1.47	6.12	500.00
Esperanza	Drillhole	MD143	34.90	35.30	0.40	0.27	3.05	200.00
Esperanza	Drillhole	MD165	37.00	38.00	1.00	0.58	0.65	4.60
Esperanza	Drillhole	MD166	47.30	48.20	0.90	0.57	0.34	7.30
Esperanza	Drillhole	MD294	35.80	40.00	4.20	2.48	4.21	107.66
Esperanza	Drillhole	MD315	121.86	122.91	1.05	0.84	4.18	364.34
Esperanza	Drillhole	MD330	119.00	121.00	2.00	1.35	0.01	1.11
Esperanza	Drillhole	MD333	103.00	103.95	0.95	0.74	0.10	9.21
Esperanza	Drillhole	MD334	110.97	112.25	1.28	0.82	0.95	56.13
Esperanza	Drillhole	MD336	95.33	95.72	0.39	0.46	10.00	100.00
Esperanza	Drillhole	MRC035	24.00	27.00	3.00	1.34	1.39	97.00
Esperanza	Drillhole	MRC036	54.00	60.00	6.00	3.63	0.18	22.00
Esperanza	Drillhole	MRC037	108.00	114.00	6.00	3.02	0.17	32.00
Esperanza	Drillhole	MRC038	33.00	39.00	6.00	4.76	2.15	158.50
Esperanza	Drillhole	MRC039	15.00	18.00	3.00	2.06	1.96	92.00
Esperanza	Drillhole	MRC056	6.00	8.00	2.00	1.58	2.01	286.00
Esperanza	Drillhole	MRC057	38.00	40.00	2.00	1.36	5.79	466.50
Esperanza	Drillhole	MRC059	20.00	23.00	3.00	1.97	4.28	354.33
Esperanza	Drillhole	MRC060	15.00	18.00	3.00	2.05	1.01	115.33
Esperanza	Drillhole	MRC244	50.00	55.00	5.00	3.07	0.51	9.08
Deborah	Drillhole	DDH16	37.70	44.90	7.20	8.51	2.23	26.31
Deborah	Drillhole	DDH17	30.10	32.30	2.20	2.74	1.25	81.16
Deborah	Drillhole	MD063	80.13	84.13	4.00	3.19	1.30	37.98
Deborah	Drillhole	MD078	99.62	100.66	1.04	1.42	1.77	32.70
Deborah	Drillhole	MD079	87.57	88.50	0.93	1.17	1.01	5.30
Deborah	Drillhole	MD080	79.86	80.75	0.89	1.25	1.12	27.10
Deborah	Drillhole	MRC023	12.00	15.00	3.00	3.41	0.75	16.00
Deborah	Drillhole	MRC024	27.00	33.00	6.00	6.74	1.33	23.50
Deborah	Drillhole	MRC025	21.00	27.00	6.00	7.10	1.58	32.00
Deborah	Drillhole	MRC026	69.00	72.00	3.00	3.85	2.15	40.00
Deborah	Drillhole	MRC027	27.00	33.00	6.00	7.41	2.61	66.50
Deborah	Drillhole	MRC028	27.00	30.00	3.00	3.74	4.00	93.00
Deborah	Drillhole	MRC029	24.00	30.00	6.00	7.49	3.67	88.50
Deborah	Drillhole	MRC030	12.00	21.00	9.00	9.65	3.05	67.67
Deborah	Drillhole	MRC032	27.00	30.00	3.00	3.66	3.97	74.00
Deborah	Drillhole	MRC033	21.00	24.00	3.00	3.87	1.75	18.00
Deborah	Drillhole	MRC052	59.00	64.00	5.00	5.45	2.92	73.80

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Prospect

Type

Hole ID

From

Length

HW (m)

Au g/t

Ag g/t

Deborah	Drillhole	MRC053	46.00	55.00	9.00	4.42	3.60	79.11
Prospect	Type	Hole ID	From	To	Length	HW (m)	Au g/t	Ag g/t
Deborah	Drillhole	MRC054	59.00	62.00	3.00	3.78	1.42	30.33
Deborah	Drillhole	MRC055	31.00	34.00	3.00	3.79	2.39	17.00

Table 94 List of Geological Intercepts

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15.	MINERAL RESERVES ESTIMATE

The resources at Cerro Moro have not been proven to a reserve level at the stage of writing this study.

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16.	MINING METHODS

16.1	Preliminary Mining Study

16.1.1

Summary

This mining study is based on Cube Consulting’s April, 2011, 43-101 compliant mineral resource statement for Cerro Moro which includes both Indicated and Inferred Category Mineral Resources. This resource estimate for Cerro Moro has been discussed in Chapter 14 of the present report.

The study is categorised as a Preliminary Economic Assessment (“PEA”) instead of a Prefeasibility Study (“PFS”) to facilitate the incorporation of additional mineralized zones. However, the engineering inputs are to the confidence level of a PFS but a portion of the resource included in the study is still Inferred category. Although the lower confidence resource would not be mined until the latter years of the project they do contribute to the mine life.

From a mining standpoint, the principal differences between this current PEA and Extorre’s maiden PEA issued in December, 2010, lie in the following parameters:

1. Mining rate has been increased from 750 t ore/day (270,000 t ore/year) to 1,000 t ore/day (360,000 t ore/year),

2. Metal price assumptions have increased to $US 1,320/oz for gold and $US 26/oz for silver;

3. Mining of the Escondida Far West ore shoot is initially to be undertaken using open pit mining methods, rather than all underground,

4. Use of the April 2011 interim mineral resource as the base for the current study.

Mining of the narrow, near-vertical gold-silver veins at Cerro Moro is expected to be undertaken using a combination of open pit and underground methods. The mine development schedule commences with 8 months of “pre-stripping” (Year 0) at both the Escondida Far West and Escondida Central sectors, in which 6.23 Mt of waste are removed together with 108,900 tons of ore grading 11.55 g Au/t and 573 g Ag/t Au for stockpiling. During the first full year of commercial operation (Year 1), production will be derived from open pit mining at both the Escondida Far West and Escondida Central sectors (264,000 t of ore grading 19.51 g/t Au and 811 g/t Ag). Decline access will concurrently be constructed in to the Escondida Far West, Escondida Central, and Gabriela sectors, with first production (total of 53,901 t of ore grading 5.03 g Au/t and 425 g Ag/t) being achieved in the Fourth Quarter.

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The Cerro Moro open pits will be conventional with hydraulic top hammer rock drill rigs and blasting, backhoe excavators, front-end loaders, and 50t mining trucks, and designed to reach a maximum vertical depth of 70m.

Geomechanical work completed to date has indicated that the host rocks to mineralization at Cerro Moro are highly competent, allowing pit slopes angles to be fixed at 65 - 68 degrees and crown pillars of 10m height to be left between the base of the completed open pits and the subsequent underground mining operations.

A general layout for the Project is shown in both Figure 1 and Figure 143.

The total estimated mine resources amount to 1.7 million tonnes of ore at grades of 6.6g/t gold and 341.9 g/t silver contained in the open pits, with a strip ratio of 26.9 to 1; and 1 million tonnes of underground ore at grades of 4.8 g/t gold and 352.2 g/t silver. The summary by sector is shown in Table 95.

Figure 143 Conceptual Mine Development for Escondida

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Open Pit	ORE				Waste ktonnes	Strip Ratio
Open Pit	ktonnes	Au g/t	Ag g/t	AuEq g/t (*)	Waste ktonnes	Strip Ratio
Escondida	629	14.5	595.4	26.4	23,600	37.5
Gabriela	538	2.0	297.7	8.0	17,838	33.2
Esperanza	307	1.3	89.4	3.1	1,714	5.6
Deborah	212	3.0	67.0	4.3	2,204	10.4
TOTAL	1,686	6.6	341.9	13.5	45,663	27.1
Underground	ORE
Underground	ktonnes	Au g/t	Ag g/t	AuEq g/t (*)
Escondida	467	8.4	395.7	16.3
Gabriela	575	1.9	316.8	8.2
Esperanza
Deborah
TOTAL	1,042	4.8	352.2	11.8
TOTAL	ORE
TOTAL	ktonnes	Au g/t	Ag g/t	AuEq g/t (*)
Escondida	1,096	11.9	510.4	22.1
Gabriela	1,113	1.9	307.6	8.1
Esperanza	307	1.3	89.4	3.1
Deborah	212	3.0	67.0	4.3
TOTAL	2,727	5.9	345.8	12.9

(*) The gold equivalent value is calculated by dividing the silver (ppm / ounces) by 50 (approximate ratio of gold/silver US$ value) and adding it to the gold (ppm / ounces).

Table 95 Mine Resources Summary by Sector

16.1.2

Open Pit Mining

The open pit limits were obtained using the optimization package Whittle Four-X, applying the alternative underground option to determine the economical limit between open pit and underground mining. The optimization process considers that if a particular block can be mined by either above-ground methods or underground methods, then the value that is given to it during open pit optimization is the difference between its value when mined above-ground and its value when mined underground.

For example, suppose a block that is worth US$1,000 if mined by above-ground methods and US$800 if mined by underground methods.

●

If no underground mining is to take place, then the correct value to use during open pit optimization is US$1,000.

●

If it is decided to mine underground below the open pit, then the correct value to use for open pit optimization is US$200.

This is because the advantage to the company obtained by mining the block by above-ground methods is only US$200, since the company will still make US$800 if the block is omitted from the open pit.

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However, if the block is included in the open pit, then US$1,000 is made from it. Therefore, there is a difference between the profit gained from mining the block and the advantage gained from choosing to mine it by above-ground methods.

It is this advantage which is relevant to open pit optimization. Open pits which are designed to maximize advantage to the company are usually smaller than pits which are designed to maximize the value of the pit itself.

Pit optimization parameters shown in Table 96 were used for the Whittle Four-X and results shown in Table 97 were obtained.

Item	Unit	Escondida	Esperanza	Gabriela	Deborah
OP mine cost	$/t	2.6	2.6	2.6	2.6
UG mine cost	$/t	50.0	50.0	50.0	50.0
Process Cost	$/t	50.0	50.0	50.0	50.0
G&A	$/t	22.0	22.0	22.0	22.0
Selling Au	$/oz	5.0	5.0	5.0	5.0
Selling Ag	$/oz	1.0	1.0	1.0	1.0
Recovery Au	%	95.0	95.0	92.5	95.0
Recovery Ag	%	90.0	90.0	87.5	90.0
Au Price	$/oz	1320	1320	1320	1320
Ag Price	$/oz	26	26	26	26
Royalty % of revenue	%	6	6	6	6
Royalty % of net profit	%	3	3	3	3
Royalty %NSR	%	2	2	2	2
IR Slope angle	°	65/68	68	68	68
Cut-off	g/t Au eq*	2.02	2.02	2.07	2.02
(*) Au eq: Gold equivalent value is calculated by dividing the silver (ppm / ounces) by 50 (approximate ratio of gold/silver US$ value) and adding it to the gold (ppm / ounces).

Table 96 Pit Optimization Parameters

The final selection of the Whittle pit shells to be used as guides for mine design was a combination of the classical open pit optimisation with the open pit/underground approach, in order to have a good match between the feed to the plant and the time require to develop the underground mines.

Whittle Four-X Results	Ore			Waste ktonnes	Ore+Waste ktonnes	Strip Ratio	Contained Ounces
Whittle Four-X Results	ktonnes	Au g/t	Ag g/t	Waste ktonnes	Ore+Waste ktonnes	Strip Ratio	Au koz	Ag koz
Gabriela	533	2.17	315.92	13,729	14,261	25.8	37	5,411
Esperanza	145	2.30	157.26	1,353	1,497	9.3	11	732
Deborah	209	2.71	61.03	1,538	1,747	7.4	18	410
Escondida (Central + Loma+FW)	599	17.48	739.23	35,081	35,680	114.0	336	14,228
Total	1,485	8.43	435.18	51,701	53,186	34.8	402	20,781

Table 97 Pit Optimization Results

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Using the above obtained pit limits, operative pit designs were developed, adding 12m ramps and slots to recover as much ore as possible at the pit bottoms. These effects cause a significant increase on the strip ratio on the operative designs when compared with the optimized pit shells. Table 98 summarized the open pit tonnages and grades by sector and Figure 144 through Figure 147 show those designs, including waste storage areas

Open Pit	ORE				Waste ktonnes	Strip Ratio
Open Pit	ktonnes	Au g/t	Ag g/t	AuEq g/t (*)	Waste ktonnes	Strip Ratio
Escondida	629	14.5	595.4	26.4	23,600	37.5
Gabriela	538	2.0	297.7	8.0	17,838	33.2
Esperanza	307	1.3	89.4	3.1	1,714	5.6
Deborah	212	3.0	67.0	4.3	2,204	10.4
TOTAL	1,686	6.6	341.9	13.5	45,357	26.9

(*) The gold equivalent value is calculated by dividing the silver (ppm / ounces) by 50 (approximate ratio of gold/silver US$ value) and adding it to the gold (ppm / ounces).

Table 98 Final Pits Summary

Figure 144 Escondida Open Pits

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Figure 145 Gabriela Open Pits

Figure 146 Esperanza Open Pits

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Figure 147 Deborah Open Pits

16.1.3

Underground Mining

Mining methods for narrow veins can be open stopes and cut and fill systems. A conservative solution has been adopted for Cerro Moro and it was agreed between NCL and Extorre to use a Bench and Fill method. The produced designs were reviewed and validated by AKL, the geotechnical consultants, concluding that a conservative approach was adopted and further optimizations can be made on the following stages of the project.

The main reasons for the adopted mining method are:

	●	Mineralization distribution (sub-vertical veins, width of 1 to 10 m, average of 1.5m)
	●	Poor geotechnical knowledge of rock quality relative to the areas of interest
	●	Poor geotechnical knowledge of rock quality relative to the areas of interest

The method consists of:

	●	Ramps that will give access to the veins, these were designed to be used as haulage routes.
	●	Along the vertical development of these ramps, a ventilation raise will provide fresh air to operations
	●	Every 24m (vertically measured), an access has been designed to allow access to 2 drifting levels (12m each)

	●	Productions starts with drifts, as accesses reaches the vein, and continue in the 2 opposite directions
	●	As soon as the last drift is completed (at bottom of the stope), benching starts from the end of drift towards the center
	●	Mucking will be done with remote control 3.5 yd3 LHD in the base drift

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	●	Once the total ore is mined out, the stope will be filled from the upper drift with waste material derived from development
	●	A new cycle starts when the stope is completely filled

Spilt blasting will be utilized, as required, in the event of narrow veins in order to minimize unnecessary dilution,. This is defined by drilling and blasting just in ore (vein), once the ore is loaded the remaining area is drilled and blasted to complete the drift width.

A cemented sill pillar has been designed in the center of the veins, approximately at 0 m elevation.

Figure 148 Mining Method – Bench & Fill

Figure 149 Bench & Fill – Split Blasting

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Figure 150 Bench & Fill - Sequence

16.1.3.1

Dilution

Dilution was estimated based on NCL experience and agreed with Extorre. The criteria applied assumed an operational over-breaking as a function of the open pit equipment selectivity and underground type of excavation, as detailed in Table 99.

SMU	Over-breaking by side (m)
Open pit Bench	0.3
Drift Split Blast	0.3
Underground Bench	0.5

Table 99 Dilution Criteria

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Vein widths were calculated perpendicular to a best-fitted incline section for each vein. Average values were applied by zone, according to Figure 151. Dilution grade was considered as 0.0 g/t. It is a conservative assumption and future estimates should apply the block model grades.

Figure 151 Underground Dilution Estimation for Escondida & Loma Escondida

Cut-Off Grade

Cut-off grades have been calculated using the following formula:

Where:

●

Cm:

Mining Cost

●

Cp:

Process Cost

●

G&A:

General and Administration Cost

●

Pau:

Gold Price

●

Cs:

Selling Cost

●

Rau:

Gold recovery

Cut-off grades values for open pit mining are:

●

1.95 g/t of AuEq for economic

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Cut-off grades values for underground mining are:

3.3 g/t of AuEq for economic

1.9g/t of AuEq for marginal (without mining operation cost)

Mine Resources

Underground mine resources were evaluated on Selection Mining Units (SMU), defined as follows:

	●	Detail SMU drifts and benches were designed for sectors with measured and indicated resources, applying directly the dilution criteria during the design process, and then divide them into 20m units along strike.
	●	For sectors with inferred resources, on a projection over a vertical section fitted for all veins, veins were divided along drifts (height = 4 m) and benches (height = 8 m) by 20 m length.
	●	A volumetric calculation was done for each SMU, with tonnage, Au (gpt) and Ag (gpt) as resulting variables.
	●	Dilution was applied to each inferred SMU, according to dilution criteria defined previously.
	●	SMU’s above cut-off grade were identified (dilution included).
	●	First phase for mine resources estimation was done selecting resources above cut-off grade.
	●	Second phase consisted of making these SMUs accessible and economically mineable, as a function of the described mining method.

Figure 152 and Figure 153 correspond to examples of the procedure described above. It can be seen that some SMU’s, specifically those named as “benches”, above cut-off grade in the final mineable resource. The reason for this is because of the expenses incurred for access and extraction of them is higher than their income. To evaluate the benefit to mine a bench, the following criteria was applied:

Bench must pay for the drifts right above and below itself. It means that the cost to extract the bench will include the cost to mine the drifts. For this purpose an estimation of 40 US$/m3 was considered.

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Figure 152 Escondida West Mine Resources – Vertical Section

Figure 153 Economic and Non Economic SMU`s

Finally, Table 100 summarized the underground mine resource by sector, which amounts to 1.0 million tonnes at grades of 4.8 g/t gold and 352.2 g/t silver. It must be noted that any underground mine resources were estimated for Esperanza and Deborah.

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Underground	ORE
Underground	ktonnes	Au g/t	Ag g/t	AuEq g/t (*)
Escondida	467	8.4	395.7	16.3
Gabriela	575	1.9	316.8	8.2
Esperanza
Deborah
TOTAL	1,042	4.8	352.2	11.8

(*) The gold equivalent value is calculated by dividing the silver (ppm / ounces) by 50 (approximate ratio of gold/silver US$ value) and adding it to the gold (ppm / ounces).

Table 100 Underground Mineable Resource Summary

Figure 154 and Figure 155 show schematic 3D layouts of the required underground developments.

Figure 154 Escondida Schematic Underground Layout

Figure 155 Gabriela Schematic Underground Layout

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16.1.4

Mine Production Schedule at 1,000tpd

A combined open pit and underground mine production schedule was developed to show the ore tonnes and grades, tonnes of waste material and tonnes of total material by year for the life of the mine, for a plant throughput rate of 1,000 tonnes per day, with two months of ramp-up. The distribution of ore and waste contained in each of the mining phases and stopes was used to develop the schedule, assuring that criteria such as continuous ore exposure, mining accessibility, and consistent material movements were met.

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

Careful grade control must be carried out during mining to minimize misplaced ore due to the inclination of the orebody. Ore control efforts will require taking advantage of all the experience of selective mining to assure the ore classes are properly selected and processed. These efforts should include the following standard procedures:

	●	Implement an intense and systematic program of sampling, mapping, laboratory analyses, and reporting.
	●	Utilize specialized in-pit, bench sampling drills for sampling well ahead of production drilling and blasting.
	●	Use small excavators and benches no higher than 5 metres (as presently planned) to selectively mine ore zones.
	●	Maintain top laboratory staff, equipment, and procedures to provide accurate and timely assay reporting.
	●	Utilize trained geologists and technicians to work with excavator operators in identifying, marking, and selectively mining and dispatching crusher ore, ROM, and waste.
	●	Best practices will be driven by sampling heading in the veins in order to have better definition of the ore zone and potential dilution, for the underground mining.

The general adopted sequence considers mining simultaneously the open pits with the underground mines, starting from Escondida, followed by Gabriela, then Esperanza and finally Deborah, graphically shown in Figure 156.

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	PP	Year1	Year2	Year3	Year4	Year5	Year6	Year7	Year8
Open Pit
Escondida						–	–	–	–
Gabriela	–	–	–	–
Esperanza	–	–	–	–	–	–
Deborah	–	–	–	–	–	–	–	–
Underground
Escondida	–								–
Gabriela	–

Table 101 General Mining Sequence

The detailed mine production schedule by sector is shown in Table 101. As an example of the mine evolution Figure 156 through Figure 159 show Escondida situation at the end of re-production, year 1, year 2 and year 3, respectively.

The plant feed is a result of a combination between what is mined from the underground mine, open pit mine and re-handling from the stockpile; and it is shown in Table 102.

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			pp	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	TOTAL
Waste	Escondida	ton	6,228,800	6,363,900	5,468,800	5,193,500	344,400	-	-	-	-	-	23,599,400
	Gabriela	ton	-	-	-	-	5,462,400	5,977,500	4,222,300	2,053,100	123,300	-	17,838,600
	Esperanza	ton	-	-	-	-	-	-	415,200	1,193,900	105,300	-	1,714,400
	Deborah	ton	-	-	-	-	-	-	-	-	1,996,900	207,500	2,204,400
	Total	ton	6,228,800	6,363,900	5,468,800	5,193,500	5,806,800	5,977,500	4,637,500	3,247,000	2,225,500	207,500	45,356,800
Open Pit Ore	Escondida	ton	108,900	263,900	102,000	135,400	19,200	-	-	-	-	-	629,400
		Au g/t	11.55	19.51	11.57	10.18	7.30	-	-	-	-	-	14.47
		Ag g/t	573.66	811.85	365.66	416.81	222.10	-	-	-	-	-	595.35
	Gabriela	ton	-	-	-	-	158,300	182,200	121,400	59,600	15,900	-	537,400
		Au g/t	-	-	-	-	2.01	2.12	2.07	1.79	0.85	-	2.00
		Ag g/t	-	-	-	-	298.49	318.91	307.55	255.53	128.92	-	297.68
	Esperanza	ton	-	-	-	-	-	-	116,600	141,100	49,100	-	306,800
		Au g/t	-	-	-	-	-	-	1.54	1.17	1.13	-	1.31
		Ag g/t	-	-	-	-	-	-	56.40	111.00	105.63	-	89.39
	Deborah	ton	-	-	-	-	-	-	-	-	154,600	57,400	212,000
		Au g/t	-	-	-	-	-	-	-	-	2.96	2.98	2.96
		Ag g/t	-	-	-	-	-	-	-	-	66.62	67.93	66.97
	TOTAL	ton	108,900	263,900	102,000	135,400	177,500	182,200	238,000	200,700	219,600	57,400	1,685,600
		Au g/t	11.55	19.51	11.57	10.18	2.58	2.12	1.81	1.35	2.40	2.98	6.65
		Ag g/t	573.66	811.85	365.66	416.81	290.23	318.91	184.51	153.92	79.85	67.93	341.90
Total Open Pit Material	Escondida	ton	6,337,700	6,627,800	5,570,800	5,328,900	363,600	-	-	-	-	-	24,228,800
	Gabriela	ton	-	-	-	-	5,620,700	6,159,700	4,343,700	2,112,700	139,200	-	18,376,000
	Esperanza	ton	-	-	-	-	-	-	531,800	1,335,000	154,400	-	2,021,200
	Deborah	ton	-	-	-	-	-	-	-	-	2,151,500	264,900	2,416,400
	Total	ton	6,337,700	6,627,800	5,570,800	5,328,900	5,984,300	6,159,700	4,875,500	3,447,700	2,445,100	264,900	47,042,400
Open Pit Strip Ratio (Ore/Waste)			57.2	24.1	53.6	38.4	32.7	32.8	19.5	16.2	10.1	3.6	26.9

Table 102 Mine Production Schedule

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(Continue)
			pp	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	Total
Undeground Ore	Escondida	ton	-	15,604	108,726	130,878	123,148	65,056	22,792	554	-	-	466,758
		Au g/t	-	12.71	10.33	6.27	8.61	8.67	6.57	9.38	-	-	8.40
		Ag g/t	-	642.04	483.86	271.29	426.51	401.93	339.92	248.36	-	-	395.68
	Gabriela	ton	-	38,297	57,826	61,684	63,960	99,565	98,801	89,712	50,327	14,891	575,063
		Au g/t	-	1.91	2.19	2.52	1.61	2.07	1.94	1.37	1.49	0.87	1.85
		Ag g/t	-	337.07	370.37	426.16	278.46	355.71	326.31	236.45	252.39	148.29	316.84
	TOTAL	ton	-	53,901	166,552	192,562	187,108	164,621	121,593	90,267	50,327	14,891	1,041,821
		Au g/t	-	5.03	7.50	5.07	6.22	4.68	2.80	1.42	1.49	0.87	4.79
		Ag g/t	-	425.36	444.46	320.90	375.90	373.98	328.86	236.52	252.39	148.29	352.16

