Rio Alto Mining 6K Rio Alto Files Ni 43-101 Technical Report on the La Arena Project | SEC 3 Mar 15

Exhibit 99.2


	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

La Arena Project, Peru

Rio Alto Mining Limited

Technical Report (NI 43-101)

Prepared by Mining Plus Peru S.A.C. on behalf of Rio Alto Mining Limited

Effective Date: 31^stDecember 2014


INNOVATIVE AND PRACTICAL MINING CONSULTANTS	1


	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

DATE AND SIGNATURE PAGE

The “qualified persons” (within the meaning of NI 43-101) for the purposes of this report are as listed below. The effective date of this report is 31st December 2014. The report was completed and signed on the 27 February 2015.


[signed]

Mr. Enrique Garay	M Sc. (MAIG)
Vice President Geology
Rio Alto Mining Limited.
Signed on the 27 February 2015

[signed]

Mr. Ian Dreyer	B.App. Sc, MAusIMM(CP)
Corporate Development Geologist
Rio Alto Mining Limited.
Signed on the 27 February 2015

[signed]

Mr. Tim Williams,	FAusIMM
Vice-President Operation
Rio Alto Mining Limited.
Signed on the 27 February 2015

[signed]

Mr. Greg Lane,	FAusIMM
Chief Technical Officer,
Ausenco
Signed on the 27 February 2015

[signed]

Mr. Scott Elfen,	P.E,
Global Lead Geotechnical Services
Ausenco
Signed on the 27 February 2015

[signed]

Mr. Fernando Angeles,	M.Eng (Min), P.Eng
Senior Mining Consultant
Mining Plus Peru SAC.
Signed on the 27 February 2015


INNOVATIVE AND PRACTICAL MINING CONSULTANTS	2


	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

CONTENTS


1	EXECUTIVE SUMMARY			9
	1.1	Introduction		9
	1.2	Property Description and Location		9
	1.3	Ownership		9
	1.4	Geology and Mineralization		10
	1.5	Exploration		10
	1.6	Production		10
	1.7	Mineral Resource Estimate		11
	1.8	Mineral Reserve Estimate		13
		1.8.1	Oxide Mineral Reserves	13
		1.8.2	Sulfide Mineral Reserves	15
	1.9	Capital and Operating Costs		16
		1.9.1	Oxide Gold Project	16
		1.9.2	Sulfide Copper Project	17
		1.10	Interpretation and Conclusions	18
		1.11	Recommendations	19


2	ISSUER AND TERMS OF REFERENCE		20
	2.1	Sources of Information	20
	2.2	Site Visits	20
	2.3	Report Responsibilities	21
	2.4	Units of Measurements	21
	2.5	Other Abbreviations	21


3	RELIANCE ON OTHER EXPERTS	22


4	PROPERTY, DESCRIPTION AND LOCATION			23
	4.1	Property Location		23
	4.2	Mineral Tenure and Status		23
	4.3	Environmental Liabilities		26
	4.4	Permitting		26
	4.5	Annual Fees and Obligations		27
		4.5.1	Maintenance Fees	27
		4.5.2	Minimum Production Obligation	27
		4.5.3	Royalties, OEFA Contribution and OSINERGMIN Contribution	28
		4.5.4	Ownership of Mining Rights	30
		4.5.5	Taxation and Foreign Exchange Controls	30
		4.5.6	Environmental Laws	31
		4.5.7	Mine Development, Exploitation and Processing Activities	32
		4.5.8	Mine Closure and Site Remediation	32
		4.5.9	Worker Participation	32
		4.5.10	Regulatory and Supervisory Bodies	33
	4.6	Risks that may affect access, title, or the right or ability to perform work		33


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru


5	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY		35
	5.1	Project Access	35
	5.2	Physiography and Climate	35
	5.3	Hydrology	35
	5.4	Population Centres	37
	5.5	Surface Rights	37
	5.6	Local Infrastructure and Services	37
	5.7	Seismicity	39


6	HISTORY			40
	6.1	Ownership History		40
	6.2	Previous Mineral Resources		40
		6.2.1	Coffey Mining 2010	40
		6.2.2	Andes Mining Services (AMS) 2011	41
		6.2.3	Andes Mining Services (AMS) 2013	41
		6.2.4	Mineros Consultores SAC (MICSAC) 2014	42
	6.3	Previous Mineral Reserves		43
		6.3.1	Coffey Mining 2010	43
		6.3.2	Kirk Mining 2013	44
		6.3.3	Mining Plus 2014	45
	6.4	Production		47


7	GEOLOGICAL SETTING AND MINERALISATION		49
	7.1	Regional Geology	49
	7.2	Project Geology	52
	7.3	Mineralization	60
	7.4	Structural Geology	60
	7.5	Hydrothermal Alteration	61


8	DEPOSIT TYPES		65
	8.1	Introduction	65
	8.2	Deposit Types and Mineralization	65
	8.3	High-Sulfidation Epithermal Au	65
	8.4	Porphyry Cu-Au (Mo) Deposits	66


9	EXPLORATION		68
	9.1	La Arena Deposit	68
	9.2	Major and Regional Exploration Target	68


10	DRILLING			71
	10.1	Introduction		71
	10.2	Drilling Procedures		72
	10.3	Drilling Orientation		72
	10.4	Surveying Procedures		72
		10.4.1	Accuracy of Drillhole Collar Locations	72
		10.4.2	Down-hole Surveying Procedures	72
	10.5	Sterilization Drilling		73


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru


11	SAMPLE PREPARATION, ANALYSES AND SECURITY	74


	11.1	Sampling Method and Approach	74


11.1.1	Diamond Core Sampling	74
11.1.2	Reverse Circulation Sampling	74
11.1.3	Logging	74


11.2	Sample Security	75
11.3	Sample Preparation and Analysis	75


12	DATA VERIFICATION	77


12.1	Introduction	77
12.2	Analytical Quality Control	77


	12.2.1	2014 Quality Control	77


12.3	Bulk Densities	81
12.4	Drillhole Database	81
12.5	Data Type Comparisons	81
12.6	Adequacy of Data	85


13	MINERAL PROCESSING AND METALLURGICAL TESTING	86


13.1	Introduction	86
13.2	Oxide Deposit	86


13.2.1	Mineralogy	86
13.2.2	Leaching Performance of Sandstone Rock	87
13.2.3	Pre-Operations Test Program	87
13.2.4	Evaluation of Oxide Intrusive Leaching Properties	89
13.2.4.1	Test Work Description	89
13.2.4.2	Sampling	90
13.2.4.3	Oxide Intrusive Test Work Programs	91
13.2.4.4	2013 Oxide Intrusive Program	92
13.2.4.5	Early 2014 Oxide Intrusive Program	94
13.2.4.6	Late 2014 Oxide Intrusive program	96
13.2.5	Dump Leach Results for Economic Modelling	97