TOTAL Ore	Escondida	ton	108,900	279,504	210,726	266,278	142,348	65,056	22,792	554	-	-	1,096,158
		Au g/t	11.55	19.13	10.93	8.26	8.43	8.67	6.57	9.38	-	-	11.88
		Ag g/t	573.66	802.37	426.65	345.28	398.94	401.93	339.92	248.36	-	-	510.33
	Gabriela	ton	-	38,297	57,826	61,684	222,260	281,765	220,201	149,312	66,227	14,891	1,112,463
		Au g/t	-	1.91	2.19	2.52	1.89	2.10	2.01	1.54	1.33	0.87	1.93
		Ag g/t	-	337.07	370.37	426.16	292.73	331.91	315.97	244.06	222.75	148.29	307.58
	Esperanza	ton	-	-	-	-	-	-	116,600	141,100	49,100	-	306,800
		Au g/t	-	-	-	-	-	-	1.54	1.17	1.13	-	1.31
		Ag g/t	-	-	-	-	-	-	56.40	111.00	105.63	-	89.39
	Deborah	ton	-	-	-	-	-	-	-	-	154,600	57,400	212,000
		Au g/t	-	-	-	-	-	-	-	-	2.96	2.98	2.96
		Ag g/t	-	-	-	-	-	-	-	-	66.62	67.93	66.97
	TOTAL	ton	108,700	317,901	268,552	327,962	364,608	346,921	359,693	290,967	270,026	72,390	2,727,720
		Au g/t	11.57	17.05	9.05	7.18	4.45	3.34	2.15	1.38	2.23	2.54	5.94

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		Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	TOTAL
Ore	tonnes	307,901	335,752	335,662	335,908	335,821	335,893	335,867	332,427	72,491	2,727,721
Gold	Au g/t	16.70	9.83	7.40	4.97	3.60	2.39	1.89	2.32	2.59	5.94
Silver	Ag g/t	733.78	455.85	373.53	354.76	349.01	253.33	203.00	129.45	85.99	345.81
Contained Gold	oz	165,275	106,064	79,847	53,670	38,843	25,836	20,415	24,759	6,038	520,746
Contained Silver	oz	7,263,709	4,920,701	4,031,013	3,831,220	3,768,122	2,735,767	2,192,041	1,383,488	200,418	30,326,479

Table 103 Plant Feed at 1000 tpd

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			Construction	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	TOTAL
Open Pit Mining	Ore	ktonnes	22	257	196	166	166	159	111	39	1,116
	Gold	Au g/t	2.41	17.90	14.79	8.76	2.64	2.17	1.99	1.23	9.02
	Silver	Ag g/t	152.31	772.90	572.46	377.71	305.64	321.68	276.20	166.70	462.48
	Contained Gold	koz	2	148	93	47	14	11	7	2	323
	Contained Silver	koz	107	6,394	3,607	2,012	1,627	1,640	989	209	16,586
	Waste	ktonnes	3,271	8,023	6,889	6,666	5,143	5,470	4,195	987	40,643
	Total Mined	ktonnes	3,293	8,280	7,085	6,832	5,308	5,629	4,306	1,026	41,759
Underground Mining	Ore	ktonnes	-	2	43	104	127	128	59	7	471
	Gold	Au g/t		19.58	7.69	6.31	4.77	6.07	6.19	12.04	6.08
	Silver	Ag g/t		972.12	528.15	401.39	308.63	421.51	415.84	440.56	398.16
	Contained Gold	koz	-	1	11	21	20	25	12	3	92
	Contained Silver	koz	-	59	738	1,336	1,265	1,737	790	102	6,027
Total Mining	Ore	ktonnes	22	259	239	269	293	287	170	46	1,586
	Gold	Au g/t	2.41	17.91	13.50	7.82	3.57	3.91	3.45	2.92	8.15
	Silver	Ag g/t	152.31	774.34	564.42	386.82	306.94	366.30	324.60	209.49	443.39
	Contained Gold	oz	2	149	104	68	34	36	19	4	415
	Contained Silver	oz	107	6,452	4,345	3,348	2,892	3,377	1,779	311	22,613
Plant Feed	Ore	ktonnes	-	234	256	256	255	255	255	74	1,586
	Gold	Au g/t	-	18.01	13.20	7.99	3.80	4.21	3.82	3.55	8.15
	Silver	Ag g/t	-	778.42	562.22	393.35	317.31	372.13	320.98	249.17	443.39
	Contained Gold	oz	-	136	108	66	31	35	31	8	415
	Contained Silver	oz	-	5,866	4,621	3,233	2,606	3,055	2,636	596	22,613

Table 104 Mine Plan and Plant Feed at 750 tpd

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Figure 156 Preproduction Year – Escondida Section

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Figure 157 Year 1 – Escondida Section

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Figure 158 Year 2 – Escondida Section

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Figure 159 Year 3 – Escondida Section

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16.1.5

Mine Production Schedule at 750tpd

A full mine exercise, at pre-feasibility level, considering only the measured and indicated resources, and the inferred resources treated as waste, was developed at a throughout rate of 750 tonnes per day.

16.1.6

Mining Equipment Requirement for 1,000 tpd

The equipment fleet was estimated based on the mine production schedules, performance and operational indices of the main equipment, and factors related to the main equipment for the auxiliary units. A spreadsheet model was created to estimate the number of operating hours and units required. Hauling distances were obtained from the mine plan. This model was also used to calculate equipment procurement schedules, workforce productivity, capital expenses and operating costs. Equipment and workforce productivity were estimated according to experience and industry standards for the size of the equipment and mine production rate.

Open Pit Mining Equipment

The following criteria were adopted for mine equipment selection for all of the open pits:

	●	Medium capacity units, consistent with the size of the operation
	●	Proven technology, similar to many other operations of similar type
	●	Flexibility in operations, where required
	●	High productivity in waste removal, as well as high selectivity for ore extraction.

The equipment was selected for a standard mining operation with conventional drill, blast, load and haul, plus the regular activities of face and road maintenance. The equipment selection includes:

	●	diesel drills of 35/8- 5 " diameter (Pantera DP1500) for waste and ore
	●	8.3 yd3 front-end loaders (Cat 988H) for waste and ore
	●	6.2 yd3 hydraulic excavator (Cat 385) for ore and waste
	●	55 tonne capacity trucks (Cat 773).

This main fleet is complemented with ancillary units including track dozers (Cat D8), wheel dozers (834H), motor graders (Cat 14M) and 15 m3 water trucks.

The total open pit requirement is show in Table 105 including also minor support equipment.

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Description	Pre-strip	Peak
Main Equipment
Loading
FEL (8.3 yd3)	1	1
Excavator (6.2 yd3)	1	1
Hauling (50 tonnes truck)	5	5
Drilling (diesel 35/8 - 5 ")	2	2
Ancillary
Bulldozer (D8)	1	2
Wheeldozer (834H)	2	2
Motorgrader (14M)	1	1
Water truck (15m3)	2	2
Support
Backhoe	1	1
Fuel Truck	1	1
Lube Truck	1	1
Support Truck	1	1
Lowboy Truck	1	1
Lightning Plant	8	12

Table 105 Open Pit Mining Equipment Requirement

Underground Mining Equipment

Underground mining equipment was selected to work simultaneously in Escondida and Gabriela.

Drills

Jumbo equipment will be used for two types of drilling work: (1) horizontal drilling for advance drifts and bolting; and (2) single level stoping and 4 m width stopes.

Two boom jumbos have been selected due to their higher level of productivity compared with a single boom jumbo. This is important due to the large number of work areas in the mine (drilling both tunnels and stopes).

Accuracy and straightness is required for drilling in 12 m high stopes. Two types of drilling rigs can be used, down-the-hole (DTH) and hydraulic top hammer. DTH was selected for a drilling pattern of 2.0 x 1,2 m.

Bolting jumbos were selected for regular support during the developments.

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Loading Units

Loading operations will be done with 6 yd3 LHDs operating in developments and smaller units of 3.5 yd3 will be used for production (drifts and stopes).

Trucks

Trucks of 30-tonne capacity complete with push plates have been selected to transport ore from the haulage levels to the surface, and to haul waste material for backfill from surface (or from oading points inside the mine) to the stopes. Trucks of this size can dump in tunnels of normal height, so the tunnels do not have to be enlarged.

Ancillary Fleet

The estimates for equipment type and quantity took into account the maximum number of active stopes at a given time and the overall requirements for a modern standard operation.

The following support equipment was selected for the project:

●

Utility trucks:

For explosives distribution: These can be conventional diesel trucks, 1,500 kg loading capacity, equipped with all the requirements established in regulations for explosive transportation.

For materials transportation inside the mine (all-purpose): Conventional flat-bed trucks, 1,500 kg loading capacity.

For general maintenance services: This is to reduce the number of trips by equipment to the maintenance shop, particularly the Jumbos.

●

Main and Ancillary Fans:

Main fans were estimated as per the ventilation requirement. Ancillary fans are estimated as a function of active stopes and galleries in development at a given time, according to production and development plans.

Development Fleet

To develop 5.0 x 4.5 m and 4.0 x 4.0 m tunnels, the same equipment described above will be used (i.e., two boom jumbos, 6 yd3 LHDs, and 30 tonne trucks).

The total underground requirement is Table 106 shown including also minor support equipment.

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	Year 1	Peak
Main Equipment
LHD 6 yd3	2	3
LHD 3.5yd3	3	4
Trucks 30t	4	7
Jumbo horizontal	6	8
Jumbo Radial	1	1
Jumbo support	3	6
Explosives charger	2	4
Others
Compressores 7cfm	1	1
Main Fans	5	5
Secondary Fans	7	7
Service truck	1	1
Fuel truck	1	1
Lube truck	1	1
Explosives truck	1	1
Lighthning plant	6	6
Pickups	10	10
Plataforma de levante	3	3
Mixer 6 m3	1	1
Minibus	3	3
Gunning machine	1	1
Water truck	1	1
Scaler	2	2
Auxiliary truck	1	1

Table 106 Underground Mining Equipment Requirement

16.1.7

Mine Labour

The mining department will be led by the Mine Manager who needs to be a fully qualified graduate engineer with 20 years of experience for open pit and underground operations including mine planning and maintenance. The required personnel are shown in Table 107 through Table 110.

The Mine Manager needs to be hired 3 – 6 months before the start of the mining operations so that he can be involved with the hiring process for the operators and maintenance people plus be involved with the preparation of the training program.

In addition to the Mine Manager, there are four other key people for the other areas of the mining operation for maintenance, mine engineering and planning, geology. The senior people for these positions will require 15 – 20 years of experience and need to be hired two months before the start of the mine operations.

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Period	Construction	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9
Total Open Pit Operators	57	61	49	51	53	55	51	47	42	47
Loading	9	9	9	9	9	9	9	9	9	9
FEL	4	4	4	4	4	4	4	4	4	4
Hydraulic Shovel	5	5	5	5	5	5	5	5	5	5
Hauling	19	18	12	13	15	17	13	10	7	13
Haul trucks	19	18	12	13	15	17	13	10	7	13
Drilling	9	9	5	5	5	5	5	5	5	5
Diesel Drill	9	9	5	5	5	5	5	5	5	5
Ancillary	13	18	16	17	17	17	17	16	15	14
Bulldozer D8	4	9	9	9	9	9	9	9	9	9
Wheeldozer 834H	4	4	4	4	4	4	4	4	4	1
Motorgrader 14M	2	2	1	1	1	2	1	1	1	1
Water Truck 15000 l	3	3	2	2	3	3	2	1	1	1
Support	7	7	7	7	7	7	7	7	6	6
Backhoe	2	2	2	2	2	2	2	2	2	2
Fuel Truck	1	1	1	1	1	1	1	1	1	1
Lube Truck	1	1	1	1	1	1	1	1	1	1
Support Truck	1	1	1	1	1	1	1	1	1	1
Lowboy Truck	1	1	1	1	1	1	1	1

Table 107 Open Pit Operators

Period	Construction	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9
Open Pit Maintenance	42	53	53	53	53	52	46	38	31	31
Site	7	13	13	13	13	9	6	4	4	4
Drills	4	8	8	8	8	4	4	3	3	3
Shovels	3	5	5	5	5	5	2	1	1	1
Workshop	35	40	40	40	40	43	40	34	27	27
FEL	5	5	5	5	5	5	5	1	1	3
Truck	13	15	15	15	15	18	18	18	14	14
Ancillary	18	20	20	20	20	20	18	15	12	10

Table 108 Open Pit Maintenance

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Period	Construction	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9
Underground Operators & Maintenance		73	102	56	49	44	44	36	12	12
LHD Operators		9	9	11	11	9	9	6	3	3
Truck drivers		11	17	11	11	9	9	6	3	3
Frontal Jumbo Operators		17	19	9	6	6	6	6
Support Jumbo Operators		9	17	6	3	3	3	3
Explosives Charger Operators		6	11	3	3	3	3	3
Radial Jumbo Operators		3	3	3	3	3	3	3	3	3
Mechanic/Electrician		14	21	10	9	8	8	7	2	2
Mine Services		4	5	3	3	3	3	2	1	1
Supervision & Administration		33	33	33	33	33	33	33	33	33

Table 109 Underground Operators & Maintenance

Indirect Personnel	TOTAL	Common	Open Pit	Underground
Mine Management	12	2	4	6
OP & UG Mine Operations Manager	1	1
Mine Captains	1			1
OP Shift Supervisor	4		4
UG Shift Supervisor	5			5
Foreman Drill/Blast	1	1
Mine Maintenance Overhead	9	6	2	1
Maintenance Manager	1	1
Maintenance Chief	1	1
Maintenance Programming Engineer	3		2	1
Maintenance Assistant	1	1
Maintenance Shift Leaders	3	3
Technical Services	25	15	6	4
Senior Engineer	1	1
Engineers	2		1	1
Mine planning Draftsman	1	1
Statistician	1	1
Surveyor	2		1	1
Surveyor helpers	6		4	2
Samplers	6	6
Geology Manager	1	1
Geologist	1	1
Geology Draftsman	1	1
Geology Helper	2	2
Rock Mechanic Engineer	1	1
TOTAL	46	23	12	11

Table 110 Indirect Personnel

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16.2	Geotechnical Studies

The stability analysis is performed using the generalized limit equilibrium method, GLE, and the software Slide 5.0 (Rocscience (2005)), for which there were analyzed all possible alternatives fail, that is, overall slope, interamp angles, combination more than one interamp angles, sectors related to local geological contacts and/or geological faults, etc. At the same time, the probability of failure, PF, was calculated by the method proposed by Duncan (2000).

AKL recommended slopes angle can be summarized as follows:

●	Escondida West:	South-West Wall 68 °
		North Eastern Wall 65 °
●	Escondida Central:	South-West Wall 68 °
		North Eastern Wall 65 °
●	Loma Escondida:	South-West Wall 65 °
		North Eastern Wall 68 °
●	Esperanza: 68° overall
●	Gabriela: 68° overall
●	Deborah: 68° overall

Its AKL opinion the stability analysis supports a pre-feasibiity study.

	f)	According to the limited information available in some parts of Escondida, where will be an important part of the developed of the slope of each pit, recommendation interramp angle is lower than in other sectors. Furthermore, the recommended maximum height should be within the range of 60m to 70m maximum height for all prospects with 5m bench in the case of single bench, which can turn left 10m to achieve greater containment berm materials, bench face angles for this design will remain between 84 to 85

	g)	With regard to the development of new designs, modification of current changes in certain parametres such as height, etc. Reassessment should be performed geotechnical the updated geotechnical information and the proposed designs.

	h)	In a few cases of prospects considered here, such as the Esperanza pit design, an overload is detected of unconsolidated gravel with a thickness greater than a bench, for such case, it is recommended that a catch bench at the gravel contact -mass of about 10 m, with the objective of containing the instabilities in such poor quality geotechnical unit.

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As a result of this study are some important recommendations for the design of the different slopes of the Cerro Moro Project prospects.

	a)	In order to improve the quality of information concerning the properties of the geotechnical units and notwithstanding that there has been progress in this regard, it is necessary to conduct further geotechnical testing.

	b)	Must be done to build a model of major structures, which must consider and include large geological structures, such as regional or district scale structures also mine.

	c)	Hydro geological studies should be conducted more detailed, taking into account that the period of development of the slope does not provide sufficient time for the groundwater flow reaches a steady state, so that should be considered a transient state of these

	d)	According to information and characterization of the rock mass that exists today, we recommend the drilling of some additional geotechnical drilling covering some areas with little information in the some prospect involved here

As a result of this work can be noted:

	1.	Assessment was conducted geomechanical stability of underground mining method called "Cut and Fill by Ascending Bench (Bench and Fill), emphasizing in particular as regards the stability of cave and crown pillar.

	2.	The results of the stability analysis of the caves, we suggest that, because their dimensions considered 25m high and 4m wide, and also that these will be filled as they come to be mined, would not have a maximum length restrictions, Therefore, for this stage of engineering is recommended maximum length of 60m without fortification associated.

	3.	For the stage of engineering pre feasibility of this work, we consider a FS> 3.8, this mainly to calculate the thickness of the crown pillar, which can be designed for contact with the surface of the original topography or to the interaction the walls or bottom of a open pit mining. If the length of the pillar, Lp <60m, these pillar will be of 7m thick, if the length of the pillar, 60m < Lp <80m, we recommend a thickness of 10m. It should complement these studies with other types of analysis in order to validate these parametres, in the following engineering stages of Cerro Moro project.

For the next stage of engineering Cerro Moro, in underground mining is recommended:

	a)	While geomechanical analyses made are valid, they are based on empirical analysis, so they must be improved to have better information geomechanics.

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	b)	"Anomalies" in the geological condition, such as the presence of faults, shear zones, dykes or inclusions of waste rock, creating a slot or visor inside the cave and poor support installation can all, individually and collectively, bring to an inappropriate result.

	c)	Practical observations suggest that the main area of uncertainty when using the method is regarding to the density of structures in the rock mass. The number of discontinuities and other structures per unit volume of rock is highly variable

	d)	The design and recommendations derived from the use of this method should be considered a first step in the design process. It should make all appropriate adjustments based on conditions observed in the cave of interest.

	e)	Analysis should be performed by the failure factor, which successfully integrated into the design of caves the presence of major faults near the exposed surfaces.

	f)	The effects due to blasting activities are usually ignored. Poor quality blasting can create new fractures and relax rock blocks, which consequently reduces the quality geotechnical rock mass.

	g)	The faults have the most influence on the stability of a cave where the failure is sub parallel to the wall of the cave (20 ° to 30 °), the minimum effect is generated when faults are approximately perpendicular to the walls and or roof of the cave, therefore must be characterized if these exist.

	h)	In general, when studying the stability of a crown pillar, can identify six major failure modes, each of which can be given alone or a combination of more than one of them. In general, the dominant mechanism depends on the geotechnical characteristics of each sector under study. These mechanisms are: Structurally Controlled Fault, Fault rock mass stress-induced failure, Failure type fireplace, beam or plate failure rate and failure rate for blocks voussoir.

	i)	Define acceptability criteria

According to the results of stability analysis for the design of waste dump in Chapter 9 of this report, we can indicate the following:

Waste Dump considered in this assessment will be built with waste material, deposited by the hopper dump trucks extraction from mining operations of the various prospects of Cerro Moro Project, also be mixed in a ratio of 4:1 paste-like materials from the stages of mineral processing project.

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	2)	According to the designs considered, the analysis identified 06 sections to cover all surfaces of the proposed designs for which are generally determined by a design section, this being the most critical section, either by design of the dump itself, topography of the base will be deposited, or quality of the soil foundation.

	3)	The maximum height of the proposed dump designs of the project, ranging from 40m at most, considering the overall heights, the heights interramp, in these designs are variables.

	4)	To perform the stability analysis, we analyzed 06 vertical sections above, considering that these include the natural topography where dumps will be placed and also presented the highest peaks of the slopes of the previously proposed designs dumps.

	5)	Based on field data collected, review of technical literature and the experience of other mining projects with similar characteristics, we estimated the geotechnical properties of foundation soil and basement rock, which is deployed the waste dumps that interest in this assessment.

	6)	Considering the nonlinear behavior of the material will be form the body of the dumps, and although their overall heights near the height limit for geotechnical properties were defined in terms of the magnitude of the confinement stress, these were considered as allows a better representation of real behavior of granular materials free.

	7)	The stability analysis was performed by limit equilibrium, determining the safety factor, FS, using the GLE method, which was preferred over other methods because it considers the best way the effect of the nonlinearity of the envelope of failure for this type of material.

	8)	Also evaluated for each section, the probability of failure, PF (eg see Duncan (2000)), considering the variations coefficients of 10% for the friction angle, and 40% for cohesion.

	9)	The design analyzed the dumps does not consider the presence of water below the surface of these, nor runoff that eventually can penetrate the base and body of these.

	10)	According to the above and as a result of geotechnical evaluations, the designs dumps of the prospects: Escondida, Loma Escondida, Gabriela, Esperanza and Deborah, meet the acceptability criteria defined for this report, considering static and pseudo cases, as outlined in this report.

	11)	The stability of the dump is valid under the considerations of a coarse granular material in a homogeneous mixture of 4:1 with paste that coming of stage of mineral processing and properties similar to those presented in this report

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As a result of this study is presents the following recommendations for the design of the waste dumps slopes in the Cerro Moro prospects.

	a)	While the condition of stability for the proposed design is acceptable to a maximum height of 40m, and overall height, it is advisable to stay decoupling, as the results obtained give us close to the minimum values according to acceptability criteria considered this report.

	b)	Global stability of the proposed design improves considering leaving terraces free of at least 10m wide, it is advisable to reduce the width, but is likely to do so, until you get a better characterization of the dump body materials.

	c)	According to the above, for the pre feasibility in Cerro Moro Project, the design dumps parametres recommended is finally:

	●	Maximum height 40m.
	●	Interramp slope angle 32°
	●	Overall slope angle 25°
	●	Heights first layer 10m
	●	Heights top two layers 15m
	●	Wide terraces 10m

	d)	For the following stages of the project is recommended to conduct a campaign to characterize the materials involved in the dumps design.

	e)	We recommend testing homogeneity of the dump body materials, as there is still the procedure of mixing of granular materials and paste that make up the body of these.

16.3	Hydrology and Hydrogeology

As part of the hydrogeological study of the Cerro Moro Project, the followed prospects were evaluated in detail: Escondida, Esperanza, Gabriela, Silvia Carla, Casius, Lechuzo, Deborah, Moro, and Nini. The piezometric characteristics of a total of 293 drill holes were measured for the study. In addition, a total of 18 new holes were either drilled or re-drilled in order to carry out pumping tests, and a further 17 drill holes located in close proximity to these new holes were selected to carry out constant head tests (slug tests). The main purpose of the slug tests was to provide information on the boreholes’ construction quality, the ideal pumping rate, transmissivity (T), the aquifer's storage factor (S) and permeability (K), and to test for the existence of nearby impervious barriers or borders, zones of recharge / discharge, and areas of fluid flow, etc. Water samples were also taken from tested holes and piezometers.

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After completing a detailed field study and literature review of the Cerro Moro Project, the conclusion may be drawn that no permanent water courses exist in the district. Indeed, the project area contains a total of 38 endorheic superrficial water basins. From the perspective of underground water flows, the main “head” areas lie in the northern part of the project area with the main discharge areas being located in the south. Another “head” area is located along the water divide separating the Cerro Moro project from the Laguna El Mosquito.

Distinct groundwater flow lines are identifiable depending on the sector, NW to SE, NE to SW, and N to S; gradients vary from 8% (North of Deborah) to 9% (North East of Escondida Far West) on the Eastern slope of the Tertiary gravel cover.

Hydrolithology

Three main hydrolithological units have been identified in the Project area.

Fractured rock unit (Chon Aike Formation): Volcaniclastic and acid volcanic rocks (ignimbrites, tuffs, and rhyolites), which only enable water circulation via its secondary porosity (density, fracture continuity and intercommunication, presence of joints, faults, open stratification planes), and degree of cementing and argillic alteration. This unit is considered to be a poor aquifer with water flow only being developed along the main fracture systems; in areas distal to these fracture systems, this rock unit forms an aquifuge.