	13.3	Sulfide Deposit	99


13.3.1	Historical Test Programs	99
13.3.1.1	Stage 1 Testing	99
13.3.1.2	Stage 2 Testing	100
13.3.1.3	Stage 3 Testing	101
13.3.1.4	Stage 4 Testing	102
13.3.1.5	Stage 5 Testing	102
13.3.1.6	Stage 6 Test Work Description	103
13.3.1.7	Mineralogy	103
13.3.1.8	Metallurgical Sampling	104
13.3.1.9	Comminution Test Program	107


14	MINERAL RESOURCE ESTIMATES	115


14.1	Introduction	115
14.2	Database	117
14.3	Geological Modelling	118


	14.3.1	Geology	118


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru


14.3.2	Gold Estimation Domains	119
14.3.3	Copper Estimation Domains	121
14.3.4	Other Elements	121


14.4	Sample Selection and Compositing	121
14.5	Basic Statistics	121
14.6	Variography	124
14.7	Block Modelling	126
14.8	Grade Estimation	127
14.9	Model Validation	130
14.10	Ancillary Fields	132
14.11	Resource Classification	133
14.12	Mineral Resource	134


15	MINERAL RESERVE ESTIMATES	136


	15.1	Types of Materials	136


15.1.1	Oxide Material:	136
15.1.2	Sulfide Material:	136


15.2	Assumptions and Parameters	137
15.3	Pit optimization	139
15.4	Cut-off Grades	140
15.5	Mineral Reserve Statement	141


15.5.1	Oxide Mineral Reserves:	142
15.5.2	Sulfide Mineral Reserves	143


16	MINING METHODS	144


16.1	Geotechnical	144
16.2	Hydrogeology and Hydrology	147
16.3	Oxide Project Mine Layout	148
16.4	Sulfide Project Mine Layout	151
16.5	Mining	152
16.6	Mine Production Schedule	153
16.7	Mining Equipment	155


17	RECOVERY METHODS	157


	17.1	Oxide Process Plant	157


17.1.1	Processing Flow Sheet – Dump Leach	157
17.1.2	Dump Leach Process	157
17.1.3	Process plant	158


	17.2	Sulfide Process Plant	160


17.2.1	General	160
17.2.2	Design Criteria Summary	162
17.2.3	Plant Design Basis	162
17.2.4	Unit Process Selection	165
17.2.5	Comminution Circuit Sizing	166
17.2.6	Flotation Circuit Design	168
17.2.7	Concentrate Regrind	169
17.2.8	Concentrate Thickening and Filtration	169
17.2.9	Concentrate Storage and Load Out	170


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru


17.2.10	Tailings Thickening Disposal and Water Recovery	170
17.2.11	Reagents and Consumables	171
17.2.12	Water Services	172
17.2.13	Air Services	174


18	PROJECT INFRASTRUCTURE	175


18.1	Roads	175
18.2	Accommodation	175
18.3	Offices, Workshops and Storage	175
18.4	Laboratories	176
18.5	Fuel and Lubrication	176
18.6	Power Supply	176
18.7	Water Supply	176
18.8	Explosives	177
18.9	Leach Pad Design	178


18.9.1	Drainage and Geomembrane Liner System	179
18.9.2	Pregnant Solution Collection System	180
18.9.3	Operational requirements	180
18.9.4	Geotechnical Investigation	180
18.9.5	Dump Leach Stability	181
18.9.6	Access Road and Perimeter Diversion Channel	182


	18.10	Tailings Storage	182


	18.10.1	Calaorco Pit Tailings Storage Facility	182


	18.11	Sulfide Waste Rock Storage Facility	186


18.11.1	Waste Rock Production	186
18.11.2	Waste Rock Storage Facility Design	186
18.11.3	Waste Rock Disposal Sequence	188


	18.12	Site Infrastructure for the Sulfide Project	189


18.12.1	Design Criteria	189
18.12.2	Utilities	189
18.12.3	Sewage Treatment	189
18.12.4	Emergency Generator	189
18.12.5	On site Infrastructure	190
18.12.5.1	General	190
18.12.5.2	Off plot Facilities	190


19	MARKET STUDIES AND CONTRACTS		192
	19.1	Gold Sales	192
	19.2	Gold Market	192
	19.3	Copper Supply and Demand	194
	19.4	Contracts	195


20	ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT		197
	20.1	Environmental Risk	197
	20.2	Social	198
	20.3	Mine Closure	198


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru


21	CAPITAL AND OPERATING COSTS	199


21.1	Oxide Gold Project - Operating Expenditures	199
21.2	Oxide Gold Project - Capital Expenditures	200
21.3	Sulfide Copper Project - Operating Expenditures	201


21.3.1	Mining Cost	201
21.3.2	Processing Cost	202
21.3.3	General and Administration Costs	202
21.3.4	Total Operating Cost	202


	21.4	Sulfide Copper Project - Capital Expenditures	204


22	ECONOMIC ANALYSIS	205


22.1	Peruvian Mining Taxes and Royalty	205
22.2	Oxide Gold Project	206
22.3	Sulfide Copper Project	207


23	ADJACENT PROPERTIES	209


24	OTHER RELEVANT DATA AND INFORMATION	210


24.1	Oxide Project Development	210
24.2	Other	210


25	INTERPRETATION AND CONCLUSIONS	211


25.1	Mineral Resources	211
25.2	Mining and Mineral Reserves	211
25.3	Oxide Treatment	211
25.4	Sulfide Treatment	212
25.5	Project Infrastructure	212


25.5.1	General	212
25.5.2	Waste Dump Facilities	212


25.6	Contracts	213
25.7	Overall	213


26	RECOMMENDATIONS	214


26.1	Geology and Resources	214
26.2	Mining	214
26.3	Oxide Mineral Metallurgy	214
26.4	Sulfide Mineral Metallurgy	215
26.5	Sulfide Process Plant	215
26.6	Infrastructure	216
26.7	Social	216
26.8	Environmental	217


27	REFERENCES	218

28	CERTIFICATES	219


INNOVATIVE AND PRACTICAL MINING CONSULTANTS	8


	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru


1	EXECUTIVE SUMMARY

1.1 Introduction

Mining Plus Peru S.A.C. was commissioned by Rio Alto Mining Limited (Rio Alto), a reporting issuer in the Provinces of Alberta, British Columbia and Ontario whose common shares are listed for trading on the Toronto Stock Exchange (TSX), the New York Stock Exchange (NYSE), the Lima Stock Exchange (BVL) to prepare a NI 43-101 Technical Report (Report) of the La Arena gold-copper project (La Arena Project) in Peru.

The following two events have triggered an updated NI 43-101 report for La Arena Project.

1.	Updated gold oxide reserve and resource estimates for the oxide project is available as a result of additional data obtained from the 2014 reverse circulation infill drill program. Gold inventory has been updated as a result of the in-fill drilling program completed with updated cost estimates.

2.	The completion of a Pre-Feasibility Study on the sulfide Cu-Au deposit, also known as La Arena Phase II Project, has been finalized. A Pre-Feasibility Study was completed in January 2015 by Ausenco on the Cu-Au sulfide material located on the East side of the current oxide pit. There have been no changes in the mineral resources on the sulfide deposit.

This report has been completed having the effective date on December 31^st, 2014. All monetary dollars expressed in this report are in United States dollars ("$").

1.2 Property Description and Location

The La Arena Project is located in northern Peru, 480 km NNW of Lima, Peru, in the Huamachuco District. The project is situated in the eastern slope of the Western Cordillera, close to the Continental Divide at an average altitude of 3,400 metres above sea level. The region displays a particularly rich endowment of metals (Cu-Au-Ag) occurring in porphyry and epithermal settings, including the Lagunas Norte mine at Alto Chicama, the Comarsa mine, La Virgen mine, Shahuindo exploration project and Tres Cruces development project.

1.3 Ownership

The mineral concessions pertaining to the La Arena Project have a total available area of 33,140 hectares. They are fully owned and registered in the name of La Arena S.A. The mining concessions are in good standing. Based on publicly available information, no litigation or legal issues related to the mining concessions comprising the project are pending.

The mineral resource identified so far in the La Arena deposit is completely contained within the mining concession “Maria Angola 18”. This mining concession is free of any underlying agreements and/or royalties payable to previous private owners. However, the Ferrol N°5019, Ferrol N°5026 and Ferrol N°5027 mining concessions, which are partially overlapped by Maria Angola 18 (as detailed in


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

section 4) are subject to a 2% Net Smelter Return (NSR) royalty, payable to their previous owners. Mining concessions Florida I, Florida IA, Florida II, Florida IIA, Florida III and Florida IIIA are subject to a 1.6% NSR royalty. Mining concessions Peña Colorada, Peña Colorada I, Peña Colorada II and Peña Colorada III are subject to a 1.4% NSR royalty.

1.4 Geology and Mineralization

The La Arena (Au, and Cu-Au) project is located in a prolific metallogenic province that contains many precious and polymetallic mines and projects such as; Lagunas Norte (Au-Ag), Santa Rosa (Au), La Virgen (Au), Quiruvilca (Ag-Base Metals), Tres Cruces (Au), Shahuindo (Au-Ag) and Igor (Au-Cu).

The La Arena oxide project consists of gold containing oxide mineralization which is predominantly of an Epithermal High Sulfidation style, hosted in oxidized sandstone-breccia within the Chimu Formation.

The Cu-Au-(Mo) sulfide mineralization is a porphyry type, which is hosted in a multi-stage porphyry intrusion. The La Arena Porphyry Cu-Au-(Mo) outcrops to the east from Calaorco and Ethel zones. The style of mineralization is typically porphyritic with at least four intrusive events identified. The intrusive rocks vary from dacitic to andesitic; they are differentiated by texture and composition.

1.5 Exploration

Until the effective day of this report, total drilled meters at La Arena are 284,782m. These meters are evenly split between reverse circulation (RC), at 141,591m (49.7% of the total), and core drilling (DC) at 143,191m (51.3% of the total).

The oxide domain has 19,733 m of DC drilling and 114,281 m of RC drilling which makes a total of 134,014 meters drilled in this domain. The sulfide domain has 121,858 m of DC drilling and 28,910 m of RC drilling.

During the period 2014, 22,087m of RC drilling was completed into the oxide domain, and 4,487m was completed into the sulfide domain as part of the sterilization drilling program near the Cu-Au sulfide project.

1.6 Production

Operations on site are currently exploiting the oxide gold reserve and are called the gold oxide Project. Oxide ore has been mined from Calaorco and Ethel pits, with the Ethel now being exhausted. Ore is being truck dumped in 8 m lifts onto the dump leach pad, with no crushing or agglomeration required prior to irrigation. The open pits are mined by conventional drill and blast, load and haul methods in 8 m high benches. Loading is with 170 t face shovels and a fleet of predominantly 92 t dump trucks. The Table 1.6-1 shows the historical ore and waste production since the operations began in 2011.


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

Table 1.6-1 Historical Mine Production (as Mined)


	Ore Mined			Waste	Total Tonnes
Year	tonnes	Au (g/t)	Oz	tonnes	tonnes
2011	3,663,752	0.88	103,547	4,182,371	7,846,123
2012	8,266,964	0.82	217,128	12,953,447	21,220,411
2013	13,811,137	0.60	268,223	22,997,357	36,808,494
2014	15,274,666	0.52	256,375	17,332,132	32,606,798
Total	41,016,519	0.64	845,273	57,465,307	98,481,826

Cyanide leach solution is sprayed onto each leach pad cell for a nominal period of 60 days. The pregnant solution flows onto the geomembrane underlying the pad to a central collection point and into the pregnant solution pond. Pontoon mounted pumps in this pond are used to pump the solution to the adsorption, desorption and refining (ADR) plant located approximately 300 m north of the leach pad. The plant currently has the capacity to treat 36,000 t/d of ore. The process includes absorption onto carbon pellets and desorption in high caustic/high temperature leach columns. The carbon is sent to regeneration and the enriched solution is sent to electrowinning cells where a cathode is used to produce a fine-grained precipitate. The precipitate is filtered and dried at approximately 420^oC, which also evaporates the mercury, which is then captured for later disposal. This dried precipitate is smelted to produce doré bars of approximately 80% Au.

A summary of processing for the project to date is presented in Table 1.6-2. Ore dumped in the leach pad may differ from actual mined ore tonnes due to the ore rehandle from the stockpile to the leach pad.