Tertiary clay unit: Includes all rock types in the Patagonia Formation. In the Cerro Moro district, this geological unit is characterized by the predominance of clay over sand. As a direct result, this unit is characterized by low permeability and is therefore considered to be an aquitard. The Tertiary clay unit is located beneath the conglomerates of the La Avenida Formation, and overlies the Chon Aike formation volcanic rocks and pyroclastics.

Tertiary conglomerate unit:The plateau-like mesetas of sandy gravels that constitute the La Avenida formation are the best water receivers and transmitters to the undersoil, due to their coarse texture and meager gradient. The Tertiary conglomerate unit is typically less than 10 m in thickness, although in specific locations may attain a thickness of up to 20 m. The Tertiary conglomerate unit is the region's most important phreatic aquifer, with the recharge index being higher than for the rest of the exposed rocks (values of 13% were recorded near the city of Tellier; González Arzac, CFI, 1989).

Conceptual hydrogeological model

Given the heterogeneous nature of the local geology, no one single conceptual hydrodynamic model can be applied to the Cerro Moro District. Indeed, it is possible to summarize the likely underground recharge, circulation, and unload alternatives into three models:

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In the first model, a scenario in which gravels of the La Avenida Formation are exposed was considered. In response to the high permeability of the gravels, rainwater easily percolates to the phreatic level (and noting that water movement within the phreatic level generally has a regional component towards the east). A portion of the water is also transferred to the pebbles' basement, formed by the marine sedimentary rocks of the Patagonia formation. As the latter have aquitard characteristics, water is unable to permeate through them.

The second model considered the fact that volcanic and sedimentary rocks of the Bahia Laura Group rocks are found at the surface of an endorheic depression. In this scenario, water filters to the bedrock from recharge areas, in deeply fractured areas, and percolates down the rocks' fractures to the piezometric level, which occurs at varying depths. Subsequently, the groundwater flow mostly follows the principal fault zones in the Project area, which generally coincide with, or are parallel to, mineralized veins. Below the endorheic depression, there would be a vertical water flow exchange, which could also be horizontal depending on the season.

The third scenario is a variation of the second: water percolates via the surface fractures of the Jurassic-age bedrock, or by percolation in the valleys' porous sedimentary rocks. After moving through the transit area in steep slopes (greater gradient), water becomes "stagnant" under the canyons' riverbeds, thus enabling a portion of the water to evaporate. After having analyzed the geological and geotechnical profiles of the drill holes, and having evaluated the hydraulic information derived from the drill holes, it was possible to conclude, that beyond around 80-meters in depth, the rocks are compact and do not permit significant groundwater circulation.

Hydrodynamic modeling results

In order to simulate the hydrogeological model under the different geological scenarios and with time, modeling was undertaken using a Modflow combination in its original version, with Modflow2000 inside the Groundwater Vistas graphical environment.

The technical criteria and bibliographical background recommended working with a 4% underground recharge rate. This index was used for all the hydrodynamic modeling. Rainfall was estimated at 207.5 mm/year, implying an Rd of 0.00002274 m/d.

Storage indexes for the fractured mean ranged from a maximum of 0.011 to values within the 2 E10-5 area. The highest transmissivities were around 58.4 m2/day, while the lowest were around 0.08 m2/day, which reflects how heterogeneous and anisotropic the medium is. In Escondida, cones of depression obtained from modeling do not go beyond the model's limits, except for in hole GW-01, in which the cone may continue for some 300 m in a N-NW orientation. In further detail, depression cones do not cover the mining area, not even during the fifth mining year.

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According to Piper's classification, waters are generally sodium chloride, although there are certain exceptions: slightly sodium-bicarbonate-chloride waters, sodium-bicarbonate-chloride waters, and – to a lesser extent – sodium-sulphate-bicarbonate-chloride waters and sodium chloride bicarbonate waters. The modeling area's hydrogeological reserves (15.81 km2) total 753.615m3, and reach a sustainable pumping level of 615.34 cubic metres/day for the modeled area.

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17.	RECOVERY METHODS

17.1	Introduction

A treatment plant capable of processing 336,000 t/y of the gold-silver ore from the various deposits at the Cerro Moro project is proposed. The design incorporates both flotation and gravity recovery stages to maximise recovery from the variable mineralogy particularly of the silver minerals. Due to the high silver to gold ratios for the deposits – overall average ratio of 60:1 silver : gold – the Merrill Crowe (zinc cementation) process has been included instead of adsorption onto carbon for the recovery of precious metals from the leach solution. The plant design is intended to comply with the requirements of the International Cyanide Management Code. Figure 160 is the process flow diagram for the proposed circuit.

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Figure 160 Overall Flowsheet

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17.2	Crushing Circuit

A three stage closed circuit design has been proposed for the crushing circuit. The circuit is designed for two shift crushing with a design throughput rate of 100 tonnes per hour.

The primary crusher will be a single toggle jaw crusher with a 1,000 mm by 760 mm opening. An apron feeder will transfer ore from a 50 tonne capacity run of mine (ROM) bin to a vibrating grizzly feeder that will scalp the rock greater than 80 mm into the jaw crusher. Jaw crusher product and grizzly undersize will be transferred to the product screen feed conveyor by the primary crusher discharge conveyor. An in-line belt magnet will remove tramp metal from the ore stream at the discharge head of the primary crusher discharge conveyor.

The secondary and tertiary crushers will be standard and shorthead cone crushers respectively that will operate in closed circuit with the product screen. The product screen will be a double deck banana screen 1.8 m wide by 6.1 m long with polyurethane panels on both decks having nominal cut sizes of 25 mm and 11 mm for the top and bottom decks respectively. Oversize from the top deck will be conveyed to the secondary crusher feed bin while the oversize from the bottom deck will be conveyed to the tertiary crusher feed bin. Static tramp metal magnets will be positioned over each of the cone crusher feed conveyors and metal detectors used to trip the conveyors in the event tramp metal that could damage the crushers has not been removed by the magnets.

The cone crushers will be 1,394 mm in size with an installed motor power of 132 kW and will operate with a 25 mm closed side setting for the secondary crusher and 8 mm for the tertiary crusher. Metso Minerals Bruno crushing simulation package was used to determine the equipment sizes and tonnage rates for the mass balance in the crushing plant.

Crushed product with a P100 of 12 mm and a P80 of 8.5 mm will be conveyed for storage in a fine ore bin. The crushing rate will be monitored by a weightometer located on the fine ore bin feed conveyor.

All transfer points will be enclosed and fitted with dust extraction hoods. This will minimise the use of water for dust suppression and avoid exacerbating the material handling issues associated with handling ore with a high proportion of oxide or altered ore. The dust will be discharged from the dust collector into the crushing area sump pump and pumped to the mill discharge hopper.

17.3	FineOre Storage and Reclaim

To cater for single shift crushing a fine ore bin with a live capacity of 1,000 t will provide 24 hours of milling capacity and act as a buffer between the crushing and grinding circuits. The fine ore bin will be provided with an opening to permit the bin to be emptied in the event of a hang up. An emergency reclaim system will enable the grinding circuit to be fed if no crushed ore is available from the bin.

A variable speed belt feeder will be used to remove the ore from a slot in the base of the fine ore bin.

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17.4

Grinding

The proposed grinding circuit design consists of a single stage overflow ball mill that will operate in closed circuit with hydrocyclones, flash flotation and gravity concentration machines. The fine crushed ore feed size and use of a ball mill will enable stable throughput rates to be achieved on each ore type. The treatment rate will be 42 t/h and the grinding circuit will operate 24 hours per day, seven days per week. The inclusion of both flash flotation and gravity recovery devices will maximise the recovery of liberated precious metals minerals into a concentrate which will be reground to increase liberation and exposure of the gold and silver.

The grinding mill will be a 4.27 m diameter (inside shell) by 4.27 m long (EGL) ball mill powered by a 1,300 kW motor. The mill will operate with a ball charge of 35% by volume and a top ball size of 80 mm.

The cyclone cluster will consist of four 250 mm diameter hydrocyclones (three duty – one standby). There will be dual cyclone feed pumps arranged in a duty – standby configuration. The cyclone overflow will have a product particle size P80 of 75 µm.

17.5	Flash Flotation and Gravity Concentration

To maximise the gravity recoverable precious metals a flash flotation machine and a centrifugal gravity separator will treat 80% of the cyclone underflow. The water added in the gravity circuit will displace ball mill feed water to minimise the impact on the density in the grinding mill.

The feed to the gravity circuit will be split from the cyclone underflow which will report to a screen to remove the +3 mm particles. The cyclone underflow stream has been selected as the gravity circuit feed because the precious metals will tend to be concentrated in this stream due to the preferential separation of higher density minerals to the cyclone underflow. The cyclone underflow will also be essentially deslimed so that the effects of viscosity will be minimised in the gravity circuit.

The flash flotation machine will be a SkimAir SK240 fitted with a dual outlet whereby the top outlet will remove a low density stream containing mostly water that will be directed to the mill discharge hopper. This arrangement will reduce the water reporting to the mill feed and help maximise the mill density. The SK240 with a cell volume of 8 m³ will provide a residence time of 5 minutes. The tailing from the flash machine will report directly into a 760 mm diameter centrifugal gravity recovery machine. Tailing from the centrifugal concentrator will report to the feed of the ball mill. Concentrate from the gravity concentrator will combine with the flash flotation concentrate and be directed to the concentrate regrind circuit.

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17.6	Gravity and Flotation Concentrate Handling

The flash flotation and gravity concentrates totalling approximately 1.0 t/h of solids will be reground in an M100 IsaMill to a product size P80 of 30 µm and then thickened in the leach feed thickener along with grinding circuit cyclone overflow.

Theconcentrate regrind circuit will include a dewatering cyclone to provide a feed density for the IsaMill of 50% solids. The use of ceramic media in the mill will minimise the effect of iron on cyanide consumption and leach kinetics. Due to the low tonnage of concentrate – 20 to 25 t/d – the IsaMill will be an M100 pilot plant sized machine powered by a 75 kW motor.

The reground concentrate will be pumped to a 4m diameter high rate thickener where it will be thickened prior to being leached on a batch basis. Leaching will be conducted at 35% solids using a solution of 1% sodium cyanide and 0.2% sodium hydroxide solution with oxygen addition as required. The leach cycles will be 24 hours in duration. Two leach tanks will be provided to permit the cycling between filling and leaching duties. At the completion of each leach cycle the leached residue will be dewatered by a 24 chamber pressure filter with 30 mm chambers and a filtration area of 87 m². The filter cake will be washed to remove soluble gold and silver. The resulting pregnant solution will then be pumped into the Merrill Crowe circuit over a 24 hour period to blend in the high solution tenors with the pregnant solution from the main leaching circuit. The wash filter cake will be repulped and pumped back into the main leach feed circuit to maximise the gold and silver leach extraction.

17.7	Leaching and CCD

The main leaching circuit design includes a leach feed thickening stage to maximise the recovery of non-cyanide contaminated process water prior to leaching and to minimise the volumetric capacity of the leaching train. A 8 m diameter high rate thickener was selected based on a settling flux rate of 1.0 t/m2h determined from test work to achieve an underflow density of 40% solids.

The leaching circuit design has been based on a leach feed density of 38% solids, an initial leach feed pH of 10.5 and the maintenance of a free sodium cyanide level of 1,000 ppm throughout the leach train via staged addition of cyanide. Oxygen will be sparged down the agitator shaft of each leach tank to maintain high dissolved oxygen levels in the pulp and thereby maximise leaching kinetics. The leaching circuit will provide a residence time of 48 hours. Five stages of leaching have been incorporated into the design to minimise short circuiting and maximise turn-over of slurry in the leaching stage. The pH of the leach tail slurry will be maintained at above 10.7 to provide the best conditions for flocculation in the CCD thickeners.

The leach tail will flow to the first of five CCD thickeners which will be used to separate and recover the solution phase carrying the dissolved precious metals from the residue solids. Pregnant solution will be removed from the first CCD thickener. Barren solution from the Merrill Crowe zinc cementation circuit will be added to the fifth (last) CCD thickener as the wash solution.

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A five stage CCD circuit has been selected to achieve a wash efficiency of 99% due to the difficulty filtering the slurry, which has a high inherent fines content due to the presence of illite clay. The CCD thickeners will be 11 m diameter high rate thickeners. Underflow hoppers will be provided to maximise mixing of the wash (‘barren’) solution from the second thickener down the line, with the underflow from any particular CCD thickener. Washing efficiency will therefore be improved.

17.8	Merrill Crowe

The Merrill Crowe circuit has been designed on the basis of information from existing operations and laboratory test work.

The Merrill Crowe circuit will be a vendor supplied package with a volumetric capacity of 120 to 240 m³/h. It will be designed to handle pregnant tenors between 2 to 6 ppm gold and 52 to 207 ppm silver.

The design envisions recycling the barren solution from the Merrill Crowe circuit to the back of the CCD circuit as wash water and also using this solution to provide the flocculant dilution water in the CCD circuit. Excess cyanide containing water will be directed to the cyanide destruction circuit.

17.9

Reagents

The following process additives will be necessary to operate the processing facilities:

●

Methyl Isobutyl Carbinol;

●

Potassium Amyl Xanthate;

●

Copper Sulphate;

●

Hydrated Lime;

●

Sodium Cyanide;

●

Sodium Hyroxide;

●

Oxygen;

●

Sodium Metabisulphite;

●

Zinc Powder;

●

Sulphuric Acid;

●

Flocculant; and

●

Liquefied petroleum gas (LPG).

Reagents will be received to site, mixed and dosed as outlined in Table 111.

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Reagent	Packaging	Mixing	Storage	Dosing
Frother	1000 litre IBC	N/A	1.0 m³	VS metering pump
Potassium Amyl Xanthate	25 kg bags	Bag splitter to 0.25 m³ enclosed, ventilated, agitated tank	0.5 m³	VS Metering pump
Copper Sulphate	1,200 kg bags	Bag splitter to 12 m³ agitated tank	24 m³	VS metering pump and ring main dosing valve
Hydrated Lime	Bulk	70t silo, feeder, vortex mixer	30 m³	Circulating pump, ring main, dosing valves
Sodium Cyanide	1000 kg boxes	Bag splitter to 10 m³ enclosed, ventilated, agitated tank	20 m³	Circulating pump, ring main, dosing valves
Sodium Hydroxide	Bulk	N/A	20 m³	Helical rotor pump
Oxygen	Bulk	N/A	6.59 m³	Sparged down leach tank agitator shafts
Sodium Metabisulphite	1,200 kg bags	Bag splitter to 12 m³ enclosed, ventilated, agitated tank	24 m³	Circulating pump, ring main, dosing valves
Zinc Powder	400 kg bags	N/A		VS feeder
Diatomaceous Earth	400 kg bags	Bag splitter to enclosed, ventilated, agitated tanks		Centrifugal pumps
Sulphuric Acid	1,000 litre IBC	N/A	1.0 m³	Air diaphragm pumps
Flocculant	25 kg bags	200 kg hopper, feeder, wetting head, 6 m³ agitated mixing tank	12 m³	VS metering pumps (one to each thickener)
LPG	Bulk		7.5 m³

Table 111 Details of reagent systems

17.10

Cyanide Destruction

The underflow from the last CCD thickener will be combined with excess barren solution and pumped at a pulp density of 38% solids to the cyanide destruction circuit. Cyanide destruction will be based on the Inco cyanide destruction process. The circuit will consist of two 175 m³ capacity agitated cyanide destruction tanks with a combined nominalresidence time of 4 hours. Sodium metabisulphite, copper sulphate and lime will be added to the first cyanide destruction tank. The WAD cyanide level can be reduced to less than 1.0 ppm using this circuit if required.

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17.11

Tailings Thickening

The tailings from cyanide destruction will be pumped to the tailings thickener for thickening to 55% solids prior to disposal in the tailings storage facility. The tailings thickener will be a 8 m diameter high rate thickener.

17.12

Water Services

Raw water will be produced from desalination plants on the coast and be pumped to site. Raw water will discharge into the raw water pond (nominal capacity 1,500 m³). The raw water pond will be fitted with 1.0 mm thick high density polyethylene (HDPE) liner.

One of two raw water pumps, arranged in a duty – standby configuration, will draw water from the base of the raw water pond to feed non-contaminated raw water to the following:

●

Crusher area and water sprays,

●

Merrill Crowe circuit,

●

Reagent mixing.

One of two gland water pumps, arranged in a duty and standby configuration, will draw water from the base of the raw water pond to feed raw water to gland seals on the horizontal slurry pumps in the process plant.

The gravity fluidising water pump will draw water from the base of the raw water pond to supply fluidising water to the centrifugal gravity concentrator.

The raw water pond will overflow into the process water pond (nominal capacity 1,500 m³). The process water pond will be fitted with 1.0 mm thick HDPE liner. One of two process water pumps, arranged in a duty - standby configuration, will draw from the base of the process water dam to feed process water to the following:

●

General 25 mm NB hose down points throughout the plant,

●

Grinding area dilution water,

●

Flocculant dilution water for leach feed and tailings thickeners.

A diesel operated fire water pump will draw from the base of the process water pond to supply hydrants distributed throughout the process facility.

Tailings return water, which will be contaminated with some fine solids, will be pumped to a settling pond to remove suspended solids. The overflows from the leach feed thickener and the tailings thickener will also be directed to the settling pond. The settling pond will overflow into the process water pond. The settling pond will be dosed with sulphuric acid to ensure the process water is maintained at approximately neutral pH to ensure the flotation performance is not adversely impacted.

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Potable water will be supplied to both the potable water tank and fresh water tank from the raw water supply. The potable water tank will be a 9 m³ capacity HDPE tank. One of two potable water pumps, arranged in a duty – standby configuration, will draw water from the potable water tank to supply the process plant and the safety shower water tank. The fresh water tank will be a 23 m³ capacity HDPE tank. One of two fresh water pumps, arranged in a duty – standby configuration, will draw water from the fresh water tank to supply seal water to the vacuum pump and the regrind mill. The safety shower water tank will be a 23 m³ capacity HDPE tank. One of two safety shower circulating water pumps, arranged in a duty – standby configuration, will draw from the safety shower water tank to feed a set of accumulators and safety showers distributed throughout the processing facility. The safety shower accumulator bank will maintain the pressure in the safety shower piping network in the event of power failure or shutdown of the safety shower circulating pumps.

Given the presence of both flotation and leaching circuits in the process facility the design philosophy has been to ensure that non-cyanide and cyanide containing process water are effectively segregated. Non-cyanide process water from the leach feed and tailings thickeners will be directed to the process water pond via the settling pond. Cyanide containing process water from the Merrill Crowe circuit will be recycled as wash water in the CCD thickeners and also to provide the flocculant dilution water in the CCD circuit. Excess cyanide containing water will be directed to the cyanide destruction circuit.

17.13

Air Services

A set of three wet screw air compressors will generate plant air which will be stored in a plant air receiver with 5 m³ capacity. Plant air will be distributed around the plant to hose points in 25 mm NB galvanised steel pipeline. 1.0 m³ capacity air receivers will be situated at the laboratory and plant workshop facilities.

A pipe from the plant air receiver will be filtered and dried in a refrigerated drier before being directed to a 1.0 m³ capacity instrument air receiver. Instrument air will be reticulated to instruments throughout the plant from this air receiver.

17.14

Electrical

Power will be supplied to the site by the local supply authority at 11,000 Volts.

An 13,000 V switchboard will be installed in the wet plant switchroom. The switchboard will comprise an incoming circuit breaker, two transformer feeders and a starter for the mill drive motor.

The transformer feeders will supply a 1,500 kVA 13,000 V/415 volt step-down transformer for the Wet Plant 415 V Motor Control Centre (MCC) and a 1,000 kVA 13,000/415 Volt transformer for the Crushing and Screening 415 V MCC.

The transformer will be located in a bunded and fenced compound adjacent to their switchrooms.

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The wet plant 415 V MCC will also supply power to the Gold & Reagent Area MCC, site offices, and stores and workshop areas.

The switchrooms will be prefabricated, air-conditioned buildings constructed of Colorbond polystyrene sandwich panels on a substantial steel base and fitted with VESDA smoke detection and hand held fire extinguishers. The switchrooms will be elevated above the ground on either concrete or steel plinths to allow for the installation of cables entries from below.

The switchrooms will house the medium voltage switchboards, 415 V MCCs, variable speed drives, lighting and small power distribution boards and programmable logic controller (PLC) cubicles.

The 415 V MCC will be of form 4 design either single sided or back to back construction and arranged for connection from below. The PLC equipment associated with the motor control modules will be built into one or more tiers of the MCC and the PLC inputs and outputs (I/O) will be hard wired between drive modules and the PLC racks.

Communications between the MCC and control system human machine interface (HMI) will be via Ethernet and will be by fibre or copper as appropriate.

Generally low voltage (LV) drives above 200 kW will be equipped with electronic soft-starters with integral bypass contactors to reduce heat dissipation and these will be housed in the MCC tiers. LV variable speed drives will be VVVF six pulse type and will be either wall or floor mounted depending upon their size and weight.

All drives will have local control stations with start and stop buttons adjacent to the drive to provide local control for maintenance. All major drives will also be remotely operable from the central control room via the operator interface terminal. The operating status of all drives will be displayed on the operator interface mimic pages. Any drive fault will be reported by the control system and an alarm will be initiated and logged.

Control voltage for all drives and local controls will be 24 V DC.

Cable ladders to be installed throughout the plant will be NEMA 20B type, hot dipped galvanised. Where necessary cable ladder bends, risers, tees and reducers of the same specification will be installed. Peaked cable ladder covers will be installed where cables in the ladder are subjected to direct sunlight or the potential for mechanical damage.

All LV power and control cables will be either CPVCPVC or CXLPEPVC armoured cables. Screened cable will be used for all variable speed drive applications. All HV cables will be CXLPEPVC with copper screened conductors and wire armouring.

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High pressure sodium and metal halide light fittings will be used for general plant lighting. Battery back- up lighting will be installed in all switchrooms and access ways to ensure safe evacuation in the event of a blackout. All LV power circuits and sub circuits will be protected by instantaneous residual current devices (RCD).

The process plant control system will be a PLC based system. The HMI will utilise standard personal computers running Citect software. The main control room will be located in the wet plant area with a subsidiary control room in the crushing and screening area.

Two visual display units (VDU) will be installed in the main control room and one in the crushing and screening control room these will provide the operator interface. These VDU’s will present the operator with graphical process information in the form of trends, mimic pages, alarm summaries, logs and reports. This interface will also enable the operator to start and stop equipment, control variable speed drives and alter process set-points.

The adjustment of controller parameters will be made from the controller face plate and it will be possible to password protect this adjustment to prevent unauthorized adjustments. Display screens will be configured for the trending of individual or related parameters and a number of alarm pages will be developed to allow the setting of alarm points attached to various parameters. All analogue input signals including outputs from flow, pressure, temperature and weighing instruments will be displayed appropriately on mimic pages. A short term trend plot for each input and output from the system can be provided where required on the mimic pages.

The analogue and digital input and output (I/O) associated with the plant instrumentation will be cabled to one or more process control cubicles (PCC) within the plant areas. The PCC’s will be located within the area switchroom and will house the PLC racks, instrumentation power supplies and communication hardware. Communications between the PCC’s and control system HMI will be via Ethernet and will be by fibre optic or copper cable as appropriate.

17.15

Projected Energy, Water, Process Material Requirements

The projected energy, water and process material requirements are itemised in Table 112.

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Item	Usage
Interfroth 50 (kg/t)	0.140
Potassium Amyl Xanthate (kg/t)	0.025
Copper Sulphate (kg/t)	1.252
Hydrated Lime (kg/t)	4.421
Sodium Cyanide (kg/t)	2.300
Sodium Hydroxide (kg/t)	0.138
Oxygen (m3/t)	0.100
Sodium Metabisulphite (kg/t)	3.589
Diatomaceous Earth (kg/t)	0.280 – 1.600
Zinc (kg/t)	0.816 – 3.166
Sulphuric Acid (kg/t)	0.456
LPG (L/t)	0.110
Flocculant (kg/t)	0.140
Grinding Balls (80 mm) (kg/t)	2.100
Ceramic Media (kg/t)	0.010
Power (kwh/t)	59.3
Water (m3/t)	0.8

Note: Zinc and diatomaceous earth consumption dependent on head grade.

Table 112 Projected Energy, Water and Material Requirements

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18.	PROJECT INFRASTURUCTURE

18.1	Cerro Moro Site

This section outlines the infrastructure associated with the Cerro Moro Gold-Silver Project.