Table 1.6-2 Leach Pad Statistics


Year	Ore Dumped (tonnes)	Head Au Grade (g/t)	Ounces Au Dumped (oz)	Ounces Au Poured (oz)	Recovery (%)



2011	2,466,882	1.01	80,452	51,145	77.0%
2012	7,964,954	0.84	214,090	201,733	86.8%
2013	13,148,713	0.62	261,232	215,395	85.6%
2014	16,232,916	0.50	263,940	222,492	86.1%
Sub-Total 2012-2014	37,346,583	0.62	739,262	639,620	86.1%
Total 2011 - 2014	39,813,465	0.64	819,714	690,765	85.2%

1.7 Mineral Resource Estimate

An updated model for the oxide resource was created in September 2014. The sulfide resource quoted in this report is the pre-existing January 2013 model. A small drilling program was completed in 2014 testing the sulfide breccia at the top of the sulfide domain. However, this was completed after the sulfide Cu-Au project study was commenced, and subsequent analysis shows that any changes to the sulfide model are not material to the economic viability of the Cu-Au sulfide project.


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

There have been additions to the oxide resource during 2014 due to additional data received from the RC infill drill program undertaken in the Calaorco area. The major changes to the oxide resource model for 2014, are:

The addition of a moderate amount of oxide resource on the western side of the Calaorco deposit and at depth in Calaorco due to a large infill RC drill program.
The deepening of the indicated oxide resource by between 20 and 100 m due to the additional data received from the drill program, and
The inclusion of more high-grade, Tilsa Style domains in the Calaorco area due to acquiring additional resource data in 2014.

Estimation methods for Au oxide domains have not changed since 2013, being Localised Uniform Conditioning (LUC) for low grade oxide mineralized domains and Ordinary Kriging (OK) for the oxide background material and Tilsa structures.

A summary of mine production reconciliation is presented in Table 1.7-1.

Table 1.7-1 Reconciliation of 2014 Resource Model – Oxide Total


	Resource Model (2014)			As-Mined			Variance to As-Mined
Year	Mt	Au (g/t)	Au (koz)	Mt	Au (g/t)	Au (koz)	Mt	Au (g/t)	Au (koz)
2011	4.6	0.77	112.7	3.7	0.88	103.5	-20%	15%	-8%
2012	10.0	0.71	230.0	8.3	0.82	217.1	-18%	15%	-6%
2013	13.2	0.57	240.8	14.5	0.59	273.1	10%	3%	13%
2014	13.2	0.51	217.4	15.3	0.52	256.4	15%	3%	18%
Project to Date	41.1	0.61	800.8	41.7	0.63	850.1	1%	5%	6%

The oxide resource is reported within an optimized undiscounted cash flow pit shell using metal prices of $1,400 / oz for Au and updated cost parameters. The resource in Table 1.7-2 is quoted at a 0.07 g/t Au cut-off grade with no constraints on copper, as high Cu grades can be blended and diluted in the open pit operation.


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

Table 1.7-2 Mineral Resource – Oxide Total (In Situ as at December 31st 2014)


La Arena - Oxide Gold Mineral Resources
(In Situ as at December 31st, 2014)
Within Optimized Pit Shell @ $ 1,400 /oz, cut-off grade 0.07 g/t Au
Classification	Material Type	Tonnes (‘000,000 t)	Au g/t	Cu %	Ag g/t	Mo ppm	Au (´000 oz)
Measured	Sediments	1.1	0.23	0.07	0.3	32.6	8
	Intrusive	9.4	0.28	0.15	0.4	61.6	86
	Colluvium	-	-	-	-	-	-
	Total	10.5	0.28	0.15	0.3	58.5	94
Indicated	Sediments	100.8	0.38	0.01	0.5	4.1	1,234
	Intrusive	19.7	0.22	0.06	0.7	9.7	137
	Colluvium	2.6	0.34	0.01	0.2	2.5	28
	Total	123.1	0.35	0.02	0.5	5	1,399
Measured and Indicated	Sediments	102.0	0.38	0.01	0.5	4.5	1,243
	Intrusive	29.1	0.24	0.09	0.6	26.5	223
	Colluvium	2.6	0.34	0.01	0.2	2.5	28
	Total	133.6	0.35	0.03	0.5	9.2	1,494
Inferred	Sediments	2.2	0.34	0.01	0.4	2.9	24
	Intrusive	0.3	0.14	0.01	0.1	2.1	1
	Colluvium	-	-	-	-	-	-
	Total	2.5	0.31	0.01	0.4	2.8	25

The sulfide resource is reported within an optimized undiscounted cash flow pit shell using metal prices of $1,400 / oz for Au and $3.50 / lb Cu and updated cost parameters. The resource in Table 1.7-3 is quoted at a 0.12% g/t Cu cut-off grade.

Table 1.7-3 Mineral Resources - Sulfide Total (In Situ as at December 31st 2014)


La Arena – Sulfide Copper/Gold Mineral Resources
(In Situ as at December 31^st, 2014)
Within Optimized Pit Shell ($ 1,400 /oz Au, $3.5 /lb Cu), Cut-Off Grade 0.12 %Cu
Classification	Tonnes (‘000 000 t)	Au g/t	Cu %	Ag g/t	Mo Ppm	Au (´000 oz)	Cu (‘000 lbs)
Measured	-	-	-	-	-	-	-
Indicated	274.0	0.24	0.33	0.4	38.5	2,124	2,013,930
Measured and Indicated	274.0	0.24	0.33	0.4	38.5	2,124	2,013,930
Inferred	5.4	0.10	0.19	0.4	40.7	18	22,074

1.8 Mineral Reserve Estimate

1.8.1 Oxide Mineral Reserves

Oxide mineral reserves have been updated as a result of material changes in the mineral resources and updated estimating inputs. The reasons for those changes are:


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

An updated resource estimate based on the new drilling campaign.
A new set of inputs for the sulfide Cu-Au pit which were taken from the latest Pre- Feasibility Study report from January 2015.
The cost and operating information gained from an additional year of commercial production of the oxide gold dump leach mine.

Oxide mineral reserves have been constrained to the final pit design based on an optimized pit shell. The mineral reserve has been estimated with measured and indicated oxide mineral resources only. The pit optimization input parameters used are listed in Table 1.8-1.