18.1.1

Water Supply

Water demand for the process plant is based on initial operating conditions and the conditions that would prevail should there be no reclaim water available from the tailings storage facility. As such, the plant will require a borefield capable of producing up to 50 m³/h; when operating with the design return water from the tailing storage facility, the demand when operating will reduce to 45 m³/h.

A borefield has been identified and developed to the top of the bore casing that is capable of producing 31 m³/h and additional capital costs have been included for the development of additional bores to meet the plant throughput requirements.

The capital costs allow for the design and supply a borefield pumping system from the top of the bore holes (including pumps and rising mains) to the raw water tank at the process plant. Each bore pump will be driven by a diesel generator feeding switchgear in a control panel controllable via telemetry.

The capital costs include provision for the establishment of pipeline access track. The track will provide maintenance access for the borefield equipment.

18.1.2

Process Plant Earthworks

The plant and its associated infrastructure is contained in an area approximately 45,000 m2 and has been located by topographical information as provided by Extorre. The location has been selected to avoid, as much as possible the prevailing winds, local water courses and sandy river beds while taking advantage of hills for building the ROM pad.

The cost estimates assume that 100 mm of top soil will be removed across the site and stored within 500 m of the plant site for future reuse during mine closure. Based on geotechnical investigative work carried out for the plant site, the estimate assumes that detailed excavation and backfill with selected material will only be required under the heavy foundation loads expected for the ROM bin, crushers, mill and major building foundation areas. The balance of the site will be cut to provide appropriate cross falls for drainage to site open drains and backfilled with material using cut and fill. Allowance has been made in the cost estimate for the mobilisation of a contract crushing plant for the production of suitable gravels for site earthworks and engineered fill.

Plant site drainage will be via these cross falls and then into drains directed to a lined site drainage dam. A two compartment water (raw and process water) dam will also be constructed as part of the site earthworks scope of work.

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18.1.3

Run of mine (ROM) Area

The ROM pad will be constructed from development waste provided by the mining contractor using mine waste.

The plant estimate includes provision for the supply and erection of a concrete wall for the crushing plant.

18.1.4

Roads

Internal process plant roads will be formed and compacted from local gravels. A graded track to provide access to the borefield from the treatment plant has also been included within the water supply capital costs.

Access from and to the site will be via 13 kilometres of access road from the plant site interior roads to the provincial road. Provision has been made in the infrastructure costs to construct this 13 km access road to allow for truck access.

It is assumed, for the purposes of the capital cost estimate that suitable gravel material will be available in borrow pits to be established along the access road.

Construction water will be sourced from existing bores located on the project site. Water will be pumped from these bores into temporary ‘turkey nest’ dams established at suitable locations along the road. Diesel hire pumps will be used to transfer water from the temporary dams to water trucks provided by the site earthworks contractor.

18.1.5

Site Accommodation

Provision has been made for the construction of exploration, permanent and construction accommodation facilities at the site. Costs associated with the supply of food and accommodation at Cerro Moro and payment of any travelling time per employee for the duration of the construction period have been included in the man-hour rate build-up.

The construction facilities have been assumed to be hire facilities for the duration of the construction period, with demobilisation and off hire of the units once the process plant is complete.

Sewage treatment facilities, potable water supply facilities and power supply facilities will be installed to handle the full capacity of the exploration, permanent and construction camps.

18.1.6

Power Supply

Electrical power will be supplied to the Cerro Moro Project via a 132 kV line that will connect to the Argentine national grid (“Sistema Argentino de Interconexion”) at a point located on National Highway 281 between the communities of Antonio de Biedma and Tellier. From Antonio de Biedma, the Cerro Moro High Voltage (HV) power line will broadly follow Provincial Route 66 southwards, crossing the

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Deseado River in the area of Paso Gregores and then continuing southwards to the Estancia El Mosquito and to the Cerro Moro Project. At Cerro Moro, a transformer station with two secondary coils will reduce the 132kV input to 33kV and 13.2kV. Peak electrical consumption for the Cerro Moro Project has been calculated at 7.5MW.

Three possible alternative routes for the HV power line between Antonio de Biedma and Cerro Moro were originally evaluated, with the final option (Figure 161) having been selected on the basis of representing the shortest route (71.1 km) and crossing the least number of third-party surface right holders (“estancias”). Preliminary designs for the 132 kV line contemplate the use of towers 25 m in height, spaced either 250 m (steel towers) or 300 m (concrete towers) apart, with a minimum ground clearance for transmission cables of 7 m in rural sectors and 7.5 m in the vicinity of roads / highways. An Environmental Impact Assessment for the proposed power line will be presented to Santa Cruz Authorities during Q3, 2011.

Figure 161 Proposed power line route for the Cerro Moro Project

18.1.7

Fuel Storage

The major site fuel storage will be located adjacent to the mining contractors workshop. The fuel farm will be constructed using self bunded fuel tanks. This will provide storage for the mining fleet and the process plant mobile equipment. The tanks will be installed directly onto an earthen pad constructed for the fuel storage area.

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18.1.8

Potable Water Supply

A reverse osmosis (RO) plant will be installed on the bank of the plant water dam. The bore water transfer pumps will pump to a RO plant feed tank situated alongside the RO plant. The RO plant feed tank will overflow to the plant water dam. Cleaned water from the RO plant will be directed to an ultraviolet stabiliser and then to the plant potable water tank. A set of pressure sustaining pumps and accumulators will be used to reticulate potable water to service points around the plant and associated infrastructure areas.

Potable water storage and distribution to safety showers in the process plant and to administration, workshop and control rooms in the plant along with power supply to the equipment in the potable water area are included in the process plant capital costs.

18.1.9

Sewerage Treatment

A waste treatment system has been allowed in the capital cost estimates for waste streams from the project. The waste water system will be established adjacent to the administration building.

Provision for the supply and installation of pits and submersible machination pumps have been allowed adjacent to the plant ablutions building and the mine services workshop. These two submersible pumps will transfer effluent to the waste water system adjacent to the administration building.

18.1.10

Vehicle Washdown

A light vehicle wash down facility will be provided capable of cleaning vehicles up to 8 tonne truck size.

A fresh water tank fitted with a high pressure water pump will be installed along with a buried oil/water separator. A purpose built concrete slab will be designed to drain to the oil water separator.

Cleaned effluent from the separator will be returned to the process water streams in the process facility.

18.1.11

Heavy Vehicle Workshop

Heavy vehicle workshop facilities for the mining fleet will be provided by Extorre.

18.1.12

Administration Office Complex

A centralised administration complex will be provided at the process plant. The office will consist of transportable modules complexed together to form a single building. Site installed colorbond verandas will be installed around the perimeter of the building.

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18.1.13

Stores and Plant Maintenance Workshop

A combined stores and workshop building will be constructed. The building will be a steel frame building with a central partition segregating it into a stores building and a maintenance workshop building.

The workshop will also have a personnel access (PA) door connecting to a transportable workshop/stores office. The workshop will be provided with hi-bay lighting. Provision has been made in the capital cost estimates for workshop fit out, tools or equipment.

A roller door at the front and rear of the maintenance building will be installed to provide drive through access into a fenced stores yard. A double access gate at the front of this yard will enable vehicle access to the stores yard.

Provision for a three sided reagents stores building has been made to be located adjacent to the reagent mixing and storage facility.

Two 12 m x 3 m transportable building will be installed near the stores/workshop building. One of these transportable buildings will be outfitted as a male/female ablutions building, the other will be established as a crib room. Provision has been made to outfit this building with a sink, pie warmer, microwave oven, hot water facilities and benches.

Refuse from the ablutions and crib rooms will be plumbed to a below ground concrete tank fitted with a submersible grinder pump and automatic level control, this pump will transfer effluent to a central sewerage treatment facility situated adjacent to the main administration office.

18.1.14

Communications

Provision of a PABX system (mobile satellite connection) and a local UHF radio system for the mining process plant and administration facilities will be established.

18.2	Puerto Deseado Facilities

18.2.1

Office Facilities

The office facilities in Puerto Deseado will be utilised for administration of the project requirements for liaison with local authorities, expediting cargo through the port and procurement requirements.

The office facilities in Puerto Deseado will be the main contact point for labour employment and subcontractor prequalification.

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18.2.2

Laboratory

A contract analytical laboratory will be established in Puerto Deseado that will handle all grade control assays for mining (approximately 100 samples per day) and all process plant assays required for operating of the process plant.

The laboratory will consist of all solution and sample preparation equipment, AAS instrumentation, dust extraction systems and fire furnace for assays necessary to handle the high quantity of samples require to undertake grade control for the mine.

18.3	Mining Facilities

18.3.1

Magazine

The explosives magazine has been designed taking into account an estimated replenishment time of 45 days. The minimum stock was calculated as being 40,000 kg of ammonium nitrate, 57 kg of detonating cord, and 930 kg of detonators,which will be stored in a magazine area of 3,000 square metres with 2 containers for high explosives and 2 equally-sized containers for the detonators warehouse, all protected by a border fence. A temporary underground magazine is considered, capable to store explosives for 15 days.

18.3.2

Workshop

The maintenance workshop for both heavy mining equipment and light vehicles will cover an area of approximately 1,945 square metres, and will be located within a precinct covering a total area of 5,093 square metres. Both the vehicle washing area and mud separation unit will be located separately.

18.3.3

Underground Ventilation

Fresh air will be introduced into the mine through a central raise equipped with one 200 kW fan, and then into the stopes, through the corresponding accesses. Mine air will exhaust from the stopes through the upper levels connected to the exhaust raises where 56 kW fans will be installed; exhaust air will also exit the mine through the main haulage decline and secondary vein raises.

18.3.4

Refuge and Escape Shaft

Mine refuges consist of a transportable, hermetically sealed, 20-person cabin. In the case of fire, personnel may take shelter within the cabin. An air scrubbing system constantly removes he excess CO2 to supply breathable air, and there is a CO detection system. The refuges are stocked with water, food and basic facilities for 48 hours. Cabin dimensions are 6.4 m length x 2.2 m wide x 2.2 m height. The location for refuges will be located at the side of the decline. Escape shaftways consist of a set of metal stairs installed in the main raise that connect to every level of the mine. The stairs will have a rest landing every 26 m and will exit to surface alongside the principal ventilation fan.

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18.4	Tailings Storage Facility

18.4.1

Introduction

Extorre Gold Mines Limited (hereinafter, Extorre) hired from Vector Argentina S.A. (hereinafter, Vector) the engineering services for the development of the tailings storage conceptual level design for Cerro Moro mining project, in the province of Santa Cruz, Argentina.

The conceptual engineering developments have given preliminarily definition to the embankments characteristics, storage capacity, required area, earthworks quantities, waterproofing system, perimeter roads and monitoring system.

18.4.2

Scope and objectives

The objective of this report is to describe the works necessary for the execution of the Tailings Storage Facility. The design criteria, the developed methodology and the information resulting from the Tailings Storage Facility conceptual engineering design are presented.

The scope of this report is the development of technical documents containing enough detail to provide basic information for the development of a cost analysis and the assessment by Extorre of Ausenco Vector’s proposal.

It should be noted that, when sufficient data was not available, the experts involved resorted to their previous experience.

Taking into account Extorre’s requirements, a three-stage storage alternative is proposed. Each stage will be defined according to the mine production planned by Extorre.

It is important to mention that for the development of this engineering stage, existing information on Geology, Geotechnics and Hydrology was used. For subsequent stages, the referred information will have to be confirmed.

The engineering work developed by Ausenco Vector responds to the requirements and guidelines expressed by Extorre through different communication channels.

18.4.3

Design Criteria

The criteria used for the conceptual engineering design were determined based on the information provided by Extorre. The criteria are mentioned below.

	Process Data

	Plant production rate	1,000tpd
	Tailings specific gravity:	1.185t/m3(Estimated by Ausenco Vector)
	Annual tailings production	365,000 t/year

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	Operation cycle	365 d/year
	Percentage of tailings conducted to the storage facility:	100%

	Required storage (accumulated)

	Stage 1	700,000m3
	Stage 2	1,500,000m3 (Raise)
	Stage 3	2,540,000m3 (Raise)

	Storage Facility life (partial)

	Stage 1	2.6 years (2 years and 7 months)
	Stage 2	2.3 years (2 years and 3 months)
	Stage 3	3.4 years (3 years and 4 months)
	Life at the end of Stage 3: (accumulated)	8.2 years (8 years and 2 months)

	Embankments

	Slopes	2H:1V(it may vary according to subsequent geotechnical analyses of the soils to be used)
	Minimum safety factor for static conditions	1.3
	Minimum safety factor for pseudo-static conditions	1.1
	Freeboard	2m (between crest level and maximum dam level)

	Rain water management within the storage facility

	Conveyance:	open
	Section	trapezoidal
	Minimum slope	0.5 %

18.4.4

Tailing Storage Facilities

Tailings Storage Facility

The Tailings Storage Facility was projected in a natural depression area located southwest of the site determined by Extorre for the process plant. The depression presents moderate to gentle slopes to the south, and is part of a natural runoff system toward an endorheic depression.

The structuring of a storage facility required the projection of at least three embankments so as to restrict the required surface to reasonable technical-economic margins, thus preventing extension of the area to unfeasible values.

At this stage of study, an excavation of at least 0.30 m deep was planned for extraction of top soil, which shall be kept isolated from other materials in a special place to be determined by Extorre. The proposed excavation aims at reducing the undulations of the surface to make the waterproofing system works easier.

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In view of the large amount of material that, according to Extorre, will be available for the construction of the embankments, material from stripping (barren material) and in order not to exceed the excavation volume mentioned above and that corresponding to embankment foundations, the profiling of the facility bottom has not been considered. The profiling of the facility bottom aims at favoring the formation of a recovery water pond in an area determined in advance.

Ausenco Vector’s criterion is to favor the formation of the clear water pond in an area opposite the internal parameter of the embankments, in such a way that the tailings are, during the first instances of operation, disposed behind the embankments. Thus, the tailings will begin their liquid expulsion process and density increase (increasing their similarity with the soil) behind the embankments. This brings about the formation of the water pond in areas near the edges of the facility, which makes water intake works for re-entry into the process considerably easier. At the detail engineering stage, it will be convenient to develop a detailed plan for filling the facility, starting from the discharge of tailings from the dams located in the south.

Dams

The location and elevation that the embankments must have to respond to the design requirements was defined based on the topography obtained by satellite restitution, the information obtained by Ausenco Vector’s staff during the field visit, and the guidelines provided by Extorre.

The embankments location was determined in such a way that together, each of the embankments and their subsequent rises are located in the alignment with the greatest narrowing with respect to the elevation corresponding to the volume requirement. That is to say, the each embankment’s alignment bears relation to the following stage embankment’s alignment, so that both meet at the greatest possible narrowing in order to comply with the storage requirements. This selection criteria aims at reducing soil movements.

Stage 1 requires 3 embankments to make up a 700,000 m3 storage facility at a crest elevation of 61.00 m.At Stage 2, it is necessary to raise the 3 embankments to an elevation of 64.25 m to reach a 1,500,000 m3 capacity, while in Stage 3 the embankments situated to the South are unified and jointly with the Dam 1 rise to an elevation of 67.25 m make up a 2,540,000 m3 storage facility.

Figure 162 presents a cross-section of South Dam 1, identified as Dam 1 in the drawings, which shows the elevations at the different stages of the project.

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Figure 162 Construction stages

The embankments were pre-designed with 2:1 (H:V) slopes, according to the kind of material observed during the field visit and experience in similar projects. It is worth mentioning that according to Extorre’s indications, the material to be used for the construction of embankments will be obtained from mine stripping, and at this design stage, no data were available that described the soils to be used; therefore, after the corresponding studies are conducted, the slopes gradient may vary.

In order to prepare the cost estimate, it is planned to remove soils from a 3.00 m deep layer for the embankments foundation, an area limited by the embankments’ perimeter or “footprint”. The expected foundations depth may vary according to studies that allow defining the soil stratigraphy in the area and that provide geotechnical parameters which define the depth to a competent soil layer.

Dam crests and transition between construction stages

The dam crest for each stage was designed with a total width of 10 m, of which 5.60 m will be passable and the rest are perimeter safety berms and anchoring trenches, as observed in Figure 163. The crest downstream margin safety element can be accomplished by a compacted soil berm or a “guard rail”-type metallic flexible defense.

As part of a good operative practice, a 1-m-wide lined berm was planned at the upstream face of the dam, for the layout of the tailings transportation pipelines. In the pipeline berm sector, a second geomembrane will be placed to protect the lower geomembrane.

Figure 163 Dam crest diagram

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Figure 164 shows in greater detail the geomembrane anchoring trench that will be constructed on the dams’ crests at each stage.

Figure 164 Geomembrane anchoring detail

Figure 165 shows the solution proposed for the transition between the construction stages, which considers removal and replacement of a 0.30 m thick soil layer on the crest and liner binding to the previous stage by means of extrusion welding.

Figure 165 Transition between stages

The welding between the previous phase geomembrane and the new phase will be performed on the downstream slope, in an area after the second sheet placed on the pipeline berm.

The Tailings Storage Facility has been designed to operate as a sub-aquatic storage, in order to avoid acid drainage formation and dust production. Accordingly, a 2 m freeboard is established from the crest level to the tailings deposition maximum level. Within that freeboard, 1 m corresponds to the storage maximum level and the remaining meter, to a safety freeboard.

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Perimeter Service Roads

For each of the defined stages, a service road was planned around the tailings storage facility perimeter during construction and operation. The road final elevation will be the same as the waterproofness limit elevation (or crest elevation), with a total width of 8.00 m and a passable width of 5,50 m at the operation stage.

The storage facility liner anchoring trench and a safety berm will be set out on the road. Figure 165 shows a diagram of the road.

Figure 166 Perimeter road

According to the topography, land profiling will be necessary in some sectors for the construction of the road, which is preliminarily considered with 1.5:1 (H:V) slopes; these values may vary in subsequent stages, in agreement with the corresponding studies.

Waterproofing system

A waterproofing system made up of a compound primary barrier including a 0,30 m thick low-permeability compacted soil layer and a 1.5 mm LLDPE geomembrane was projected. The geomembrane has been pre-selected according to Ausenco Vector’s experience in similar projects; however, the type and/or thickness may vary in subsequent design stages in which parameters are defined in greater detail. Likewise, in later design stages, other materials or a leak detection system may be required.

Surface water diversion system

In order to quantify construction costs, a possible layout was determined considering the site natural drainage network and the location of the tailings storage facilities.

A trapezoidal channel was planned, which intercepts the runoff upstream of the permanent camp and diverts it downstream the deposit, returning it to a natural flowline. The channel diverts the runoff from the North and the West to the South, while towards the East, diversion works were considered unnecessary, as the watershed is close to the storage limits.

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The channel was pre-sized in excavated section in 1.50 m in all its alignment, with a base width of 2.00 m and 2H:1V lateral slopes.

To catch the flows that would be originated in an area not draining to the channel -located between the channel and the storage facility limits- a 2.00-m-deep excavated pond with a capacity of 50,000 m3 was designed. It should be noted that the area previously mentioned is part of the natural topographic depression where the storage facility was projected.

The pond was planned in the sector with the lowest elevation of the depression upstream of Dam 1 (North Dam). The water received can be sent to the plant to be used in the process, or it can be disposed for evaporation. The figure in Annex 2 shows the storage facility general plan and the proposed diversion system.

Control and monitoring system

In order to control dam stability variations, the use of three topographic monoliths has been planned. They will be located on the dam crests; one will be central and two, peripheric, in each of the stages. These points will be controlled with respect to three fix points, performed with three additional monoliths constructed on the deposit zone hillsides.

The control fix points will be located close to the dam crests. These devices will be used to control displacements in horizontal or vertical (settling) planes.

In order to control the waterproofing system performance, two monitoring wells were projected, where samples will be gathered to assess groundwater quality; if an unexpected anomaly occurred, the flow will be returned to the tailings dam or the process plant.

According to the project site seismic characteristics and the small height of the embankments in the projected stages, instruments such as accelerometers, inclinometers, vibrating wire piezometers and open piezometers were not deemed necessary.

If the embankments are raised beyond Stage 3, Ausenco Vector deems it necessary to consider the use of the previously mentioned instruments.

The drawings include a detail of the position for each of the elements of the monitoring system proposed.

18.4.5

Storage Facility Capacity

The design considers tailings disposal in only one storage facility. According to the mine production information received by Ausenco Vector, and a tailings characterization based on Ausenco Vector’s experience and on international bibliography, the dams are projected to have an anticipated disposal capacity according to the volumes mentioned in Table 113.

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Stage	Required accumulated capacity (Ton)	Designed storage accumulated capacity (m3)	Crest elevation (m)	Tailings storage elevation (m)
1	700 000	720,474	61.00	59.00
2	1 500 000	1,536,503	64.25	62.25
3	2.540.000	2,537,040	67.25	65.25

Table 113 Tailings storage facility Capacity

18.4.6

Earthworks Quantities

Table 114 shows the grubbing and embankment volumes for each stage, estimated based on the topography obtained by satellite restitution.

Stage	Embankment volume - partial (m3)	Grubbin volume - partial (m3)	Embankment Volume - Acum. (m3)	Stripping volume - Acum (m3)
1	68,400	36,150	68,400	36,150
2	111,600	49,900	180,000	86,050
3	188,500	60,600	368,500	146,650

Table 114 Earthworks volumes

18.4.7

Required Areas

Table 115 shows the required area for the Tailings Storage Facility and the dams corresponding to each stage. The area corresponds to the surface defined by the perimeter roads and the embankments downstream toe.

Stage	Affected area - Accumulated (m2)
1	290,580
2	391,065
3	468,110

Table 115 Required Areas

18.4.8

Conclusions

The conceptual design engineering was developed with a sufficient level of detail to estimate costs within a range of +/- 30%. This document informs Extorre about the criteria and guidelines used by Ausenco Vector.

Three stages were projected at a conceptual engineering level in order to respond to Extorre’s requirements; they are outlined in Figure 166 to 18.8 below:

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Figure 167 Tailings Storage Facility conceptual view

Figure 168 Dam 1 (North Dam) conceptual view

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Figure 169 Dam 2 and 3 (South Dam) conceptual view

18.4.9

Recommendations

Vector consider appropriate to provide the following recommendations:

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At a subsequent engineering stage, the diversion system individual works design must be developed in detail.

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The embankments’ slopes and foundation must be assessed at a later engineering stage, when geotechnical parameters describing the properties of the soils in the deposit and borrow areas are available. Slopes have been pre-designed according to estimated values and observations during the site visit.

18.4.10

Restrictions and exceptions

This report merely represents the technical conditions identified in and near the premises, exactly as they were at the moment this Tailings Facility Conceptual Design Report, Cerro Moro Project was written, and the conclusions drawn on the basis of the compiled information and the aspects assumed during the evaluation process. This Tailings Facility Conceptual Design Report, Cerro Moro Project is restricted to the scope of the works previously required and executed until the drawing up of this report. The conclusions included in this report represent professional opinion and judgment based on the information studied during the course of this evaluation, not scientific certainties. Limited by the scope of the agreed service, this Tailings Facility Conceptual Design Report was undertaken and carried out in a professional manner, according to the rules of good conduct and generally accepted

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practices, and using the level of proficiency and caution usually exercised by respectable environmental consultants in similar conditions. No other guarantees are provided, whether explicit or implied.

This report shall only be used in complete form. It is based on the scope of the services and is subject to the Restrictions and Exceptions and other limitations herein established. It has been drawn up for the exclusive use of Silex Argentina S.A. and its legal advisors. No other individual or entity is authorized to disclose or rely on any part hereof without the written consent of Vector Argentina S.A. or its legal representative is authorized to assign or authorize the assignment to third parties of this report in full or in part. Nevertheless, any third party using or relying on this report without the express written consent of Vector Argentina S.A., covenants and agrees to have no legal right against Vector Argentina S.A. or against its parent company or affiliates/subsidiaries, or against its consultants and subcontractors, and to indemnify and hold (them) free and harmless from and against any and all claims that may arise from or in conjunction with the said use or assignment.

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19.

MARKETING

Cerro Morro will produce and sell a gold and silver doré to generate revenue for the project.

Over the life of the mine, the process plant will produce doré containing 494.7 koz of gold and 26.7 Moz of silver. The doré will be sold to a refinery for separation into gold and silver bullion.

The annual production for the project is provided in Table 116.

	Au (Ozs)	Ag (Ozs)
Year 1	157,011	6,392,064
Year 2	100,760	4,330,217
Year 3	75,854	3,547,291
Year 4	50,987	3,371,473
Year 5	36,901	3,315,947
Year 6	24,544	2,407,475
Year 7	19,395	1,928,996
Year 8	23,521	1,217,469
Year 9	5,736	176,368
TOTAL	494,709	26,687,301

Table 116 Forecast Annual Production

The dore produced by Cerro Morro can be considered high grade with no impurities that wouldaffect its acceptance by refineries. The anticipated product quality is 95% precious metal and 5% base.

The current proposed mining plan incorporated selective blending of the ores from the various mineralised areas to provide a feed to the plant that will provide consistent high recoveries and low levels of impurities. The composition of the doré may vary depending the blend of ore feed. The various inferred resources from the mineralised zones are as shown in Table 117:

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	Au (%)	Ag (%)
Escondida – Central	38.5	19.2
Escondida – West	0	0
Escondida - Far West	39.1	37
Loma - Escondida	4.7	4
Deborah	0	0
Gabriela	13.4	36.5
Esperanza	4.3	3.3

Table 117 Location of Mineralisation as a percentage of inferred resource

For reference Table 118 from Argor-Heraeus sets out the allowable impurities for major refineries:

Mercury		Not Acceptable
Radioactivity		Not Acceptable
Antimony	<	0.01%
Arsenic	<	0.01%
Bismuth	<	0.01%
Cadmium	<	0.01%
Copper	<	10.00%
Lead	<	0.50%
Molybdenum	<	0.10%
Selenium	<	0.01%
Sulphur	<	0.05%
Tellurium	<	0.01%
Tin	<	0.30%
Zinc	<	0.30%

Table 118 Allowable impurities for major refineries

19.1	Market Outlook

Annual Revenue generated by the project as reported in the base case financial model includes:

●

49 % revenue from Au

●

51 % revenue from Ag

Initially the larger proportion of the revenue is generated from gold sales, but in the latter stages silver generates the greater proportion of revenue. This results in sensitivity to the variability of the market price of both gold and silver.

The charges in market price for gold and silver affect cash flow and profits generated by Estelar. Figure 170 and 19.2 from Kitco show the variability of gold and silver prices from January 2000 until January 2011.

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Figure 170 Gold Pricing – London PM Fix 2000- Present

Figure 171 Silver Pricing – London PM Fix 2000- Present

19.2	Gold Demand

The current demand and supply relationships affects the gold and silver prices, but not necessarily in the same way as the current demand and supply affects the price of other commodities.

Historically, gold has tended to maintain its value compared to basic goods in times of inflation, currency crisis, and macroeconomic events.

Central banks, financial institutions and individuals have historically acquired gold for reserves, and the production for any given year constitutes only a part of the total potential supply of gold. Given that the potential world supply of gold is largely independent of annual mine production, normal variations

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in current production do not necessarily have a significant impact the supply or price of gold, as can be seen in Figure 172 and 19.4.

Figure 172 Gold Annual Total Demand

Figure 173 Gold Annual Total Supply

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19.3	Silver Demand

Strong investment demand, the single most important factor in influencing the price of the metal, is expected to keep silver prices at elevated levels during the projected period.

Fabrication demand for silver is forecast to rise over the projected period, providing additional support to prices along with strong investment demand. Even in present economic conditions there is strong demand for some of the products in which silver is used, including various electronic components used in a full range of consumer and industrial equipment.

Fabrication demand for silver is expected to rise further over the next few years, due to an anticipated improvement in global economic activity coupled with increased use of silver in some of its new and relatively new uses, such as solar panels, silver-zinc batteries. Current silver fabrication demand and Silver total supply are indicated in Figure 173 and 19.5.

Figure 174 Silver Fabrication Demand Rising

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Figure 175 Silver Total Supply

19.4	Available Refineries

The following refineries that are located in Europe and North America and are sufficiently close to the project to be cost effective are :

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Johnson Matthey Limited — Brampton, Canada

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Royal Canadian Mint — Ottawa, Canada

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Argor-Heraeus SA — Mendrisio, Switzerland

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Cendres & Metaux SA — Biel-Bienne, Switzerland

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Metalor Technologies SA — Marin, Switzerland

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PAMP SA — Castel San Pietro, Switzerland

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Valcambi SA — Balerna, Switzerland

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Johnson Matthey Inc — Salt Lake City, United States

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Metalor USA Refining Corporation — North Attleboro, United States.

The availability and suitability of these refineries will be investigated further during the feasibility study.

19.5	Refining Contracts

The following estimated refining costs and recoveries are based on an analysis of existingcontracts and market practice:

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Refining costs per ounce for gold and silver: US$0.25 to 0.30/oz (per Agor)

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Return rate for gold content: 99.85% to 100.00%

●

Return rate for silver content: 99.75% to 100.00%.

Gold and silver sales will be at the precious metal spot prices fixed by the London Metals

19.6	Exchange (LME).

The terms of payment by the refiner are as follows: 99% of the value of the shipment, calculated using Estelar’s assays and the LME spot prices, is due for payment to Estalar the day after the dore is received in the refinery. The 1% outstanding is withheld for adjustment subject to metal content and is due for payment the day following the exchange of the gold and silver assays between the refinery and Estelar. Should a dispute arise with the assay values, samples will be sent to a mutually agreed-upon independent laboratory of verification.

19.7

Shipping

The doré bars produced at Cerro Morro will be stamped, numbered and weighed prior toshipping for refining. The bars will be transported by a security vehicle in a maximum 5,000 kg per shipment to Commodoro Rividavia. From Commodoro Rividavia, the shipment will be air freighted to Buenos Aires for transfer to an international carrier for air freight to the refinery’s location.

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20.	ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

20.1	Environmental Studies

Based on the Environmental Impact Report (Exploration Stage) submitted by Mincorp Exploraciones SA (1997), five Updatings of such document have been submitted between 1997 and 2010 according to File 403.897/M/97 – Province Mining Bureau, Ministry of Economy and Public Works, Province of Santa Cruz.

In September 2010, the Environmental Impact Report (Exploration Stage) was submitted. It was approved by the competent authority in May 2011; thus, they granted the permit to start the construction of the infrastructure required for Cerro Moro Project development and the deposit mining/operation stage.

Baseline environmental studies started in 2008. The consulting company chosen to perform the Environmental Baseline survey was Vector Argentina SA (currently Ausenco/Vector)choosing the environmental consulting company to carry out these studies entailed analysing the following aspects:

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National and international experience of the consulting company in similar jobs performed.

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Consulting company’s experience and background in the Province of Santa Cruz, Argentina.

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Experience of each of the expert consultants in the different subject areas addressed.

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Assessment of the Technical and Economic Proposals submitted by each of the companies which took part in the bid.

In order to perform the Baseline studies, the following environmental aspects were analysed:

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Physical Aspect.Geology and Geomorphology. Seismology. Soils. Hydrology. Water Quality (Surface and Groundwater).Air Quality. Climate. Hydrogeology, ARD.

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Biological Aspect. Ecosystemic Characterization. Flora. Fauna. Limnology.

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Socio-Economic and Cultural AspectArchaeology and Palentology. Landscape. Traffic. Socio-Economic Baseline Study in Puerto Deseado.

The general methods used for performing the Environmental Baseline Study consisted in: to begin with, collection of secondary background; preparation of background and required maps for field visits; field visits (in subject areas such as flora, fauna and water quality, seasonal monitoring has been required), and finally, preparation of Partial Report per specialty.

Estelar is currently performing environmental monitoring activities in the Project’s areas of direct and indirect impact; these include surface and groundwater samples, limnology, meteorology, etc.

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20.1.1

Main Environmental Findings

Physical Aspect

The Region has arid plateau cold-to-warm weather, with annual mean temperatures which do not exceed 16°C.It is severely arid, with annual rainfall under 200 mm. Generally speaking, there are harsh conditions, which are typical of the southernmost part of South America. Summers are cool and winters are intensely cold, slightly moderated by the closeness to the sea, which —especially during the summer— provides humidity thanks to a sea breeze front developed during the day’s central hours along the coast. Evaporation coming from the own meteorological station installed at Cerro Moro was 1600 mm/year, which indicates that there is a serious water deficit.

A geological description which identified and described the regional and local lithological and structural features was prepared. Some of the existing geomorphic units identified in the Project area are: Jurassic ignimbrite environment, La Matilde formation environment, tertiary sedimentary rocks environment, plateau-like relief, endorheic depression relief and alluvial deposits.

The Project is located in a zone where seismic activity is much reduced, classified as Zone 0, according to the National Institute for Seismic Prevention (Instituto Nacional de Prevención Sísmica). The region’s seismic risk, the earthquake occurrence rate and potential seismic sources were estimated with an expected acceleration rock value equal to 0.03 g.

For Air Quality assessment, PM10, NO2, SO2, SH2, and environmental noise concentrations were calculated. These hourly concentrations were extrapolated to equivalent daily and annual “USEPA (2005)” concentrations, and it was noted that base values did not exceed the maximum Environmental Air Quality Standards’ concentration levels set forth in the referenced legislation.

The study area is homogeneous as to the type of soils it has. They are shallow soils, with little evolution, where no horizons can practically be differentiated. On the runoff lines, protected from the wind, material build-up generates profiles deeper than the mean, on which high shrub-like patches develop. Based on the Soil Taxonomy classification, soils were classified in the Aridisoles Order, and in turn, in two suborders: Argides, including great group Natridurids; and Ortides, including great group Cambortides. Soil use capabilities are limited to natural pasture and use by wild fauna. The main limiting factors are the region’s soil useful depth, wind erosion and severe weather conditions. Soils are currently used mainly for sheep farming, with a significant mining development in the region over the last decade.

An Acid Rock Drainage (ARD) research program was carried out for Deborah, Escondida (East, Central and West), Esperanza, Gabriela and Loma Escondida deposits, both for waste rock or hosting rock and process tailings from test works performed, by means of static and kinetic geochemical tests. Static tests (ABA tests) allowed to analyze the types of rocks before performing kinetic tests (Humidity Cell Test), which are more complex and more expensive. Taking into account that the laboratory conditions in which ARD kinetic tests are performed differ largely from those actually existing in the zone, it is estimated that the weather, geological and environmental conditions present in the region under study would substantially inhibit metal solubilization, therefore the process prevailing in the project area would be leaching in the long-run. To this end, for each deposit under study, sulphates and critical metals (those metals that during the tests exceeded the environmental regulations in force) should be monitored considering the pH.

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A Surface and Groundwater Quality Study was performed. It included the Analysis of Physico-Chemical and Microbiological Parameters for 17 sampling stations (8 stations correspond to groundwater and 9 to surface water samples). All analyses were performed pursuant to “Standard Methods for the Examination of Water and Wastewater”, (APHA, AWWA y WPCF), “Methods for Chemical Analysis of Water and Wastes”, as established by EPA/600/4-79-020. The main aspects to be taken into account are as follows:

Surface Waters:

●

In the project area there are only surfaces drainage network which channels waters towards endorheic saline depressions. These collect surface runoff in winter, and are an evaporation area in spring-summer, naturally concentrating salts, with neutral to alkaline pH. Due to strong evaporation processes and little rainfall, there is salt concentration throughout the spring-summer period until it dries up.In the monitored endorheic depressions there are relatively high values of fluorides, boron, manganese, and vanadium.

●

With regard to the Deseado River, it has a seasonal regime all through its run.At “Paso Gobernador Gregores” bridge, for instance, the river has water in winter, but remains dried up for the rest of the year. Downstream, to 40 km approximately, east-westward, the river water comes from the elevation of tides, which are typical in the zone (seawater entering the continent).

Groundwater

We can mention two groundwater types or qualities:

●

Water extracted from mills to provide for the hacienda and for local people's consumption (6 sampling points) is taken from 5 to 20 m deep, and is water stored in alluvial sands and silts which fill the most depressed areas of basins and valleys, creating surface aquifers. This water has a relatively low saline content, with an electrical conductivity from 1,500 to 5,000 µS/cm. As for the anionic composition of these waters, they mainly have chloride content and bicarbonate content as well. Cationic composition is mainly sodic, and calcic to a lesser degree.

●

The deepest waters (samples from mining holes) (2 sampling points in Escondida and Déborah) catch the water in greater depths (over 40 m) and correspond to rock aquifers with secondary porosity and high salinity, around 7,000 µS/cm for the Déborah well and 15,000 µS/cm for the Escondida well. They primarily have sodium chloride content.

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Biotic Aspect

Fauna was characterized by means of seasonal survey campaigns. The wild fauna living in the Project area corresponds - biogeographically - to the Neotropical Region, Andean-Patagonian Dominion and within this, to the Patagonian Province. The latter is known for its ways of adapting to extreme living conditions, namely a severe aridity and very low temperatures in winter. The severe weather conditions deeply influence the temporal distribution patterns of the main vertebrate groups; this is most obvious in birds. The region's most typical reptile, bird, and mammal species were observed in the Project's area.

●

Among large and micro mammals, 16 out of the 68 species mentioned for the province of Santa Cruz (native and exotic) were found. Guanacos (Lama guanicoe) were the most representative, accounting for 23.5% of the total. The main carnivore species is the gray fox (Pseudalopex griseus). This species' population is still in considerable recovery given the lack of hunting and poisoning.In turn, the guanaco and choique (Rhea pennata) groups' behavior is consolidating in the area and thus they are increasing in number.

●

Among birds, 40 out of the 205 species listed for the province where found; they account for 19.5% of the total. Among the species linked to water bodies, the cauquén real (Chloephaga poliocephala) was the most abundant; together with the playerito unicolor (Calidris bairdii) and the monjita chocolate (Neoxolmis rufiventri), they are the 3 migrant species of the study area.

●

With regard to herpetofauna, 7 of the species listed as endemic were observed in the study area.

Flora was identified during four seasonal campaigns. Vegetation was characterized phytosociologically, and the data was presented in a vegetation map which included vegetation units of the Project's area, fodder availability, grazing land condition, monitors, and a complete list of species. In the Project's area there are three basic vegetation communities: desert, barren land or sub-shrub steppe, and shrub-steppe. Saline depressions or lakes with little or no water account for around 1% of the Project's area. The following Table shows a brief description of the composition of these communities and a picture of the community.

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Desert
This vegetation unit is mainly located in the Project's S and NW areas. These are places where vegetation is scarce and scattered. Topsoil values range from 5% to 10% and are mainly represented by sub-shrubs. This vegetation unit accounts for 11% of the Project's area.

Sub-shrub steppe
This is the predominant vegetation unit within the study area. Topsoil accounts for 30%-40% of the area. The remaining area is covered by pebble paving.In these places, the soil has been completely lost, and the surface is covered by gravel. Leaf litter is also lost; there is only some dead material standing. This vegetation unit accounts for 67% of the Project's area.

Sub-shrub steppe
Topsoil accounts for 46% of the area.The predominant life form is shrubs, covering 26%. The dominant species is Senecio filaginoides; others are Berberis heterophylla, Nardophyllum obtusifolium, Schinus sp., Lycium ameghinoi. This vegetation unit accounts for 21% of the Project's area.

Shrub-steppe

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Socio-economic and Cultural Aspect

For socio-economic and cultural factors, the Area of Direct Impact (ADI) was defined as the towns linked to the Project (Puerto Deseado, Tellier, Jaramillo, and Fitz Roy) and the communication paths, considering commercial interactions (demand for goods and services, source of human resources, etc.). The area of indirect influence (AII) for socio-economic and cultural factors would be comprised of the towns of Comodoro Rivadavia, Puerto San Julián, Pico Truncado, Las Heras, and Río Gallegos. Moreover, it includes Jaramillo and Tellier, although the Project's activities will virtually not impact the latter, except for the impact on the workforce, for a lack of logistic availability.

The town nearest to the Project’s area is Puerto Deseado.This town has a reduced population (according to 2001 Census, 10,237 inhabitants, reaching 12,963 inhabitants in 2009). Its main commercial structure relies on retailing, which is focused on the supply and sale of goods to satisfy such basic needs as food, clothes and cleanliness, and supplies for dwellings and motor vehicles. This town’s predominant production is basically generated by the industrialization of fishing products and fishing-related services. Although tourism constitutes a significant activity, it is not reflected on the town’s commercial and service structure.In the rural area, livestock breeding is the major activity, and mining has sensibly expanded during the last decade.

Other minor towns nearby the Project’s area are: Tellier, Jaramillo and Fitz Roy, located at 20, 121 and 142 km away, respectively. Tellier (~80 inhabitants), Jaramillo (~220 inhabitants) and Fitz Roy (~180 inhabitants) have been recently urbanized; their population is scarce, located inward and originated in the railway branch line. These are towns that have been “stagnant” for the last three decades, with the same absolute population values, which precisely account for a "vegetative existence", with no dynamism at all. As an exception, it should be noted that there is a “floating population” associated with the personnel transferred by contracting companies from current road works.

There are no natural protected areas in the Project’s area.

The landscape was assessed in terms of quality, frailty, and visual accessibility. The following Landscape Units were identified: El Espinazo, El Mosquito Lake, Plateau Plain, Lakeside Hills, Quebrada de Gabriela, Valleys, and Hills.

With regard to cultural aspects, different archaeological and paleontological survey campaigns were carried out. The archaeological findings in the study/work area, obtained as a result of the prospecting activities performed, can be - operatively - grouped in three categories:sets of lithic material or isolated lithic findings; caves and/or rock shelters; and structures possibly linked to funerary activities. Outside the Project's area of impact, 5 archaeologically-relevant places were identified: one is a cave or rock shelter with pictographs inside, North of the Project; two are places with lithic material concentrations which could correspond to lithic material extraction quarries; and two are possible sites related to funerary activities (chenques). All these places are far away from the infrastructure facilities proposed. Four sites with paleontological findings have been defined; three of them are not related to the Project's facilities. The fourth site is next to the Prospecto Esperanza SE’s facilities, where sea macrofossils remains have been found.

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20.2	Permitting

The following Chart shows the schedule of the filing of the Environmental Impact Assessment (EIA), granting and notification of the Environmental Impact Statement (EIS) and consequently the expiry date of the environmental permits obtained for the Cerro Moro Project. Further, it is indicated whether the permit corresponds to an exploration and/or development.

Company	Eia Filing	Eis Grant	Eis Notification	Eis Expiry	Project Stage
MINCORP EXPL. SA	23/05/97	23/06/97	15/07/97	15/07/99	Exploration
MINCORP EXPL. SA	29/06/99	N/D	N/D	N/D	Exploration
CERRO VANGUARDIA SA	08/04/05	20/02/06	24/02/06	24/02/08	Exploration
CERRO VANGUARDIA SA	08/04/05	20/02/06	24/02/06	24/02/08	Exploration
ESTELAR RESOURCES LTD.	09/10/08	30/03/09	28/05/09	28/05/11	Exploration
ESTELAR RESOURCES LTD.	October 2010	13/12/10	15/12/10	15/12/12	Advanced Expl.
ESTELAR RESOURCES LTD.	17/09/2010	16/05/11	16/05/11	16/05/13	Development

20.3	Environmental Impact

While most of the generated environmental impacts are moderately significant, they are localized and may be reversed or mitigated in the short, mid and long term.They are detailed below:

●

Decreased air quality as a result of the emission of particulate matter and combustion gases.

●

Increased level of existing noise due to higher sound pressure generated by the operation of equipment, machinery and vehicles, blasting, and the operation of the Process plant.

●

Alteration of vegetation processes and its population dynamics, due to the emission of dust and environmental fragmentation resulting from the presence of the components of the Project, thus altering the habitat of the land fauna as a consequence of the noise, vibration, introduction of barriers and human presence.

●

Worsening of erosive processes due to soil movements (grubbing, excavations, and stockpiled barren material) generated during the construction of the proposed infrastructure, pit opening, and waste dump development.

●

Caving and subsidence processes as a result of ore extraction in veins (underground mines).

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Quality of surface and underground water will be affected due to accidental spills during operation of mining equipment, transport, handling and storage of process supplies.

●

On archaeological resources and paleontological heritage, due to the risk involved in soil movement activities.

●

On road infrastructure, due to the deterioration that may be caused by increased vehicle traffic on the haul road.

●

The sense of wellbeing of the population will be affected, and the increased risk of accidents and/or the sense of insecurity as a consequence of the increased vehicle traffic generated by the activities related to the transport of material, supplies, finished product and staff from and/or to the Project location on the haul road. The city affected will bePuerto Deseado and the towns of Tellier, Jaramillo and Fitz Roy, on National Road No. 281.Such disturbances will be mainly the result of the emissions of noise, vibration and particulate material produced by the increased traffic on Provincial Road No. 47.

●

On water resources, given the characteristics of the region where the Project is located, this is an environmentally critical factor. Regarding this environmental factor, the Project has considered the sustainable extraction from the aquifer by calculating the safe flow rate extraction, in order to avoid any impact on the current users of this resource. This flow rate is the volume that may be extracted from the total reserves consumed by the replenishment flow plus a part for storage (hydro-geological reserves), taking levels to sustainable values in the mid term, without affecting water catchment points for cattle.

The above-mentioned impacts are non-significant because their occurrence rate is medium or low, since it is possible to effectively implement preventive control measures on the activities entailed; these impacts may be reversed or mitigated immediately.

Others, which, as a result of their permanence and reversibility, will cause residual impacts on environmental factors, are:

●

Topography alteration, generated by the deepening of open pits and barren material stockpiled, and storage of process tailings from the plant.

●

Loss of soil as a resource and modification of its current use, loss of topsoil and destruction of the habitat for land fauna, as a result of open pits excavation and barren material stockpiled, and process tailings.

The above-mentioned impacts are considered moderately significant on the surfaces affected, although their extent is partial or local in the frame of a regional context, and they can’t be reversed.

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20.4	Social and Community Impact

The impacts caused from a socio-economic perspective are detailed below.

●

The Project development will bring about social changes on the local community, which entail negative impacts, such as changes in the urban centers’ profiles resulting from the pressure of immigration, which also has effects such as temporary shortage of dwellings and the resulting increase in the value of real estate and rents; increase in social problems related to health and safety due to higher income, and to the circulation of a higher number of people who do not belong to the communities related to the Project; together with deficiencies in the availability of education and health infrastructure.

The above-mentioned impacts are considered moderately significant, due to the fact that the affected factor holds a high environmental value and to the permanence of the impact.

●

The development of the Project will cause a mildly significant positive impact on employment levels, as a result of the demand of direct and indirect jobs that will bring dynamism to the provincial and local markets; the community gives a higher value to the benefits obtained for such employment of individuals belonging to the relatively poorer socio-economic levels, which would be the benefited local residents.

●

The economic factor will be positively impacted, given that the Project stands out for the income and wealth generated by the commissioning phase for the whole production circuit. These activities involve a gross production value in which the value added will be of great relevance, because during operation only a minor proportion of the investment will go towards buying supplies and services. The Project will generate a substantial retribution in terms of value added, since it will generate an important proportion of wealth through taxes and royalties withheld by the jurisdiction involved and the wages paid to employees, among others.

The above-mentioned positive impacts with residual effect are moderately significant in the area of direct influence (Puerto Deseado) and the area of indirect influence (Santa Cruz and the Argentine Republic).

20.5	Environmental Management Plan

The Environmental Management Plan developed by Estelar states the preventive, control and mitigation measures for environmental and socio-economic impacts resulting from the execution of the Project, in a framework of a series of plans and programs which will be complied with by Estelar and the contractors. The prime objective is to comply with the legal environmental regulations of Argentina and with Estelar’s environmental policy.

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The Environmental Management Plan is based on the objectives of ensuring that, by preparing documented procedures, negative environmental impacts that may be caused during the different phases of the Project, be they of construction, operation or closure, are prevented, controlled, minimized and mitigated, while positive impacts on the socio-economic and technological spheres are promoted, thus forging a good relationship with the community.

The Environmental Management Plan is made up of the following Prevention, and Control and Preservation Measures Programs:

Prevention Programs developed:

●

Safety and Occupational Health Program.

●

Contingency Plan.

●

Community Relations Plan.

●

Monitoring, Control, and Follow-up Program.

●

Environmental Training Program.

●

Affected Areas Recovery Program.

Control and Preservation Measures developed:

●

Measures to control and mitigate atmospheric emissions.

●

Measures to control and mitigate noise emissions.

●

Measures to control erosion and settlement.

●

Measures to control the physical stability of the Project components (open pits, underground mines, and waste dumps).

●

Measures to preserve water resources.

●

Measures to control hazardous materials transport, storage, and handling.

●

Waste management measures.

●

Biodiversity protection measures.

●

Cultural heritage preservation measures.

●

Measures to preserve soil and soil use.

●

Measures of landscape value preservation control.