Table 1.8-1 Pit optimization input parameters for Oxide Mineral Reserves


Pit Optimization Parameters for Oxide Mineral Reserves
Mining Parameters	Units	Value
Mining Dilution Factor	factor	1.05
Mining Recovery Factor	factor	0.98
Mining Cost Sediments (direct & indirect)	$/t mined	2.08
Processing Parameters
Ore processing rate	Mt/y	13
Processing Cost Sediments	$/t leached	1.55
Processing Cost Intrusive	$/t leached	1.55
General & Administration Cost	$/t leached	1.22
Gold leaching recovery intrusive	%	83
Gold leaching recovery sediments	%	86
Economics Assumptions
Gold price	$/oz	1,200
Payable proportion of gold produced	%	99.9
Gold Sell Cost	$/oz	12.37
Royalties	%	1

The oxide mineral reserve, based on the December 31st 2014 Measured and Indicated Resource only, is summarized in Table 1.8-2. The Inferred resource and material below the cut-off grade of 0.1 g/t Au was reported as waste rock. The mineral reserves are reported as in-situ dry million tonnes and include 5% mining dilution and 98% mining recovery.


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Table 1.8-2 La Arena - Mineral Reserve Statement for Oxide Ore


Classification	Material	Tonnage	Au	Cu	Ag	Au
Classification	Type	(‘000 000 t)	g/t	%	g/t	(´000 oz)
Proven	Sediments	1.2	0.22	0.07	0.32	8.6
Proven	Intrusive	8.7	0.28	0.15	0.33	77.7
Proven Stockpiled	LG stockpile	0.3	0.24	0.14	0.33	2.3
Total Proven	Total	10.2	0.27	0.14	0.33	88.6
Probable	Sediments	80.9	0.42	0.01	0.42	1,085.2
Probable	Intrusive	12.3	0.27	0.06	0.84	105.7
Total Probable	Total	93.1	0.40	0.02	0.48	1,190.9
Proven and Probable	Sediments	82.1	0.41	0.01	0.42	1,093.8
Proven and Probable	Intrusive	21.0	0.27	0.10	0.63	183.4
Proven Stockpile	LG stockpile	0.3	0.24	0.14	0.33	2.3
Total Proven and Probable	Total	103.3	0.39	0.03	0.47	1,279.5

Intrusive ore hosted within the oxides cannot be separated as a different ore type for processing, as it needs to be blended with sediments in order to be leached effectively. The colluvium deposit was not included in the mineral reserve due to cost of moving the national highway at this time. However, the colluvium material inside the Calaorco Pit was included in the mineral reserves as sediments. The colluvium deposit is a small shallow unconsolidated deposit of approximately 2.0 million tonnes grading 0.34 g/t gold and it is located immediately South-East of the main Calaorco Pit.

1.8.2 Sulfide Mineral Reserves

When calculating the reserve for the sulfide resource a small CAPEX constrained project was considered with strict financial hurdles. The resulting reserve at this stage is only a small portion of the total resource. The main economic assumptions used in the sulfide pit optimisation are presented in the Table 1.8-3.

The sulfide pit will be operated as an extension of the current gold mine operation. As Rio Alto operates the oxide project, the assumptions used for the sulfide project were adapted to represent the operation of the porphyry pit.

Mineral resources classified as Measured and Indicated are reported as Proven and Probable mineral reserves respectively. There were no resources classified as Measured Resources within the sulfide pit limits. Table 1.8-4 presents the Reserves Statement as of 31st of December 2014.

The tonnage and grades reported as in-situ dry tonnes and using 98% mining recovery and 5% dilution.

The cut-off grade is 0.18% Cu.


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

Table 1.8-3 Pit optimization input parameters for Sulfide Mineral Reserves


Pit Optimization Parameters for Sulfide Mineral Reserves
Mining Parameters	Units	Value
Mining Dilution Factor	factor	1.05
Mining Recovery Factor	factor	0.98
Mining Cost	$/t mined	1.92
Processing Parameters
Ore processing rate	Mt/y	6.57
Processing Cost	$/t milled	4.61
Process Copper Recovery	%	Avg. 91.1%, Range 75.9 - 92.0
Process Gold Recovery	%	Avg. 38.9% Range 29.5 – 45.5
General & Administration Cost	M$/y	22.6
Economics Assumptions
Copper price	$/lb	3.0
Payable proportion of copper produced	%	96.5
Copper Sell Cost	$/lb	0.37
Gold price	$/oz	1,200
Payable proportion of gold produced	%	88.6
Gold Sell Cost	$/oz	8.0
Royalties	%	1.0

Table 1.8-4 La Arena - Mineral Reserve Statement for Sulfide


Category	Tonnage	Au Grade	Cu Grade	Cu Content	Au Content
Category	('000 000 t)	g/t	%	'000 lb	(‘000 Oz)
Probable	63.1	0.312	0.430	579,407	633.2

1.9 Capital and Operating Costs

1.9.1 Oxide Gold Project

The capital cost of the oxide project has been estimated by Rio Alto based on current operations.

Annual capital cost estimates are detailed in Table 1.9-1.


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Table 1.9-1 Annual Capital Cost for Oxide Gold Project (in ‘000 $)


		Capex Additions for Oxide Gold Project (‘000 dollars)
	2015	2016	2017	2018	2019	2020	Total
Construction	17,259	24,712	2,300	12,600	-	-	56,871
Plant	2,166	8,012	-	-	-	-	10,178
Community Relations	2,560		-	-	-	-	2,560
Permits & Engineering	1,396	-	-	-	-	-	1,396
Other Capex	2,103	2,000	-	-	-	-	4,103
Road Diversion	-	-	7,000	8,000	-	-	15,000
Land Purchases	14,548	-	-	-	-	-	14,548
Total Capex	40,032	34,723	9,300	20,600	-	-	104,656

Annual Operating cost estimates for the oxide gold project, broken down by major element, are detailed in Table 1.9-2.

Table 1.9-2 Annual Operating Cost for Oxide Gold Project (in ‘000 $)


		Total Operating Costs and Operating Profit (‘000 $)
	2015	2016	2017	2018	2019	2020	Total
Net Revenue	257,284	222,737	212,365	187,506	181,111	171,881	1,232,884
Total Operating Expenses	117,967	137,879	141,303	127,538	126,869	98,758	750,314
Closure Expenditures	1,500	1,500	1,500	1,500	1,500	1,500	9,000
Operating Profit (EBITDA)	137,817	83,358	69,562	58,467	52,743	71,624	473,570
Operating Profit Margin	53.57%	37.42%	32.76%	31.18%	29.12%	41.67%	38.41%

1.9.2 Sulfide Copper Project

An overall summary of operating costs for the sulfide project is presented in Table 1.9-3.