Exeter has developed the Conceptual Closure Plan (CCP), for all its components, to be implemented progressively in the different stages, from construction, operation and closure of the Project. The main objective of the Conceptual Closure Plan is to ensure that all areas where mining or ore processing activities took place are restored in such a way that they provide adequate public safety, and a use similar to the original one to the ground affected by mining activities.

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21.	CAPITAL AND OPERATING COSTS

21.1

Capital

The overall capital costs for the mining and process plant are summarised in Table 119:

Category (MUS$)	Initial Capital	Sustaining Capital	Total Capital
Direct Costs
Open pit mining Equipment	22.9	7.2	30.2
Pre-strip (Escondida FW)	12.8		12.8
Underground mining equipment & Development & Infra/Services	20.0	32.0	52.0
Plant	67.2		67.2
Tailings Storage Facility	7.2		7.2
Water Borefield	5.6		5.6
Power Supply	17.9		17.9
Camp facilities	10.2		10.2
Access roads	3.3	8.5	11.8
Mobile Equipment	2.8		2.8
Fuel Storage	1.4		1.4
Site Buildings	1.5		1.5
Assay Laboratory	2.4		2.4
Communications	0.3		0.3
Closure		5.0	5.0
Subtotal Direct Costs	175.5	52.7	228.2
Undirect Costs
EPCM, Commissioning, Temporary facilities, Initial Fill and Spares	28.8		28.8
Owners Costs	4.0		4.0
Subtotal Undirect Costs	32.8		32.8
TOTAL	208.4	52.7	261.1

Table 119 Overall Capital Costs

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21.1.1

Process Plant and Infrastructure

General

The capital cost estimate is quoted to an accuracy level of ±15%. The following qualifications are made:

●

The estimate is in United States of America dollars USD;

●

The estimate is only for capital costs from the commencement of the process plant design and construction award; all costs prior to this are excludedincluding exploration, pre-feasibility and feasibility related studies;

●

The estimate excludes all exploration and development drilling expenditures. Extorre’s “Owners costs” for project development are included, but Extorre’s corporate costs are excluded;

●

Charges related to financing (e.g., fees, consultants, etc.) are excluded;

●

Capitalised interests and standby fees from third-party lenders are excluded;

●

There is no escalation added to the estimate.

Capital cost estimates for the process plant and associated infrastructure have been based upon:

●

Preliminary engineering including process design criteria, quantity take-offs;

●

Budget price quotations (for major equipment) and current cost data (for the remaining equipment and material);

●

Unit rates have been based on budget prices received from subcontractors and fabricators in Argentina.

●

The costs of engineering, procurement, construction management and commissioning have been estimated from knowledge on similar projects.

●

topographical information obtained from site survey provided by Extorre

●

Preliminary geotechnical investigation information provided by Extorre

Summary

The capital cost estimate for the processing plant and associated infrastructure are summarised in Table 120 by discipline. Table 121 provides a summary of the capital cost by process plant area.

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Item	Materials $	Install $	Freight $	Cont’ncy $	Subtotal $	VAT $	TOTAL $
Earthworks	5,274,739	930,268	33,350	670,349	6,908,706	1,450,828	8,359,535
Civil works	3,727,263	2,679,131	8,642	775,460	7,190,495	1,510,004	8,700,499
Buildings	8,665,810	901,724	623,324	1,064,172	11,255,030	2,363,556	13,618,586
Mechanical equipment	26,085,667	2,140,599	1,208,560	3,066,949	32,501,774	3,867,345	36,369,119
Platework	3,871,896	1,478,544	203,881	629,359	6,183,680	1,298,573	7,482,253
Structural steel	4,972,995	1,154,533	580,927	728,572	7,437,027	1,561,776	8,998,803
Piping	2,909,045	1,308,017	41,006	491,208	4,749,275	997,348	5,746,623
Electrical installations	18,057,887	1,511,193	241,740	2,056,642	21,867,461	4,478,990	26,346,451
Construction equipment	2,513,400	1,029,635	172,000	422,985	4,138,020	868,984	5,007,004
Total Direct Costs	76,078,700	13,133,643	3,113,431	9,905,696	102,231,469	18,397,404	120,628,873
INDIRECT COSTS
Construction facilities	1,528,115	1,411,482	50,750	369,609		705,591	4,065,547
Temporary construction facilities	234,720	1,411,482	50,750	240,269	1,937,221	406,816	2,344,037
Mobilisation and demobilisation	1,293,395	0	0	129,340	1,422,735	298,774	1,721,509
EPCM Costs	6,026,030	14,224,670	0	2,025,070	22,275,770	228,755	22,504,525
Commission, spares & initial fills	1,837,539	1,686,793	132,229	365,656	4,022,216	236,299	4,258,515
Owners cost	3,639,667	0	0	363,967	4,003,633	311,850	4,315,483
Total Indirect Costs	13,031,351	17,322,944	182,979	3,124,302	33,661,576	1,482,494	35,144,070
TOTAL	89,110,051	30,456,587	3,296,409	13,029,997	135,893,045	19,879,898	155,772,943

Table 120 Capital Cost Summary by Discipline

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Description	Equip / Matl Cost$	Labour Cost$	Freight Cost$	Contingency $	TOTAL $	VAT $	TOTAL INCL VAT $
AREA 200 - BULK EARTHWORKS & DAMS	729,400	921,934	11,600	212,390	1,875,324	393,818	2,269,142
AREA 310 - CRUSHING	6,250,286	1,166,738	320,599	832,099	8,569,722	1,277,508	9,847,230
AREA 320 - ORE STORAGE	1,833,381	798,690	87,020	311,844	3,030,935	559,244	3,590,180
AREA 330 - GRINDING AND CLASSIFICATION	4,749,077	661,859	441,903	618,377	6,471,216	635,518	7,106,734
AREA 336 - CONCENTRATE TREATMENT	1,876,311	440,123	77,510	261,401	2,655,345	328,178	2,983,523
AREA 334 - CONCENTRATE REGRIND	529,873	68,777	12,477	64,552	675,679	64,115	739,793
AREA 340 - LEACH & ADSORPTION	3,157,401	1,308,116	152,379	536,703	5,154,600	868,052	6,022,651
AREA 341 - CCD CIRCUIT	3,195,200	631,901	124,921	432,595	4,384,617	516,115	4,900,732
AREA 350 - GOLD RECOVERY	2,580,601	403,266	290,672	347,617	3,622,157	319,212	3,941,369
AREA 360 - REAGENTS	1,758,663	493,696	109,711	260,892	2,622,962	476,765	3,099,727
AREA 370 - POWER AND RETICULATION	4,018,429	1,518,689	241,740	653,820	6,432,678	1,237,685	7,670,364
AREA 390 - WATER STORAGE	861,359	126,065	41,468	109,192	1,138,084	238,998	1,377,082
AREA 400/402 - TAILINGS	1,356,287	234,786	68,699	178,847	1,838,619	246,320	2,084,939
AREA 420 - COMPRESSED AIR	298,733	38,208	26,129	38,217	401,287	30,644	431,930
AREA 499 - PLANT PIPING	3,096,582	1,377,756	60,297	522,351	5,056,987	1,061,967	6,118,954
AREA 804 - CONSTRUCTION EQUIPMENT	2,513,400	1,029,635	172,000	422,985	4,138,020	868,984	5,007,004
DIRECT COST	38,804,983	11,220,239	2,239,125	5,803,883	58,068,230	9,123,124	67,191,354
AREA 201 - ACCESS ROADS	2,479,339	0	0	247,934	2,727,273	572,727	3,300,000
AREA 205 - MOBILE EQUIPMENT	2,066,000	8,334	21,750	210,025	2,306,109	484,283	2,790,392
AREA 375 - POWER SUPPLY (OHP LINE)	14,049,587	0	0	1,404,959	15,454,545	3,245,455	18,700,000
AREA 391 - WATER SUPPLY (BOREFIELD)	3,405,000	596,090	205,000	450,413	4,656,503	977,866	5,634,369
AREA 401 - TAILINGS STORAGE FACILITY	5,398,755	0	0	539,875	5,938,630	1,247,112	7,185,742
AREA 410 - FUEL STORAGE	851,864	142,809	24,232	109,031	1,127,935	236,866	1,364,801
AREA 430 / 440 - SITE BUILDINGS	875,696	173,840	105,800	124,226	1,279,561	268,708	1,548,269
AREA 460 - ASSAY LABORATORY	1,787,006	27,380	0	182,808	1,997,194	419,411	2,416,604
AREA 480/481/485 - CAMP FACILITIES	6,171,721	964,951	517,524	813,667	8,467,863	1,778,251	10,246,115
AREA 490 - COMMUNICATIONS	188,750	0	0	18,875	207,625	43,601	251,226
TOTAL INFRASTRUCTURE	37,273,717	1,913,404	874,306	4,101,813	44,163,239	9,274,280	53,437,519
AREA 500 - ENGINEERING	6,026,030	14,224,670	0	2,025,070	22,275,770	228,755	22,504,525
AREA 510 - COMMISSIONING	459,000	1,686,793	0	214,579	2,360,372	106,029	2,466,401
AREA 550/560 - INITIAL FILLS AND SPARES	1,378,539	0	132,229	151,077	1,661,845	130,270	1,792,115
AREA 600 - TEMPORARY FACILITIES	1,528,115	1,411,482	50,750	369,609	3,359,956	705,591	4,065,547
AREA 700 - OWNERS & PRE-PRODUCTION	3,639,667	0	0	363,967	4,003,633	311,850	4,315,483
TOTAL INDIRECT COSTS	13,031,351	17,322,944	182,979	3,124,302	33,661,576	1,482,494	35,144,070
TOTAL COST ESTIMATE	89,110,051	30,456,587	3,296,409	13,029,997	135,893,045	19,879,898	155,772,943

Table 121 Capital Cost Summary by Area

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21.1.2

Process Plant and Infrastructure Sustaining Capital

The main are of sustaining capital for the process plant and infrastructure is for the tailings storage facility stages two and three which will occur in years three and six of operation respectively.

The estimated sustaining capital costs for these stages are $3,550,000 for stage two and $5,500,000 for stage three.

21.1.3

Mining

Mine capital costs estimate was divided into open pit and underground. Using the developed mine schedule, equipment requirement was estimated and underground developments.

The initial mine capital for the open pit amounts to US$ 22.9 million, from which US$ 16.4 million is for main equipment, US$ 1.2 million for support equipment and US$ 5.4 million in other investments, as detailed in Table 122. An additional US$ 7.2 million is required for the open pit as sustaining capital due to fleet increase.

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Period name	%VAT		Construction	1	2	3	4	5	TOTAL
Main Equipment		MUS$	16.41	5.09	-	-	-	1.22	22.72
Loading	21%	MUS$	3.50	1.27	-	-	-	-	4.77
Hauling	21%	MUS$	6.11	1.22	-	-	-	1.22	8.55
Drilling	21%	MUS$	1.82	1.82	-	-	-	-	3.63
Ancillary	21%	MUS$	4.99	0.79	-	-	-	-	5.77
Support		MUS$	1.16	0.06	-	-	0.06	-	1.27
Backhoe	21%	MUS$	0.36	-	-	-	-	-	0.36
Fuel Truck	21%	MUS$	0.12	-	-	-	-	-	0.12
Lube Truck	21%	MUS$	0.17	-	-	-	-	-	0.17
Support Truck	21%	MUS$	0.11	-	-	-	-	-	0.11
Mobil Crane	21%	MUS$	-	-	-	-	-	-	-
Lowboy Truck	21%	MUS$	0.30	-	-	-	-	-	0.30
Tires Handler	21%	MUS$	-	-	-	-	-	-	-
Lightning Plant	21%	MUS$	0.09	0.06	-	-	0.06	-	0.20
Other Investments		MUS$	5.37	0.07	-	0.12	0.61	-	6.17
Small backhoe 0,5 - 1 yd3	21%	MUS$	0.13	-	-	-	-	-	0.13
Sump Pumps	21%	MUS$	0.29	0.07	-	-	-	-	0.36
Pickups	21%	MUS$	0.61	-	-	-	0.61	-	1.21
Explosives Truck	21%	MUS$	0.12	-	-	-	-	-	0.12
Forklift 7t	21%	MUS$	0.17	-	-	-	-	-	0.17
Manlift 25t	21%	MUS$	0.15	-	-	-	-	-	0.15
Compactor	21%	MUS$	0.36	-	-	-	-	-	0.36
Computer Equipment (Tech Service)	21%	MUS$	0.12	-	-	0.12	-	-	0.24
Mining - Geology Soft Ware	21%	MUS$	0.42	-	-	-	-	-	0.42
Furnitures	21%	MUS$	0.12	-	-	-	-	-	0.12
Radio Equipment	21%	MUS$	0.07	-	-	-	-	-	0.07
Spare parts (6% Main Equip.)	21%	MUS$	1.05	-	-	-	-	-	1.05
Consulting / Engineering	21%	MUS$	0.73	-	-	-	-	-	0.73
Survey Equipment	21%	MUS$	0.36	-	-	-	-	-	0.36
Geology & Geotechnical Equipment	21%	MUS$	0.48	-	-	-	-	-	0.48
Buses	21%	MUS$	0.18	-	-	-	-	-	0.18
Total Open Pit Capital Cost		MUS$	22.93	5.22	-	0.12	0.66	1.22	30.16

Table 122 Open Pit Capital Cost

Underground mine capital cost is detailed in Table 123 through Table 126. The estimate had been divided into development, equipment and infrastructure & services. The initial capital cost amounts to US$ 20.1 million on equipment, because even that they are required during the first year needed to be ordered one year before. Additional US$ 32.0 million is required as sustaining capital, mainly because of development requirement and fleet increase and substitution.

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Year	Development	Equipment	Infrastructure and Services	Total
0	-	20.01	-	20.01
1	6.27	7.99	2.22	16.47
2	10.43	0.80	0.42	11.65
3	1.99	0.73	0.20	2.92
4	0.93	-	-	0.93
5	-	-	-	-
6	-	-	-	-
7	-	-	-	-
TOTAL	19.62	29.52	2.83	51.97

Table 123 Underground Capital Cost Summary (MUS$)

Developments (m)	US$/m	Year 1	Year 2	Year 3	Year 4
Horizontal Developments	1,018.79	4,882.58	6,282.17	1,213.33	657.58
Vertical Developments (RB)	4,009.50	322.00	1,006.00	188.00	64.00
TOTAL	MUS$	6.27	10.43	1.99	0.93

Table 124 Development Capital Cost

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Equipment	US$/un		Year 0	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9
LHD 6 yd3		Requirement		2	3	1	1	1	1	1
LHD 6 yd3	$ 680,000	Adquisition	2	1
LHD 3.5yd3		Requirement		3	3	4	4	3	3	2	1	1
LHD 3.5yd3	$ 560,000	Adquisition	3		1
Trucks 30t		Requirement		4	7	4	4	3	3	2	1	1
Trucks 30t	$ 520,000	Adquisition	4	3
Jumbo horizontal		Requirement		6	8	3	2	2	2	2
Jumbo horizontal	$ 620,000	Adquisition	6	2
Jumbo Radial		Requirement		1	1	1	1	1	1	1	1	1
Jumbo Radial	$ 750,000	Adquisition	1
Jumbo support		Requirement		3	6	2	1	1	1	1
Jumbo support	$ 680,000	Adquisition	3	3
Explosives charger		Requirement		2	4	1	1	1	1	1
Explosives charger	$ 250,000	Adquisition	2	2
Others
Compressores 7cfm	$ 79,000		1
Main Fans	$ 100,000		5
Secondary Fans	$ 30,000		7	7
Service truck	$ 95,000		1	1
Fuel truck	$ 100,000		1			1
Lube truck	$ 140,000		1			1
Explosives truck	$ 100,000		1
Lighthning plant	$ 12,000		6			3
Pickups	$ 50,000		10			5
Plataforma de levante	$ 380,000		3
Mixer 6 m3	$ 420,000		1
Minibus	$ 75,000		3			1
Gunning machine	$ 275,000		1
Water truck	$ 100,000		1
Scaler	$ 175,000		2	1
Auxiliary truck	$ 100,000		1	1	1
Subtotal	MUS$		16.54	6.60	0.66	0.60	-	-	-	-
VAT	21%		3.47	1.39	0.14	0.13	-	-	-	-	-	-
TOTAL	MUS$		20.01	7.99	0.80	0.73	-	-	-	-	-	-

Table 125 Underground Equipment Capital Cost

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Infrastructure and Services
US$	1	2	3	4	5	6	7	8
Surface workshop	$ 550,000	100%
Underground workshop	$ 110,000	100%
Magazine	$ 88,000	100%
Backfill plant	$ 275,000	100%
Portals (3)	$ 381,392	100%
Fire refugies	$ 440,000	50%	50%
Transformers	$ 330,000	60%	20%	20%
Cables, protections, etc	$ 440,000	60%	20%	20%
Comunications	$ 220,000	60%	20%	20%
TOTAL	MUS$	2.22	0.42	0.20	-	-	-	-	-

Table 126 Underground Infrastructure and Services Capital Cost

21.2

Operating

21.2.1

Process Plant

General

The operating cost estimate is quoted to an accuracy level of ±15%.The following qualifications are made:

●

The estimate is in United States of America dollars USD;

●

The estimate is only for the process plant and associated borefield operating costs. No allowance has been made for any corporate overhead or offsite costs;

●

No allowance has been made for offsite bullion transport and refining costs;

●

There is no escalation in the estimate.

Operating cost estimates for the process plant and associated borefield have been based upon:

●

Budget price quotations for reagents and grinding media;

●

Reagent and grinding media consumptions based on testwork;

●

Mining schedule provided by Extorre dated 24th June 2011.

Summary

The operating cost estimate on an annual basis is summarised in Table 127.

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Production	Y00	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08
Tonnes Treated (t)	307,901	335,752	335,662	335,908	335,821	359,893	335,867	332,427	72,391
Gold (g/t)	16.7	9.8	7.4	5.0	3.6	2.4	1.9	2.3	2.5
Silver (g/t)	733.8	455.8	373.5	354.8	349.0	253.3	203.0	129.5	84.5
Cost Centre	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)
Cost Centre Labour	17.67	16.20	16.21	16.20	16.20	15.12	16.20	16.37	75.15
Power	5.82	5.34	5.34	5.33	5.34	4.98	5.34	5.39	10.46
Regents and Grinding	44.60	35.93	34.39	33.82	33.64	30.39	29.53	27.49	31.32
Media Maintenance	3.78	3.47	3.47	3.46	3.47	3.23	3.46	3.50	16.08
Linings	1.38	1.27	1.27	1.27	1.27	1.18	1.27	1.28	1.18
Other	8.40	7.77	7.78	7.77	7.77	7.31	7.77	7.84	32.94
Total	81.66	69.98	68.44	67.85	67.68	62.22	63.57	61.87	167.13

Table 127 Operating Cost Estimate by Cost Centre

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Production	Y00	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08
Tonnes Treated (+)	307,901	335,752	335,662	335,908	335,821	359,893	335,867	332,427	72,391
Gold (g / +)	16.7	9.8	7.4	5.0	3.6	2.4	1.9	2.3	2.5
Silver (g / +)	733.8	455.8	373.5	354.8	349.0	253.3	203.0	129.5	84.5
Cost Area	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)	Unit Cost ($/t)
Crushing and Screening	1.96	1.80	1.80	1.80	1.80	1.68	1.80	1.82	5.40
Grinding and Classification	9.95	9.56	9.56	9.56	9.56	9.27	9.56	9.60	12.51
Regrind and Leaching	1.17	1.09	1.09	1.09	1.09	1.02	1.09	1.10	3.00
Leaching and CCD	9.91	9.81	9.81	9.80	9.81	9.73	9.80	9.82	13.74
Tailings Thickening and Disposal	7.79	7.76	7.76	7.76	7.76	7.73	7.76	7.76	8.90
Gold and Silver Recovery	24.51	15.72	14.18	13.61	13.43	10.10	9.32	7.30	15.24
Reagents Mixing and Distribution	0.41	0.38	0.38	0.38	0.38	0.35	0.38	0.38	1.50
Water and Air Services	1.92	1.83	1.83	1.83	1.83	1.76	1.83	1.84	4.58

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Workshop	0.56	0.51	0.51	0.51	0.51	0.48	0.51	0.52	2.38
Laboratory	0.47	0.43	0.43	0.43	0.43	0.40	0.43	0.44	2.01
Administration	23.01	21.10	21.11	21.09	21.10	19.69	21.09	21.31	97.87
Total	81.66	69.98	68.44	67.85	67.68	62.22	63.57	61.87	167.13

Table 128 Operating Cost Estimate by Cost Area

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Operating Cost Centres

Labour

The labour cost has been estimated at $5,440,500 for a full production year. Table 129 provides a breakdown and details of the labour cost estimate for a full production year.

Position	Number	Position Salary ($)	Total Salary ($)	On Cost Factor	Total Cost ($)
Production
Processing Plant Manager	1	300,000	300,000	1.35	405,000
Production Superintendent	1	250,000	250,000	1.35	337,500
Plant Metallurgist	2	200,000	400,000	1.35	540,000
Laboratory Technicians	2	50,000	100,000	1.35	135,000
Panel Supervisor (shift)	4	80,000	320,000	1.35	432,000
Process Technician (shift)	24	50,000	1,200,000	1.35	1,620,000
Goldroom Supervisor (day)	1	80,000	80,000	1.35	108,000
Process Technician (day)	8	50,000	400,000	1.35	540,000
Maintenance
Maintenance Superintendent	1	250,000	250,000	1.35	337,500
Maintenance Planner	1	50,000	50,000	1.35	67,500
Fitter	4	50,000	200,000	1.35	270,000
Boilermaker	4	50,000	200,000	1.35	270,000
Mechanical Apprentice	2	25,000	50,000	1.35	67,500
Electrical Leading Hand	1	80,000	80,000	1.35	108,000
Electrician/Instrument Technician	2	50,000	100,000	1.35	135,000
Electrical Apprentice	2	25,000	50,000	1.35	67,500
TOTAL	50		4,030,000	1.35	5,440,500

Table 129 Labour Cost Estimate Breakdown.

The salaries for the Processing Manager, the Maintenance Superintendent, the Production Superintendent and the two metallurgists are based on expatriate personnel. All other salaries are based on local labour rates. There will be a total of 50 personnel comprising twenty-four operations personnel, sixteen maintenance personnel, and ten supervisory or management personnel.

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The working roster for the process plant will be two 12 hour shifts per day, seven days per week. Each of four operating shift crews will comprise a Shift Supervisor and four operators who will work a 4 days on, 4 days off roster. The crushing plant will be operated on day shift only by a day crew which will comprise six operators who will also be responsible for reagent mixing, gold room support activities, general site housekeeping and shift operator relief. There will be a dedicated gold room supervisor.

Maintenance personnel will work day shift only seven days per week with a call out system for night shift. The tradespersons will work a 4 days on, 4 days off roster which will provide two mechanical tradespersons, two boilermakers and one electrician each day.

Power

The power cost has been estimated at AU$1,791,859 for a full production year. Table 130 provides a breakdown of the power cost estimate. The unit cost for power of AU$0.09 per kWh is based on off-site generation by local power providers.

Cost Centre	Total Power Consumption (kWh)	Total Power Cost ($)
Crushing and Screening	1,395,889	125,630
Grinding and Classification	10,430,495	938,745
Regrind	1,341,027	120,692
Leaching and CCD	2,025,122	182,261
Tailings Thickening and Disposal	689,009	62,011
Gold and Silver Recovery	466,305	41,967
Reagent Mixing and Distribution	359,684	32,372
Water Services	985,290	88,676
Air Services	1,849,455	166,451
Workshop	39,858	3,587
Laboratory	113,880	10,249
Administration	213,525	19,217
TOTAL	19,909,540	1,791,859

Table 130 Power Cost Estimate Breakdown.

Reagents and Grinding Media

Table 131 provides a breakdown the inputs to the reagents and grinding media cost estimate. Reagents and grinding media prices are based on vendor quotations and estimated transport costs.

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Vendor supplied facilities have been assumed for oxygen and LPG delivery and storage. Facility changes for these have been included in the operating cost estimate.

It should be noted that cost of the cyanide, zinc and copper sulphate cost is over 65% of the total reagent and grinding media operating cost. Further test work to optimise of the consumption of these reagents has the potential to reduce to significantly reduce operating costs.