Table 1.9-3 Total Operating Cost Summary


		Units Cost
	MINING
	Mining Direct Cost ($ / t mined)	1.32
	Mining Maintenance ($/mined)	0.02
	Mining Indirects ($/mined)	0.60
	PROCESSING
	Power	1.73
	Reagents, Consumables	1.30
	Labour	0.66
	Maintenance, Mob, Equipment	0.93
	Total Processing Cost ($ / t milled)	4.61
	GENERAL AND ADMINISTRATION	$22.6 M/y


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

Ausenco has estimated an initial CAPEX of $314 million for the processing plant and all associated infrastructure such as camp relocation, power, drainage system, tailings facilities and contingency. An additional $61.2 million has been allocated for sustaining capital during the 10 year mine life, which totals an estimated CAPEX of $415.5 million (including $ 40 million Mine Closure), for the 63Mt starter project.

The sensitivity analysis on the NPV and IRR for different Cu prices is shown in the Table 1.9-4:

Table 1.9-4 Sensitivity Analysis for the Sulfide Project


					After Tax NPV
Interest Rate		2.5/lb Cu		2.75/lb Cu		3.0/lb Cu		3.5/lb Cu		4/lb Cu
0%	$	115,931,418	$	202,256,504	$	285,492,823	$	448,325,631	$	606,298,959
4%	$	57,148,889	$	124,695,433	$	189,213,692	$	314,925,219	$	436,670,624
6%	$	34,108,807	$	94,249,838	$	151,479,447	$	262,807,486	$	370,539,024
8%	$	14,441,544	$	68,205,343	$	119,203,512	$	218,270,395	$	314,068,574
10%	$	(2,366,623)	$	45,879,743	$	91,521,670	$	180,078,151	$	265,657,574
					After Tax IRR
		2.5/lb Cu		2.75/lb Cu		3.0/lb Cu		3.5/lb Cu		4/lb Cu
IRR		9.70%		15.41%		20.15%		28.22%		35.16%

1.10 Interpretation and Conclusions

The increase in the gold oxide Resource is primarily due to the definition of extra resource to the west and at depth in Calaorco, as a direct result of the 2014 drilling program.
Oxide Mineral Reserves have increased due to the physical extension of the mineralization of the oxide deposit reflected in the new Mineral Resource estimates. The gold price of $ 1,200 per ounce was not changed from the previous estimates and only costs were updated based on the performance of the year 2014.
The La Arena oxide mine continues to exceed budget expectations due to positive grade variances between resource models and mining, and the definition of additional resources at the mine.
The sulfide project reserve pit at 63 Mt is the starter pit which provides 10 years of steady mill feed at 18,000 t/d to the processing plant. A trade-off analysis conducted in Section 15 shows that this pit size represents the best discounted value for the project with lowest CAPEX. However, this pit is only a portion of a potentially larger pit from the 274 Mt resource.
The La Arena mine site has been connected to the Peru grid power supply since September 2014.


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1.11 Recommendations

Define the northern strike extensions to the current gold oxide Resource through ongoing RC infill and extensional drilling.
Review the mine production plan for the sulfide project and smooth the total rock moved per period. An opportunity to reduce the peaks and lows on the mine production schedule was identified which allow better equipment utilization.
An opportunity to reduce the size of the low grade stockpile exists with a detailed mine schedule.
Carry out additional leaching test on both blended and unblended material.
Carry out carbon adsorption tests with pregnant liquor in order to determine carbon loading capacity when high soluble copper samples are being leached.
Additional variability flotation tests at optimized conditions should be conducted with new samples from drill holes samples inside the current pit design.
Conduct leaching tests of pyrite concentrate and cleaner scavenger tailings using a regrinding stage to determine if economic recovery of gold from these streams is feasible.
Future plant investigations may include gravity gold recovery tests on the cleaner scavenger tailings to determine if gold losses in tailings can be reduced and to produce a gravity concentrate that can be combined with the final copper concentrate.
Perform additional testing on the sulfide waste rock facilities to better refine their physical and mechanical properties to further develop the stacking and the PAG waste rock encapsulation and leachate collection strategy;
A revision to the Calaorco tailings feasibility study is required to incorporate the changes in Calaorco pit geometry.
Complete purchasing the land required for the gold oxide project, for the public road deviation, and continued land purchases for the sulfide project.
The site closure plan needs to be updated with the new details of the proposed Sulfide operation.
Revise the detailed re-logging of the sulfide deposit, in 2015, and determine if a more selective model can be constructed with sufficient geological confidence to potentially lift grade and therefore advance the project further.


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2	ISSUER AND TERMS OF REFERENCE

Mining Plus Peru S.A.C. has been commissioned by Rio Alto Mining Limited (Rio Alto), a reporting issuer in the Provinces of Alberta, British Columbia and Ontario whose common shares are listed for trading on the Toronto Stock Exchange (TSX), the New York Stock Exchange (NYSE), the Lima Stock Exchange (BVL) to prepare an Technical Report (Report) of the La Arena gold-copper projects (La Arena Project) in Peru.

This report complies with the disclosure and reporting requirements set forth in the Toronto Stock Exchange Manual, National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101), Companion Policy 43-101CP to NI 43-101, and Form 43-101F1 of NI 43-101.

The report is also consistent with the ‘Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves’ of 2012 (the Code) as prepared by the Joint Ore Reserves Committee of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia (JORC).

Furthermore, this report has been prepared in accordance with the Code for the Technical Assessment and Valuation of Mineral and Petroleum Assets and Securities for Independent Experts Reports (the “VALMIN Code”) as adopted by the Australasian Institute of Mining and Metallurgy (“AusIMM”). The satisfaction of requirements under both the JORC and VALMIN Codes is binding on the authors as Members of the AusIMM.

Rio Alto, a producing issuer, nominated three of its employees to assist in the preparation of this report: Mr Tim Williams - Vice-President Operations, Mr Enrique Garay - Vice President Geology, and Ian Dreyer, Corporate Development Geologist. Each is a qualified person under NI 43-101 rules. Messrs Williams and Garay are both shareholders in Rio Alto; however Ian Dreyer does not hold any shares of Rio Alto.

The relationship of Mining Plus and Rio Alto is solely one of professional association between client and independent consultants. Mining Plus does not have any material interest in Rio Alto or related entities or interests.

2.1 Sources of Information

The authors have made all reasonable enquiries to establish the completeness and authenticity of the information provided and identified, and a final draft of this report was provided to Rio Alto along with a written request to identify any material errors or omissions prior to final submission.

2.2 Site Visits

Rio Alto’s employees visit the mine site in a regularly basis. Mr. Fernando Angeles from Mining Plus has visited La Arena mine from 08^thto 11^thof September 2014. Greg Lane from Ausenco has visited La Arena mine in December 2014. Scott Elfen from Ausenco visited site in August 2014.


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2.3 Report Responsibilities

Specific sections of the report that the Qualified Persons are responsible for are provided in Table 2.3-1 and are detailed further in the attached Qualified Persons certificates.

Table 2.3-1 Qualified Persons-Report Responsibilities


Who		Section
Enrique Garay	(Rio Alto)	Section 7, 8, 9, 10
Ian Dreyer	(Rio Alto)	Section 11, 12, 14
Tim Williams	(Rio Alto)	Section 4, 5, 15, 16, 17.1, 18 (except 18.10 and 18.11), 19, 20, 21, 22
Fernando Angeles	(Mining Plus)	Section 2, 3, 6, 23, 24
Greg Lane	(Ausenco)	Section 13, 17.2
Scott Elfen	(Ausenco)	Section 18.10 18.11
Combined (All)		Section 1, 25, 26

2.4 Units of Measurements

All monetary dollars expressed in this report are in United States dollars (“$”). Quantities are generally stated in the International System Units. Metal content is expressed in troy ounces (Au) and pounds (Cu).

2.5 Other Abbreviations

A listing of other abbreviations used in this report is provided in Table 2.5-1 below.

Table 2.5-1 List of Abbreviations


Abbr.	Description	Abbr.	Description
$	United States of America dollars	koz	Thousands of troy ounces
“	Inches	kW	kilowatt
	microns	lb	pound (weight)
AAS	atomic absorption spectrometry	M	million
ADR	adsorption, desorption and refining	Moz	million troy ounces
ARD	acid rock drainage	Mt	millions of dry metric tonnes
Au	gold	Mt/y	million tonnes
bcm	bank cubic metres	MW	Megawatt
CaO	calcium oxide	NI	National Instrument
Cu	copper	NPV	net present value
DMT	dry metric ton	NSR	net smelter return
EIA	environmental impact assessment	oz	troy ounce
g/t	grams per tonne	ppm	parts per million
ha	Hectare	QA/QC	quality assurance quality control
hp	horse power	t	metric tonnes
IRR	internal rate of return	t/y	tonnes per year
k	Thousand	t/d	tonnes per day


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3	RELIANCE ON OTHER EXPERTS

This report has been prepared by Mining Plus Peru SAC (Mining Plus) for Rio Alto Mining Limited (Rio Alto) based on a high level assessment conducted on La Arena gold-copper Project in Peru. The information, conclusions, opinions, and estimates contained herein are based on:

Information available to Mining Plus at the time of preparation of this report;
Assumptions, conditions, and qualifications as set forth in this report; and
Data, reports, and other information supplied by Rio Alto and others sign-off parties.

For the purpose of this report, Mining Plus has relied on ownership information provided by Rio Alto. Mining Plus has not researched property title or a mineral rights for the property and expresses no opinion as to the ownership status of the property.

Mining Plus has relied on Tim Williams - Vice President Operations of Rio Alto, Ian Dreyer - Corporate Development Geologist of Rio Alto and Enrique Garay - Vice President Geology of Rio Alto, for the information regarding Mineral Resources and Mineral Reserves Estimates. It has also relied on the information provided by Ausenco on mineral processing aspects for the sulfide project and the leaching processing aspect for the oxide project.

Mining Plus has relied on the outputs resulting from Rio Alto financial model of La Arena Project and the application of taxes, royalties, and other government levies or interests, applicable to revenue or income from La Arena Project.


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4	PROPERTY, DESCRIPTION AND LOCATION

4.1 Property Location

The La Arena Project is located in Northern Peru. It is 480 km NNW of Lima, the capital of Peru, see Figure 4.2-1. Access to La Arena is 710 km on paved highways or upgraded roads from Lima. Politically, La Arena falls within the Huamachuco district, Sánchez Carrión province and Region of the La Libertad. The average altitude is 3,400 meters above sea level (m.a.s.l.) and the project is located in the eastern slope of the Western Cordillera, close to the Continental Divide with local rivers flowing towards the Atlantic Ocean.

The geographic and UTM coordinates of the gold and copper mineralization are as Table 4.1-1:

Table 4.1-1 Project Coordinates


Geographic		UTM (PSAD 56)
Latitude	07 ° 50’ S	North	9,126,360
Longitude	78 ° 08’ W	Este	816,237

4.2 Mineral Tenure and Status

The mineral concessions of the La Arena Project fall within a total area of 33,140 hectares. The concessions are 100% owned by and registered in the name of La Arena S.A.

The mining concessions are in good standing. Based on publicly available information, no litigation or legal issues related to the mining concessions comprising the project are pending. See section 4.5 below for the fees and activities required to keep these concessions active.

Mining concessions Florida I, Florida IA, Florida II, Florida IIA, Florida III and Florida IIIA are subject to a 1.6% NSR royalty. Mining concessions Peña Colorada, Peña Colorada I, Peña Colorada II and Peña Colorada III are subject to a 1.4% NSR royalty.


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

Figure 4.2-1 Project Location Map


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

Figure 4.2-2 La Arena Project Mining Concessions


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

4.3 Environmental Liabilities

By means of Ministerial Resolution No. 096-2010-MEM/DM, dated March 4, 2009, the General Mining Bureau of the Ministry of Energy and Mines has updated the “Preliminary Roster of Mining Environmental Liabilities (2006)” (“Roster”). From the legal review of the publicly available version of the above mentioned document, it has been identified that the following Mining Environmental Liability has been included in the Roster (see Table 4.3-1).

Table 4.3-1 Mining Environmental Liability


Name	Type	Coordinates		Mineral Right	Titleholder of the Mineral Right
		UTM PSAD 56
		East	North
La Florida I	Mining labour	823,378	9,124,708	Florida I	- Calcáreos Industriales Perú E.I.R.L. - IAMGOLD PERU S.A. - La Arena S.A. - Sociedad Minera Cambior Perú S.A.