Item	Consumption (kg/t)	Unit Price ($/kg)
Interfroth 50	0.140	5.00
Potassium Amyl Xanthate	0.025	2.50
Copper Sulphate	1.252	3.00
Hydrated Lime	4.421	0.16
Sodium Cyanide	2.300	3.00
Sodium Hydroxide	0.138	0.50
Oxygen	0.100	1.00
Sodium Metabisulphite	3.589	1.10
Diatomaceous Earth	0.280 – 1.600	0.50
Zinc	0.816 – 3.166	7.00
Sulphuric Acid	0.456	0.25
LPG	0.110	0.60
Flocculant	0.140	4.50
Grinding Balls (80 mm)	2.100	1.70
Ceramic Media	0.010	12.00
TOTAL

Note: Oxygen consumption in m3/t and unit price in AU$/m3

LPG consumption in L/t and unit price in AU$/L

Zinc and diatomaceous earth consumption dependent on head grade.

Table 131 Reagents and Grinding Media Cost Estimate Input Breakdown

Maintenance

The maintenance (excluding labour) cost is estimated to be $1,163,736 per for a full production year. This has been based on scaling maintenance spares from the capital cost estimate and in house data for similar sized projects.

Linings

The cost of liners for the crushers and ball mill (excluding installation) is estimated to be $425,390 for a full production year. Table 132 provides a breakdown of the linings cost estimate.

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Linings	Unit Price (each)	Yearly Changes	Total Cost ($)
Jaw Crusher
-Fixed Jaw	4,448	7.9	35,187
-Movable Jaw	4,398	6.2	27,413
-Upper Cheek Plate	749	2.0	1,481
-Lower Cheek Plate	430	4.0	1,701
Subtotal			65,782
Product Screen
-Top Screen Deck	4,320	3.5	15,095
-Bottom Screen Deck	778	16.2	12,662
Subtotal			27,757
Cone Crushers
-Secondary Mantle and Bowl	4,442	9.3	41,355
-Tertiary Mantle and Bowl	4,442	5.5	24,222
Subtotal			65,577
Primary Mill
-Lining	120,000	2.0	240,000
Regrind Mill
-Lining	26,274	1	26,274
TOTAL			425,390

Table 132 Linings Cost Estimate Breakdown.

21.2.2

Mining

Open pit operating costs were developed from the recommended equipment requirements and the personnel requirements. The mine operating costs include all the parts, supplies, and labour costs associated with mine operation, and maintenance. Mining labour cost for supervision is included as part of the G&A. Table 133 through Table 135 summarize the total open pit operating costs. Total cost, unit cost per total tonne of material moved (including rehandling), and unit cost per tonne of material mined respectively.

In similar way the underground mine operating costs were estimated, shown in Table 136. The average estimated underground mining cost amounts to US$ 45.9 per tonne.

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Table 133 Open Pit Operating Cost (US$)

Table 134 Open Pit Operating Cost (US$/t moved)

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Table 135 Open Pit Operating Cost (US$/t ROM)

Table 136 Underground Mine Operating Cost

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21.2.3

General and Administrative (G&A) Costs

General and administrative (G&A) costs for the Cerro Moro project have been calculated using May, 2011 US dollars. The G&A costs are inclusive of the following:

●

On-site and off-site labour costs relating to General Management (Mine General Manager, Section Manager for excluding Process Plant Manager), HSE, community relations, finance and administration, and legal. Note that GRES has already included the cost of the Process Plant Manager in the OPEX cost for processing.

●

Cost of the Cerro Moro camp (including catering / meals and cleaning services)

●

Off-site offices in Puerto Deseado and Buenos Aires

●

Miscellaneous costs, including: camp and general site power, safety clothing and personal protection equipment, environmental monitoring and auditing, site security, first aid and medical services.

●

Management (including on site admin) $US2.5M

●

Cerro Moro Camp $US 1.5M

●

Off site offices $0.6M

Miscellaneous

–

Safety clothing $0.15M

–

Camp and General Site Power $0.4M

–

Insurance $0.8M

–

Site Security $0.25M

–

Environmental Monitoring: $0.15M

–

Recruitment and Training: $0.15M

–

Road Maintenance $0.25M

–

First Aid and Medical $0.24M

–

Auditing and Bank Fees: $0.08M

●

Total $US 7.2M

At a production rate of 1000 tpd, the average G&A costs equates to $20/t

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22.	ECONOMIC ANALYSIS

22.1	Taxes and Royalties

The tax and government incentives for the project development were not included in the evaluation. The negotiations have not been finalized with the government and its agencies so that these values can be included in the pre-feasibility study.

Taxation and royalties criteria were applied as follows:

Income tax: A regular income tax of 35% was considered.

Export tax: 5% of total revenues were applied.

New Provincial Royalty: 1% of total revenues was applied.

CVSA NSR:2% of total revenue less refining costs was applied. This corresponds to the agreement with Cerro Vanguardia SA (CVSA).

Boca Mina: 3% of operating profit was applied, which corresponds to operating cash flow less export tax and CVSA NSR.

22.2

Base Case

A simple cash flow financial model was created by NCL, based on its experience, utilizing the mine production schedule, associated gold/silver grades, gold/silver recoveries estimated from the preliminary metallurgical test program, and capital and operating costs as set out in this document.

Base case prices at US$1320 per ounce gold and US$26 per ounce silver were considered for all the life of mine, provided by Extorre.

The study entails a 1000 tonnes per day to flash flotation and gravity concentration of coarse material and a conventional cyanidation of the tails, with further treatment of the product by Merrill Crowe. Total metal production amounts to 495,000 ounces of gold and 26,687,000 million oz of silver over 8.2 years of production.

A constant tax rate of 35% was applied; and four types of royalties were considered, averaging 10% of total revenue.

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NCL has considered the cash flow on a 100% equity basis, i.e. no account has been taken of financing arrangements and associated costs.

Economical parameters used for the evaluation are shown in Table 137.

Item	Unit	Value
Average Mining Costs	Open Pit US$/tonne mined Underground US$/tonne mined	2.15 45.9
Processing Cost	US$/ore tonne	67.00
G&A	US$/ore tonne	20.00
Reclamation fee	US$ million per year	1.00
Base Case Prices	Gold US$/oz SilverUS$/oz	1320.00 26.00
Transport, Freight, Insurance, Refining	Gold US$/oz Silver US$/oz	10.00 0.50
Metallurgical Recovery	Gold (%) Silver (%)	95% 88%
Discount Rate	%	5.0
Income Tax	%	35
Royalties	Export tax - % of total revenue New Provincial Royalty - % of total revenue CVSA NSR - % of total revenues less refining costs Boca mina - % of operating profit	5.0 1.0 2.0 3.0
Depreciation	In three years, 33.3% per year

Table 137 Economical Model General Parameters

Base Case net present value (NPV) after-tax and at a 5% discount rate, amounts to US$ 274 million, generating an IRR of 58% after-tax.

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Table 138 Cash Flow Economical Model for US$ 1320 per ounce

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22.3	Sensitivity Analysis

The project’s sensitivity to changes in revenue, operating costs, capital costs and discount rate was tested with the following observations (Table 139 through Table 141):

●

When the gold price was increased to a constant US$ 1,584 per ounce gold (+20%), NPV after tax at a 5% discount rate rises by US$ 62.6 million to US$ 337 million and the IRR rises from 58% to 70%.

●

In none on the analyzed cases the project NPV became negative. The lowest obtained value was US$ 158 million for the combination of US$ 1,056 per ounce gold and 10% discount rate.

●

The lowest obtained IRR was 35% for the combination of US$ 1,056 per ounce gold and 20% increase on the capital cost.

●

All figures given in the following tables are After Tax

NPV (M$)
Gold Price & Discount Rate	Gold Price (US$/oz)
	1,056	1,320	1,584
Discount Rate	0.0%	$ 280.7 M	$ 356.4 M	$ 432.2 M
	5.0%	$ 211.9 M	$ 274.4 M	$ 337.0 M
	10.0%	$ 159.7 M	$ 212.3 M	$ 264.9 M
IRR (%)	47%	58%	70%

Table 139 Sensitivity – Gold Price and Discount Rate

NPV (M$)
Gold Price & Operating Cost	Gold Price (US$/oz)
	1,056	1,320	1,584
Opertaing Cost	-20%	$ 259.6 M	$ 322.9 M	$ 386.1 M
	Base Case	$ 211.9 M	$ 274.4 M	$ 337.0 M
	+20%	$ 186.5 M	$ 249.9 M	$ 313.2 M

Table 140 Sensitivity – Gold Price andOperating Cost

NPV (M$)
Gold Price & Capital Cost	Gold Price (US$/oz)
	1,056	1,320	1,584
Capital Cost	-20%	$ 298.2 M	$ 366.0 M	$ 430.5 M
	Base Case	$ 211.9 M	$ 274.4 M	$ 337.0 M
	+20%	$ 178.6 M	$ 241.1 M	$ 303.7 M

Table 141 Sensitivity – Gold Price andCapital Cost

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23.	ADJACENT PROPERTIES

On March 3, 2009, Exeter and Fomento Minera de Santa Cruz Sociedad del Estado 1 (“Fomicruz”), the provincial mining company, signed a definitive agreement over ten Fomicruz concessions, covering 69,100 ha (691.0 km2), adjacent to Cerro Moro. Under the terms of the agreement, Exeter (now Extorre) has the option to acquire an 80% interest in the Fomicruz properties by spending US$10 Million on exploration over a number of years. As part of this agreement, Fomicruz will acquire a 5% participation interest in Exeter´s Cerro Moro project on the granting of the required exploitation concessions and permits to commence a mining operation. Extorre will manage and fund future exploration and development on both the Cerro Moro and Fomicruz concessions. Fomicruz will repay an agreed amount of its attributable exploration and development costs from the 50% of its share of the net revenue from future mining operations.

As indicated in Figure 176 the Fomicruz properties border the south-western, southern and south-eastern portions of the Cerro Moro property.

Figure 176 Fomicruz Properties Location

Fomicruz is a provincial state entity and is not required to publicly release any results from their properties and, as far as the author is aware, they have not done so. Previous work on the property by Fomicruz is limited to preliminary minor geological mapping and rock chip sampling of selected veins.

Preliminary reconnaissance exploration commenced during 2009 on the Fomicruz properties, working from the north to the south. The exploration methodology has been threefold:

1Fomicruz is a company owned by the Government of Santa Cruz Province, Argentina

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Regional stream sediment sampling, involving the collection of;

An approximate 3 kilogram BLEG2 sample, -2mm to 5 mm size fraction.

ii.

An approximate 500 g of 2 mm to 4 mm size fraction.

iii.

Approximately 500 g of -2 mm size fraction

During this program the geology, structure and any alteration is mapped at a scale of 1:25,000, along with rock chip sampling of any potentially or obviously mineralized structures.

Ground magnetic data collection - Exeter purchased two GEM 19 Overhauser Magnetometers for this task; a base station and a ‘walking’ magnetometer. The walking magnetometer is fitted with a GPS and was set to take readings every 2 seconds. As detailed in Williams and Perkins (2009) the majority of the Cerro Moro project was covered by either 40 metre or 80 metre spaced lines. This surveying is now being extended on to the Fomicruz tenements. The data will be used to assist with geological mapping, the identification of major structures (particularly those under cover), with the principal aim of identifying potential mineralised structures.

The current third phase is follow-up of any anomalous sampling results (stream sediment sampling or rock chip results) with more detailed mapping and the collection of additional rock chip samples. Follow-up drill testing of defined target areas on a priority basis.

To May 31, 2011, a total of 61 drill holes for 12,383,50 m had been completed on the Fomicruz Properties in four principal targets – Escondida Fomicruz, Bella Vista, Veta Olvidada, and Union Domes (Table 142). Drilling in 2011 was restricted to the Union Domes target area and select drilling on the Escondida Fomicruz area.

Drilling Program	Prospect	Holes	RC (m)	DDH (m)	Total (m)	%
MINERALIZATION	Escondida Fomicruz	42	1,212.00	8,428.9	9,640.90	78%
	Bella Vista	9	430.00	883.40	1,313.40	11%
	Veta Olvidada	1	0.00	190.00	190.00	1%
	Union Domes	9	0.00	1,239.20	1,239.20	10%
Total		61	1,642.00	10,741.50	12,383.50	100%

Table 142 Drilling Completed in Fomicruz Properties

2BLEG = Bulk Leach Extractable Gold – utilises cyanide leach procedures for the extraction of gold, which are often used in grassroots exploration where cyanide extractable gold from a very large sample can sometimes detect small gold anomalies that otherwise would go undetected.

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23.1	Escondida-Formicruz Prospect

The Cerro Moro Escondida mineralised structure trends northwest into the Fomicruz property and is clearly evident in the ground magnetic data to trend for an additional 2 kilometres (refer to Figure 178).

Figure 177 Drilling Completed at Escondida (Fomicruz Extension), Bella Vista, and Veta Olvidada with Ground Magnetic Data (RTP)

The majority of the Escondida-Fomicruz structure is under cover and as such does not outcrop. Initial exploration (scout) drilling along the Escondida-Fomicruz structure was undertaken with RC percussion holes on sections spaced 160 m apart. The main objective of this drilling was to locate the trace of the Escondida-Fomicruz structure under the Tertiary marine sediment cover. Evaluation of the geochemical data obtained from this initial drill campaign led to the design of the follow-up drill program, in which nine diamond drill holes were completed on five sections. The target depth for these nine holes was 150 m below surface; all holes encountered hydrothermal alteration (silicification and illite-smectite) but returned low precious metal values. Follow-up work is being undertaken to determine whether the drilling completed to date was targeted too close to surface (i.e. potential for economic mineralization may still exist below 150 m vertically below surface).

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Recent drilling on the Escondida-Fomicruz structure has been focused on diamond drilling at the southerneastern end of the structure (essentially western extensions to the Escondida Far West orebody). Significant results have been reported from drilling in this area (Table 143).

The locations of drill holes in the Escondida-Fomicruz structure are indciated in Figure 178.

Figure 178 Escondida-Fomicruz Prospect – Drill Hole Locations

23.2	Bella Vista Prospect

The Bella Vista prospect is a 4 kilometre long, predominantly east-west striking, structure, consisting of fault breccia, chalcedonic cement, and local black silica with fine grained sulphides (Figure 177).

Surface expressions of the main structure are anomalous in gold (up to 2.6 g/t), silver (up to 30 g/t), molybdenum (up to 350 ppm), arsenic, mercury, lead and zinc. It also has a moderate to subtle signature in the ground magnetic data and the one trial MT survey line. Located approximately 300 metres to the south of this prospect is another structure represented by poor outcropping quartz/breccia outcrop, and interpreted to trend parallel to Bella Vista. Reconnaissance rock chip sampling of this area encountered gold grades up to 7.6 Au g/t.

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As a first test of the Bella Vista prospect, three RC percussion drill holes and six diamond drill holes were drilled during September 2009 and April 2010, respectively (Table 143). No significant results were returned from the initial RC percussion drilling. The April 2010 diamond drilling was focused on those portions of the Bella Vista vein system with significant surface geochemical anomalies, with drilling designed to intersect gold-silver mineralization at approximately 100 m depth below surface. This drilling succeeded in intersecting the Bella Vista structure and quartz veining; however, no potentially economic ore grades were encountered. Significant results returned by this drilling are given in Table 143.

23.3	Veta Olvidada Prospect

The Veta Olvidada is a narrow, northwest-trending vein / structure located approximately 1 km south of the Bella Vista vein. On surface, the Veta Olvidada is characterized by chalcedonic quartz with very low sulfide content and low precious metal contents (but highly anomalous trace element geochemistry). Alteration on the host rocks is domintantly illite and smectite. Only one diamond drill hole was drilled in to the central portion of Veta Olvidada (Figure 177), with only low level gold and silver values being reported (Table 143).

23.4	Union Domes Prospect

Preliminary geological mapping and a first pass reconnaissance drill program have indicated that gold(-silver) mineralisation at the Union Domes prospect, located approximately 13 km south of the Escondida structure at Cerro Moro, is associated with a shallow northwards-dipping tectonic breccia. Host rocks to the mineralisation are shallowly dipping (to the south – southeast) felsic tuffs of the Matilde Formation and minor carbonaceous epiclastics. The La Matilde Formation has been locally intruded by a fine grained rhyolitic sill. A total of 9 drill holes for 1,239.20 m were completed in late March 2011 at Union Domes; as at May 31, 2011, assay results for the drill program were still in the process of being received and reviewed.

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Table 143 Significant Drill Intercepts from Fomicruz Properties

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24.	OTHER RELEVANT DATA

24.1	Hydrogeology

The hydrography of the Deseado River, Fénix River, and Pinturas River basins was described, and the lakeside endorheic basins were characterized. In the area of direct and indirect impact of the Project, the sub-basins were delimited and geometric characteristics were determined. Design Storms for a recurrence of 5, 10, 25, 50, and 100 years event were estimated.

A Surface and Groundwater Quality Study was performed, and included the analysis of physical-chemical and microbiological parametres from 17 sampling stations (8 stations correspond to groundwater and 9 stations to surface water samples). All the decision-making process was carried out pursuant to quality control standards. Monitoring activities in the chosen stations were performed seasonally due to the site’s environmental conditions, which lacks permanent surface water bodies.

Below are the main conclusions regarding surface water monitoring.

●

In the Project area there is only one surface drainage network which channels water towards endorheic saline depressions. These collect surface runoff in winter and are an evaporation area during spring-summer, naturally concentrating salts, dissolved solids, and neutral pH. Due to the strong evaporation processes and little rainfall, there is year-round salt concentration. In the monitored endorheic depressions there are relatively high values of fluorides, boron, manganese, and vanadium. In the Deseado River sampling point (upstream) there is water in certain times of the year; there are mainly sodium bicarbonate waters with a slightly alkaline pH, and relatively high fluorine, aluminum, nickel, and manganese contents;

●

In the Deseado River sampling point (downstream) there are high contents of total dissolved solids in the samples taken during the February and April 2009 campaigns, thus evidencing sea-water inflows into mainland. These are sodium chloride waters with a slightly alkaline pH. The results obtained during the winter were analyzed and there is an increase in dissolved salts and suspended solids due to their entering the firth; there are also relatively high fluorides, aluminum, nickel, boron, and vanadium values.

Below are the main conclusions regarding groundwater monitoring.

●

Water extracted from mills to provide for the hacienda and for local people's consumption (6 sampling points) was taken from depths of 5 to 20 cm, and appears to be water stored in alluvial sands and silts that fill the depressed areas of basins and valleys, creating surface aquifers. This water has a relatively low saline content, with an electrical conductivity from 1,500 to 5,000 µS/cm;

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●

The deepest waters (samples from mining holes) (2 sampling points in Escondida and Déborah) capture the water from greater depths (over 40 m) and correspond to rock aquifers with secondary porosity and high salinity, around 7,000 µS/cm for the Déborah well and 15,000 µS/cm for the Escondida well;

●

As for the anionic composition of mill waters, these waters also contain chloride and bicarbonate. Cationic composition is mainly sodic, and calcic to a lesser degree. In Escondida, the water contains sodium chloride with low bicarbonate and calcium content and a neutral to alkaline pH;

●

A degradation of water quality exists in the water from the northwestern--western/southwestern mills: water extracted from the mills upstream from the Project has a relatively better quality than that taken from downstream. Some samples have arsenic (except the Escondida well) and fluoride contents above the legal levels;

●

The presence of total coliform bacteria was common in all the samples analyzed, exceeding the values set forth by the Argentine Food Code (except in Escondida). There are total aerobic bacteria in all stations. The reason behind the microbiological pollution found in the water from monitored mills is the latter’s poor facilities: they are constantly visited by the hacienda for water consumption and it is easy to see around the wells large amounts of dung;

●

The hydrogeological study performed in the Project area was aimed at preparing a hydrologic mass balance to determine groundwater supply for the Project’s activities. Results showed that the modeled area has a volume of 615 cubic metres per day (design flow rate for Hydrogeological Model), which ensures that there will be enough water supply for said activities;

●

New holes were also drilled to determine the basins’ parameters. Piezometers were defined, based on an analysis of the equipotential surfaces for the Escondida – Loma Escondida – Casius – Lechuzo, Silvia – Carla – Carlita, Déborah, Gabriela – Esperanza areas. The hydraulic parameters for the execution of a conceptual hydrogeological model were defined;

●

The Advanced Hydrogeological Research started in October, 2009, and finished in July, 2010. Twenty wells were drilled and/or redrilled, and 18 pumping tests were performed. Moreover, 17 piezometers were sited for constant head tests.

24.2	Mine Geotechnical

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AKL recommended slopes angle can be summarized as follows:

●

Escondida West: South-West Wall 68 °

●

North Eastern Wall 65 °

●

Escondida Central: South-West Wall 68 °

●

North Eastern Wall 65 °

●

Loma Escondida: South-West Wall 65 °

●

North Eastern Wall 68 °

●

Esperanza: 68° overall

●

Gabriela: 68° overall

●

Deborah: 68° overall

Its AKL opinion the stability analysis supports a pre-feasibiity study.

According to the limited information available in some parts of Escondida, where will be a important part of the developed of the slope of each pit, recommendation interamp angle is lower than in other sectors. Furthermore, the recommended maximum height should be within the range of 60m to 70m maximum height for all prospects with 5m bench in the case of single bench, which can turn left 10m to achieve greater containment berm materials, bench face angles for this design will remain between 84 to 85

ii.

With regard to the development of new designs, modification of current changes in certain parameters such as height, etc. Reassessment should be performed geotechnical the updated geotechnical information and the proposed designs.

iii.

In a few cases of prospects considered here, such as the Esperanza pit design, an overload is detected of unconsolidated gravel with a thickness greater than a bench, for such case, it is recommended that a catch bench at the gravel contact -mass of about 10 m, with the objective of containing the instabilities in such poor quality geotechnical unit.

As a result of this study are some important recommendations for the design of the different slopes of the Cerro Moro Project prospects.

In order to improve the quality of information concerning the properties of the geotechnical units and notwithstanding that there has been progress in this regard, it is necessary to conduct further geotechnical testing.

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Must be done to build a model of major structures, which must consider and include large geological structures, such as regional or district scale structures also mine.

Hydro geological studies should be conducted more detailed, taking into account that the period of development of the slope does not provide sufficient time for the groundwater flow reaches a steady state, so that should be considered a transient state of these

According to information and characterization of the rock mass that exists today, we recommend the drilling of some additional geotechnical drilling covering some areas with little information in the some prospect involved here

As a result of this work can be noted:

Assessment was conducted geomechanical stability of underground mining method called "Cut and Fill by Ascending Bench (Bench and Fill), emphasizing in particular as regards the stability of cave and crown pillar.

The results of the stability analysis of the caves, we suggest that, because their dimensions considered 25m high and 4m wide, and also that these will be filled as they come to be mined, would not have a maximum length restrictions, Therefore, for this stage of engineering is recommended maximum length of 60m without fortification associated.

For the stage of engineering pre feasibility of this work, we consider a FS> 3.8, this mainly to calculate the thickness of the crown pillar, which can be designed for contact with the surface of the original topography or to the interaction the walls or bottom of a open pit mining. If the length of the pillar, Lp <60m, these pillar will be of 7m thick, if the length of the pillar, 60m < Lp <80m, we recommend a thickness of 10m. It should complement these studies with other types of analysis in order to validate these parameters, in the following engineering stages of Cerro Moro project.

For the next stage of engineering Cerro Moro, in underground mining is recommended:

While geomechanical analyses made are valid, they are based on empirical analysis, so they must be improved to have better information geomechanics.

"Anomalies" in the geological condition, such as the presence of faults, shear zones, dykes or inclusions of waste rock, creating a slot or visor inside the cave and poor support installation can all, individually and collectively, bring to an inappropriate result.

Practical observations suggest that the main area of uncertainty when using the method is regarding to the density of structures in the rock mass. The number of discontinuities and other structures per unit volume of rock is highly variable

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The design and recommendations derived from the use of this method should be considered a first step in the design process. It should make all appropriate adjustments based on conditions observed in the cave of interest.

Analysis should be performed by the failure factor, which successfully integrated into the design of caves the presence of major faults near the exposed surfaces.

The effects due to blasting activities are usually ignored. Poor quality blasting can create new fractures and relax rock blocks, which consequently reduces the quality geotechnical rock mass.

The faults have the most influence on the stability of a cave where the failure is sub parallel to the wall of the cave (20 ° to 30 °), the minimum effect is generated when faults are approximately perpendicular to the walls and or roof of the cave, therefore must be characterized if these exist.

In general, when studying the stability of a crown pillar, can identify six major failure modes, each of which can be given alone or a combination of more than one of them. In general, the dominant mechanism depends on the geotechnical characteristics of each sector under study. These mechanisms are: Structurally Controlled Fault, Fault rock mass stress-induced failure, Failure type fireplace, beam or plate failure rate and failure rate for blocks voussoir.

Define acceptability criteria

According to the results of stability analysis for the design of waste dump in Chapter 9 of this report, we can indicate the following:

According to the designs considered, the analysis identified 06 sections to cover all surfaces of the proposed designs for which are generally determined by a design section, this being the most critical section, either by design of the dump itself, topography of the base will be deposited, or quality of the soil foundation.

The maximum height of the proposed dump designs of the project, ranging from 40m at most, considering the overall heights, the heights interramp, in these designs are variables.

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To perform the stability analysis, we analyzed 06 vertical sections above, considering that these include the natural topography where dumps will be placed and also presented the highest peaks of the slopes of the previously proposed designs dumps.

Based on field data collected, review of technical literature and the experience of other mining projects with similar characteristics, we estimated the geotechnical properties of foundation soil and basement rock, which is deployed the waste dumps that interest in this assessment.

Considering the nonlinear behavior of the material will be form the body of the dumps, and although their overall heights near the height limit for geotechnical properties were defined in terms of the magnitude of the confinement stress, these were considered as allows a better representation of real behavior of granular materials free.

The stability analysis was performed by limit equilibrium, determining the safety factor, FS, using the GLE method, which was preferred over other methods because it considers the best way the effect of the nonlinearity of the envelope of failure for this type of material.

Also evaluated for each section, the probability of failure, PF (eg see Duncan (2000)), considering the variations coefficients of 10% for the friction angle, and 40% for cohesion.

The results for each of the sections evaluated are presented in Appendix E of this report.

10)

The design analyzed the dumps does not consider the presence of water below the surface of these, nor runoff that eventually can penetrate the base and body of these.

11)

According to the above and as a result of geotechnical evaluations, the designs dumps of the prospects: Escondida, Loma Escondida, Gabriela, Esperanza and Deborah, meet the acceptability criteria defined for this report, considering static and pseudo cases, as outlined in this report.

12)

The stability of the dump is valid under the considerations of a coarse granular material in a homogeneous mixture of 4:1 with paste that coming of stage of mineral processing and properties similar to those presented in this report

As a result of this study is presents the following recommendations for the design of the waste dumps slopes in the Cerro Moro prospects.

While the condition of stability for the proposed design is acceptable to a maximum height of 40m, and overall height, it is advisable to stay decoupling, as the results obtained give us close to the minimum values according to acceptability criteria considered this report.

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Global stability of the proposed design improves considering leaving terraces free of at least 10m wide, it is advisable to reduce the width, but is likely to do so, until you get a better characterization of the dump body materials.

According to the above, for the pre feasibility in Cerro Moro Project, the design dumps parameters recommended is finally:

●

Maximum height 40m.

●

Interramp slope angle 32°

●

Overall slope angle 25°

●

Heights first layer 10m

●

Heights top two layers 15m

●

Wide terraces 10m

For the following stages of the project is recommended to conduct a campaign to characterise the materials involved in the dumps design.

We recommend testing homogeneity of the dump body materials, as there is still the procedure of mixing of granular materials and paste that make up the body of these.

24.3	Plant Geotechnical

The process plant area topsoil and surface has low growing shrubs and contains organic matter, the topsoil average depth is 0.40m. At the bottom of the valley to the east of the proposed ROM pad location where the terrain opens into a plain, below the surface layer of silty sand there is a very wet clay layer starting at 0.80m extending to 1.80m below the surface. Below the clay layer there is weathered sedimentary rock sandstone that breaks very easily reaching a depth of 3.20m.

The central sector of the valley below the layer of topsoil, there is a layer of sedimentary rock very hard limiting excavation to 1.40m. In the upper valley below the top soil there is a silty sand layer that extends from 0.40 to 0.70m deep below which there is weathered sedimentary rock sandstone of medium hardness that was able to be removed by backhoe up to 1.70m.

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24.4	Operations Project Organisation

A total of 370 will be employed for the project, distributed as follows:

●	General Management and Administration	80
●	Process Plant	58
●	Mining Department	232
●	Total	370

The General Manager will be responsible for the entire operation. The area managers, Environmental Coordinator and safety engineer will report directly to him.

The process group will be led by the Process Manager who will report to the General Manager. The Process Manager will require being an engineer with 10 – 15 years of process experience including a minimum of five years of experience senior supervisory with gold processing plants. This person will need to be hired about six months before the start-up of the process plant and be involved with the hiring of his team plus the development of the training program.

The Administration Manger will be responsible for accounting and cost control, human resources, site security, and procurement. This person will need to have 5 – 10 years of experience with mining operations. He will need to be hired about six months before start-up.

The Environmental Coordinator will report to the General Manager. This person will need to be a university graduate and have five years of experience with environmental control, and legal and governmental procedures. This person needs to be hired one month before the start of the mining operations.

The Safety Coordinator will report to the General Manager. This person will be a safety engineer with 5 – 10 years of experience with construction and mining operations and will be responsible for implementing the zero accident safety policies plus training.

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Contractors will provide services for cleaning the offices and dry, security, waste removal (organic, hazardous and non hazardous), exploration drilling, and transportation of people and gold.

24.5	Project Implementation

The Cerro Moro Project will be implemented under two main banners:

●

Client managed open pit and underground mining design, mining equipment and infrastructure procurement and mining implementation;

●

Engineering Procurement and Construction Management (EPCM) contract for the design and construction management of the process plant and infrastructure;

The complete project will be managed by a Clients Representative who will report to Extorre.

24.5.1

Mine Management Team

The mine design and implementation will be managed by the Mine Manager as described in section 24.4.

24.5.2

EPCM Project Management Team

The implementation strategy for the Cerro Moro process plant requires a robust EPCM team for the management of design, procurement and construction. The team will be formed from Extorre personnel, EPCM contractor personnel and Argentinean consultant engineers;

EPCM Contractors Office

EPCM design team for the generation of project drawings, equipment specifications, tenders and construction work packages for award by Extorre, including all project management and cost control functions.

Puerto Deseado, Argentina

EPCM Contractor / Extorre Project office for local procurement and logistics management, including local cost control and labour recruitment.

Cerro Moro, Argentina

EPCM Contractor / Extorre construction management office for all site construction works implementation supervision and controls.

24.5.3

Development Objective

The objective of the project implementation plan is to provide Extorre Gold Mines Limited (Extorre) with the most cost effective and timely approach to project development. The most cost effective outcome will be achieved by a project that is delivered on schedule and ramps up to full production capacity in as short a time as possible.

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These outcomes will enable Extorre to effectively bring mining, plant operating and administration resources into the project at appropriate times to minimise the phase of the project when operating costs exceed revenue.

24.5.4

EPCM Services

The EPCM engineer will provide the following services associated with the development of the Cerro Moro Gold-Silver Project:

Infrastructure

●

Design and construction of the plant access roads;

●

Management of the design and construction of the site HV power supply and distribution system;

●

Design and installation of project buildings including the accommodation camp and administration facilities;

●

Design and installation of water supply requirements;

●

Design and installation of fuel storage and handling facilities;

Process Plant

●

Finalise geotechnical engineering for the plant site;

●

Process engineering;

●

Surveying services during construction;

●

Design engineering and drafting of earthworks, civil works, structural steel, mechanical and electrical installations;

●

Project management services including cost control, scheduling, reporting and claims processing;

●

Procurement of materials, equipment and fabricated items including tendering, purchasing, expediting and contract preparation and administration;

●

Logistics coordination;

●

Construction management including site management, and control and inspection of all construction activities;

●

Commissioning including pre-commissioning and testing, dry commissioning, wet commissioning and operational assistance until handover;

●

Regulatory compliance monitoring and reporting of project activities;

●

Overall site management until completion of commissioning.

The EPCM engineer will execute the engineering, design and procurement from its home office and will maintain direct construction management from a site office at the plant site. The EPCM engineer will also be responsible for ensuring all engineering design is verified and signed off by a certified engineer in Argentina to ensure all approvals and statutory requirements are met.

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Extorre will retain responsibility for government liaison and permitting, all operating and related activities, community relations, power supply infrastructure and provider contact award and administration, asset insurances, land purchases and legal services.

24.5.5

Equipment and Materials Supplies

Generally, all equipment and materials will be competitively tendered. Award of supply packages will be made on the basis of technical compliance, price and delivery period. Supply packages will be tailored to suit individual supplier’s capacity to optimise schedule, cost and risk.

The processing and mechanical equipment required for the project will manufactured in a number of countries and where possible consolidated for sea freight utilising a combination of Buenos Aeries and Puerto Deseado for receipt of goods into Argentina.

Some specialised equipment will be purchased as design, fabricate and site installation packages. Examples include the Merrill Crowe plant and some infrastructure buildings.

Generally equipment and materials will be installed by local Argentinean construction crews not affiliated with the equipment or materials supplier, under direct supervision of the Engineers personnel.

24.5.6

Construction Packages

A construction workforce would be established by utilising Argentinean contractors to provide direct supervision and labour for the majority of packages who would be directed by the EPCM engineers supervisory staff and will be utilised to complete works associated with mechanical installations, structural and plate work installation and erection, electrical and instrumentation installation and piping installations.

Specialist subcontractors will be utilised to complete bulk earthworks associated with the plant and for concrete batching and placement.

24.5.7

Project Duration

The project implementation schedule is shown as Figure 179 at the end of this section, this schedule is based on using all new equipment. The project schedule indicates a project duration of seventy eight (78) weeks from unhindered commencement of design activities in the key process plant area to practical completion (commencement of ore commissioning). Key milestone dates required to achieve this schedule are:

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§ Commencement of process facilities design works	Week 0
§ Commencement of major equipment procurement	Week 0
§ Site access available	Week 3
§ Finalisation of PID and flowsheets	Week 3
§ Finalisation of plant layout	Week 5
§ Major equipment orders secured	Week 9
§ Commencement of plant site bulk earthworks and access road	Week13
§ Commencement of offsite fabrication works	Week 16
§ Mobilisation of civil workforce	Week 18
§ Commencement of structural steel erection	Week 41
§ Commencement of mechanical installation	Week 37
§ Commencement of piping installation	Week 55
§ Commencement of electrical installations	Week 55
§ Commencement of dry commissioning	Week 71
§ Commencement of ore commissioning	Week 78

It is anticipated that a period of 8 weeks from the commencement of ore commissioning will enable the plant to ramp up to full production rates. During this period operator training, rectification of any defects and optimisation of operation will take place.

24.5.8

Critical Path

The critical path for this schedule is defined by the time taken to complete the following activities:

●

Ball mill and flotation cell specification, ordering, manufacture and delivery.

●

Installation of the ball mill and associated structures.

●

General arrangement and mechanical drawings sufficient to commence structural steel design and drafting.

●

Structural steel design and drafting.

●

Structural steel erection sufficient to establish safe access for mechanical, piping and electrical work groups.

●

Subsequent mechanical, electrical and piping installations.

●

Electrical and plant control system installations.

●

Dry and wet commissioning.

24.5.9

Engineering Design

The schedule assumes that much of the work required for project and scope definition is already complete at the commencement of detailed engineering work. As such, after an initial minimum five week period to finalise and confirm plant layouts and project definition document, design works are shown progressing in discipline order from mechanical and equipment layouts to structural steel arrangement, concrete and piping drawings.

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Electrical and plant control system (PCS) design is indicated starting later in the design phase reflecting both the later requirement for information to the site installations crews and the fact that most of the mechanical design (such as equipment and drive selection) needs to be completed before electrical design work can be efficiently progressed.

24.5.10

Equipment Procurement

Generally (with the exception of the critical path ball mill) float of at least two weeks exists between the scheduled delivery of the major equipment and the date the equipment is required on site for incorporation into the works.

Delivery periods (to site) for major equipment are as follows:

●	Ball mill	55 weeks
●	Jaw and cone crushers	45 weeks
●	Vibrating screens	35 weeks
●	Gravity Concentrator	34 weeks
●	Flotation cell	34 weeks
●	Thickeners	45 weeks
●	Transformers	42 weeks
●	MCCs and variable speed drives	38 weeks

24.5.11

Site Construction Activity

Site construction activities are shown progressing generally on a discipline by discipline basis from bulk earthworks, concrete installations, structural steel installations, equipment installations and piping and electrical installations.

The mobilisation of site earthworks contractor, site supervision personnel and the commencement of civil works will be scheduled to allow these tasks to be substantially complete prior to the mobilisation of the critical path structural steelwork erection phase of the construction works. Sufficient time is allowed at the start of the project to confirm specialised site requirements and to finalise HSE requirements and inductions prior to mobilisation.

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Figure 179 Operations Implementation Schedule

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25.	INTERPRETATION AND CONCLUSIONS

25.1	Mineral Resource

On April 19, 2010 Extorre Gold Mines Limited (Extorre) announced the second NI 43-101 compliant mineral resource estimate of related to the Cerro Moro Project. Of particular importance to the resource statement is the high grade Escondida prospect. The April 2010 resource statement shows that all Indicated mineral resources are attributable to the Escondida prospect whilst other prospects such as Loma Escondida, Gabriela, Deborah and Esperanza remain as Inferred. The aim of Extorre’s resource definition and exploration strategy since April 2010 has been to undertake infill drilling on key Inferred resources, namely Loma Escondida and Gabriela to allow the successful conversion of a significant proportion of these prospects to the Indicated category for the purposes of this PEA. In addition, exploration drilling has been targeted at new prospects with the aim of indentifying extensions to the overall resource base.

It is Cube’s opinion that the resource definition and exploration strategy since April 2010 has been highly successful in meeting these stated aims as evidenced by the conversion of significant proportions of the Loma Escondida and Gabriela resources in April 2011 to Indicated and the recent discovery of the high grade Zoe prospect. The recently discovered Zoe prospect has the potential to make a material difference to the resource base at Cerro Moro and infill drilling is currently underway to provide sufficient data for a mineral resource estimate to be carried out in the near future.

By May 31, 2011, a total of 1,255 drill holes for 131,365 m had been completed on both exploration and infill drilling at Cerro Moro. Of this total, approximately 77% of the holes have been drilled at Escondida (50.38%), Gabriela (14.58%), Zoe (6.08%), and Esperanza (5.38%), reflecting the closer spaced drilling at the Escondida and Gabriela ore zones.

All logging, sampling and data QAQC procedures have been carried out to a high industry standard and record keeping and database management is excellent. Cube believes that the current database provides an accurate and robust representation of the Cerro Moro Gold-Silver Project and is appropriate for ongoing resource evaluation.

Extorre’s ongoing resource definition strategy is aimed at

the definition of maiden Inferred Category mineral resources in prospects such as Zoe, Carla, Martina, and Esperanza NW-Nini;

ii)

the conversion of existing Inferred Category mineral resources in prospects such as Deborah and Esperanza to Indicated Category; and

iii)

iii) the identification of potential new resources on other new and existing prospects.

Cube believes that Extorre’s ongoing resource definition and exploration strategy is well considered, appropriately funded and has a high chance of successfully fulfilling its stated aims.

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25.2	Mineral Reserve Estimates and Mining

Mining studies have been completed using the resource estimate as of April ,2011, including:

●

Pit optimization using Whittle FX to determine the ultimate pits limits for each sector.

●

Eleven pits have been proposed, for the different identified veins. Bench and overall pit slope design was based on recommendations by the geotechnical consultants (AKL).

●

Several waste storage areas close to the pits were designed according to geotechnical recommendations.

●

Pit contained resources inventory amounts to 1.7 million tonnes of ore at grades of 6.65 g Au/t and 341.9 g Ag/t. Inferred mineral resources have been included that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as mineral reserves. No Mineral Reserves have been estimated.

●

Underground mine design for cut and fill system, applying dilution and ore losses, for two of the identified sectors. Mineable underground resources inventory amounts to 1.0 million tonnes of ore at grades of 4.79 g Au/t and 352.2 g Ag/t. Inferred mineral resources have been included that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as mineral reserves. No Mineral Reserves have been estimated.

●

Development of the mine production schedule and plant feed schedule, based on the pit contained resources and underground mineable resources inventory, for a processing rate of 1000 tonnes per day and a life of mine of 8.2 years.

●

Identification of suitable mining equipment types and calculation of mining fleet requirements for both, for open pit and underground mining.

●

Estimation of mine capital and operating costs. Capital costs consider owner-operator mining, and include provision for preproduction mining. Replacement and additional equipment purchase costs have been included over the life of mine.

●

The operating cost estimate is dominated by equipment operating costs (fuel, tires, labour, and maintenance). Blasting, administration/management and services were also included. It was assumed owner maintenance for the mining fleet.

25.3	Metallurgical Processing

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25.4	Preliminary Economic Assessment Results

Base Case net present value (NPV) after-tax and at a 5% discount rate, amounts to US$ 274 million, generating an IRR of 58% after-tax. For a US$ 1,584 per ounce gold price, NPV after-tax is US$337 million and 70% IRR.

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26.	RECOMMENDATIONS

26.1

Drilling

●

Continue infill and extension drilling at existing Inferred prospects including Gabriela, Deborah and Esperanza;

●

Continue with drilling at other encouraging known prospects;

●

Continue with an aggressive infill and extension drilling program at the recently discovered high grade Zoe prospect;

●

Prepare a maiden resource estimate for the Zoe prospect as soon as sufficient drilling data is available to allow an updated PEA on the Cerro Moro Gold-Silver Project to undertaken;

●

Continue with current regional exploration strategy;

●

Progress exploration activities in close proximity to Escondida to either define additional resources or to locate appropriate 'barren' areas for location of open pit waste dumps and other near mine infrastructure;

●

Continue to develop a 'project scale' geological, structural and mineralisation model to assist with identifying additional exploration targets;

●

Continue to update relevant resource estimates for all significant mineralised zones to reflect ongoing changes to mineral resource confidence and the overall Cerro Moro resource base;

●

Continue with engineering and mine planning studies to confirm the economic viability of the project, and move forward to Pre-Feasibility stage in order to convert current resources into reserves;

●

Form an opinion as to the likely grade control requirements for both open pit and underground mining scenarios to be used as inputs in any ongoing mining studies;

26.2

Mining

For the next stage of engineering Cerro Moro, in underground mining is recommended:

●

While geomechanical analyses made are valid, they are based on empirical analysis, so they must be improved to have better information geomechanics.

●

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●

Analysis should be performed by the failure factor, which successfully integrated into the design of caves the presence of major faults near the exposed surfaces.

●

The effects due to blasting activities are usually ignored. Poor quality blasting can create new fractures and relax rock blocks, which consequently reduces the quality geotechnical rock mass.

●

In general, when studying the stability of a crown pillar, can identify six major failure modes, each of which can be given alone or a combination of more than one of them. In general, the dominant mechanism depends on the geotechnical characteristics of each sector under study. These mechanisms are: Structurally Controlled Fault, Fault rock mass stress-induced failure. Failure type fireplace, beam or plate failure rate and failure rate for blocks voussoir.

●

Define acceptability criteria

●

Using the updated resource estimation, which will include a significant proportion of higher confidence indicated category resources and utilizing the results of the pre-feasibility study, a mineral reserve estimation will be developed. With these results, detail mine planning and production schedules will be generated for both underground and open pit options.

26.3	Infrastructure

The following additional investigations are recommended for the next stage of project development:

●

Conduct further assessment of water supply locations and facilities;

●

Further define the tailings storage facilities requirements;

●

Critically asses the accommodation requirements for construction and operation labour;

26.4	Tailings Storage Facilities

The following recommendations have been made for the tailings storage facility:

●

At a subsequent engineering stage, the diversion system individual works design must be developed in detail.

●

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26.5	Metallurgical Testwork

The following additional test work is recommended:

●

Test work to optimise the consumption of the cyanide, copper sulphate in flotation and zinc because of the significant impact these reagents have on operating cost;

●

Test work to quantify the benefits of a dedicated Merrill Crowe circuit to recover precious metals from the pregnant solution generated by intensive cyanidation;

●

Test work to quantify the benefits of electrowinning of precious metals from pregnant solution generated by intensive cyanidation;

●

Test work on any new potential ore bodies to verify metallurgical performance;

●

Test work on individual ore bodies to be repeated using a P80 of 500 µm for the flash flotation test work as per vendor recommendations and with intensive cyanidation of the resultant concentrates. The previous test work used a P80 of 250 µm and did not test the intensive cyanidation of the resultant concentrates.

26.6	Site Geotechnical

It is recommended that the next phase of the project development should include site geotechnical investigations based on major equipment locations as shown in the preliminary layouts;

26.7	Risk Assessment

A high level risk assessment has been conducted in the process of developing this study, it is recommended that this risk assessment is expanded into contract execution and construction activities during the next phase.

26.8	Proposed Budget

Optimisation of the project budget based on all of the additional recommendation made above is a critical outcome of the next phase of the project development.

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Second Preliminary Economic Assessment

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27.	REFERENCES

●

Callan, N.J., 2008: Report to Accompany 1:10,000 Scale Geological Mapping, Cerro Moro Au-Ag Project, Puerto Deseado District, Santa Cruz Province, Patagonian Argentina., For Exeter Resource Corporation., Unpublished Company report dated 30 June 2008.

●

Coupland, T., 2010 (Cube, March 2010): Independent Technical Report – Update Report on the Cerro Moro Project, Santa Cruz Province, Argentina. December 31st 2009, Cube Consulting Pty Ltd, Published (on SEDAR) Company report dated March 18, 2010.

●

Coupland, T., 2010 (Cube, May 2010): Independent Technical Report – Resource Estimation for the Cerro Moro Project, Santa Cruz Province, Argentina., Cube Consulting Pty Ltd, Published (on SEDAR) Company report dated May 31, 2010.

●

Perkins, J. & Williams, M.T., 2009 (Exeter, 2009a): Technical Report – Cerro Moro Project, Santa Cruz Province, Argentina., Exeter Resource Corporation, Published (on SEDAR) Company report dated February 9, 2009.

●

Williams, M.T., 2009 (Exeter, 2009b): Technical Report – Cerro Moro Project, Santa Cruz Province, Argentina., Exeter Resource Corporation, Published (on SEDAR) Company report dated October 30, 2009.

●

Comminution Test Work Conducted Upon Samples of Ore from Cerro Moro Gold Project, AMMTEC Report No A11721, December 2008;

●

Initial Metallurgical Test Work on the Cerro Moro Silver/Gold Project, Argentina; Metcon Report M1668, February 2009 including Roger Townend & Associates mineralogy report 22301;

●

Cerro Moro Au-Ag Leach Residue Mineralogy; MODA, June 2009;

●

Metallurgical Test Work Conducted Upon Samples of Ore from Cerro Moro Gold Deposit; AMMTEC Report No A12671, May 2010 and Report No A12791, August 2010;

●

Metallurgical Characterisation Tests - Cerro Moro Mineralised Shoots - Escondida Far West, Metcon Report M2026, March 2010;

●

Flowsheet Development Test Work on the Cerro Moro Au/Ag Project; Metcon Report M2014, November 2010;

●

Flowsheet Development Test Work on the Cerro Moro Au/Ag Project; Metcon Report M2254, 2011;

●

Thickening of Cerro Moro Gold Leach Feed, Outotec Report S1474TB, February 2011;

●

Metallurgical Test Work Conducted Upon Test Work Products for Cerro Moro Project, ALS AMMTEC Report No A13274, February 2011;

●

Cerro Moro Pressure Filtration Test Work, Ishigaki Oceania Pty Ltd, February 2011;

●

Transportable Moisture Test Report TML002-96, Australian Testing Sampling & Inspection Services Pty Ltd, March 2011.

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28.	DATE AND SIGNATURE PAGES

	Effective Date of report:	30th June 2011

	Completion Date of report:	2nd August 2011

	Bill Gosling, Senior Process Metallurgist, of GR Engineering Services Limited (FAusIMM), was responsible for the information provided for the metallurgy and process plant design;

	FAusIMM



	David (Ted) Coupland (BSc DipGeoSc CFSG ASIA MAusIMM CPGeo MMICA) Director – Geological Consulting - Principal Geostatistician, Cube Consulting Pty Ltd, was responsible for resource estimation, exploration, drilling and data verification;

	MAusIMM, CPGeo
	Date Signed: 2 August 2011


	Eduardo Rosselot, CEng The Institute of Materials, Minerals and Mining (CEng MIMMM, Membership Nº448843) )with NCL Ltda, was responsible for the mining related studies and economic valuation.

	CEng MIMMM
	Date Signed: 2 August 2011

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Second Preliminary Economic Assessment

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