Additionally, the following environmental damage was identified during the field work conducted for the 2006 Pre-Feasibility Study:

In the vicinity there is an old mine called Tambo Chiquito Mine (the former Florida Mine), which drains from a coal mine on the left bank of the Yamobamba river. This is an old underground mine located 10km South East from La Arena which was abandoned approximately 50 years ago. There are still ruins from the plant, abandoned camps and offices, as well as three small waste dumps with a total of 6000m³of tailings which are not confined.
Drainage of residual acidity and mine water (pH 3.5) is occurring to the Tambo Chiquito Creek, which is a tributary to the Yamobamba River. However the creek is now stabilized and does not represent a significant environmental risk to the Yamobamba River at present.

The environmental liability that may have been generated by previous exploration activities at La Arena is not significant, and is being managed in an environmentally efficient way, in close coordination with the community and/or individual owners who may also have been involved in such activities. La Arena has completed a survey to update and identify the existence of any other environmental liabilities. The results of the study were reported to the Ministry of Energy and Mines. No significant environmental liabilities were found.

4.4 Permitting

The La Arena Project is subject to various Peruvian mining laws, regulations and procedures. Mining activities in Peru are subject to the provisions of the Uniform Code of the General Mining Law (“General Mining Law”), which was approved by Supreme Decree No. 14-92-EM, on June 4, 1992 and


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its subsequent amendments and regulations, as well as other related laws. Under Peruvian law, the Peruvian State is the owner of all mineral resources in the ground. The rights to explore for and develop these mineral resources are granted by means of the “Concession System”.

Mining concessions are considered immovable assets and are therefore subject to being transferred, optioned, leased and/or granted as collateral (mortgaged) and, in general, may be subject to any transaction or contract not specifically forbidden by law. Mining concessions may be privately owned and the participation in the ownership of the Peruvian State is not required. Buildings and other permanent structures used in a mining operation are considered real property accessories to the concession on which they are situated.

4.5 Annual Fees and Obligations

4.5.1 Maintenance Fees

Pursuant to article 39 of the General Mining Law, titleholders of mining concessions pay an annual Maintenance Fee (Derecho de Vigencia). It is due on June 30 and is effective for the following year The fee is $ 3.00 per hectare. The failure to make Validity Fee payments for two consecutive years causes the termination (caducidad) of the mining concession. However, according to article 59 of the General Mining Law, the payment for one year may be delayed with penalty and the mining concessions remain in good standing. The outstanding payment for the past year can be paid on the following June 30 along with the future year.

4.5.2 Minimum Production Obligation

Legislative Decree 1010, dated May 9, 2008 and Legislative Decree 1054, dated June 27, 2008 amended several articles of the General Mining Law regarding the Minimum Production Obligation, establishing a new regime for compliance (“New MPO Regime”).

According to the New MPO Regime, titleholders of metallic mining concessions must reach a minimum level of annual production (“Minimum Production”) of at least one (1) Tax Unit¹or “UIT” per hectare,1 within a period of ten years. The ten periods begins on January 1st of the year following granting of the concession.

In the case of mining concessions that were granted on or before October 10, 2008 (as is the case of the mining concessions of La Arena), until the ten (10) year term for reaching Minimum Production established by the New MPO Regime elapses (on January 1st, 2019), these mining concessions will be subject to the former provisions of the General Mining Law.

___________________________
¹Pursuant to Supreme Decree 304-2013-EF, dated December 11, 2013, the Tax Unit for the year 2014 was set at S/.3,800 (approximately $1,360)


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Once the deadline to comply with the minimum production of the New MPO Regime has passed, if the company fails with compliance of production, it will be obligated to pay the Penalty of the New MPO Regime, and will be subject to the termination of the mining concession.

4.5.3 Royalties, OEFA Contribution and OSINERGMIN Contribution

In June 2004, Peru’s Congress approved royalties to be charged on mining operations. By Law Nº 29788, has modified the mining royalty regime starting on October 1st, 2011.

These new mining royalties are going to be determined quarterly and shall be applied to the quarterly operational profit. This rate shall be determined according to the quarterly operating margin, according to the following chart:

Table 4.5-1 Royalties: Cumulative progressive scale


a	b		c
N°	Operative Lower Limit	Operative Upper Limit	Marginal Rate
1	0	10%	1.00%
2	10%	15%	1.75%
3	15%	20%	2.50%
4	20%	25%	3.25%
5	25%	30%	4.00%
6	30%	35%	4.75%
7	35%	40%	5.50%
8	40%	45%	6.25%
9	45%	50%	7.00%
10	50%	55%	7.75%
11	55%	60%	8.50%
12	60%	65%	9.25%
13	65%	70%	10.00%
14	70%	75%	10.75%
15	75%	80%	11.50%
16	More than 80%		12.00%

To calculate the royalty in function to the operating margin will proceed as follows:

And, Royalty = OP * EF


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Where:
OP : Operation Profit
EF : Effective Rate
Tmg_j : Marginal Rate
MgO : Marginal Operative of column b
Ul : Upper Limit of column b
Li : Lower Limit of column b
j: Sections from 1 and n-1
n: Number of the section where is the Marginal Operative

The operating margin is the result of the division of the quarterly operating profit with the income generated by the quarterly sales of the mining agent. The amount to be paid for mining royalties will be the major amount from the comparison between the rate applied to the quarterly operation profit, and the 1% of the revenues generated by quarterly sales. In the case of the small scale mining titleholders, the mining royalty would be of 0%.The payment of the mining royalty is considered an expense when determining the corporate Income Tax.

OSINERGMIN is the government agency of record to inspect and audit the compliance with safety, job-related health and mine development matters.

The Supreme Decree 128-2013-EF, published on December 19th, 2013, the government established the rate applicable for the OSINERGMIN contribution. This payment will be made by all large and medium scale mining titleholders and it is calculated on the value of the monthly operating costs, corresponding to all their activities directly related to OSINERGMIN minus the Valued Added Tax and the Municipal Promotion Tax.

Rates by year:

2014: 0.21%
2015: 0.19%
2016: 0.16%

OEFA is the government agency of record that inspects and audits mining projects operations in order to secure compliance with environmental obligations and related commitments.

The Supreme Decree 130-2013-EF, published on December 19th, 2013, the government established the rate applicable for the OEFA Contribution. This payment will be made by all large and medium scale mining titleholders and it is calculated on the value of the monthly costs corresponding to all their activities directly related to OEFA minus the Valued Added Tax and the Municipal Promotion Tax.


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Rates by year:

2014: 0.15%
2015: 0.15%
2016: 0.13%

4.5.4 Ownership of Mining Rights

Pursuant to the General Mining Law:

Mining rights may be forfeited only due to a number of enumerated circumstances provided by law (i.e. non-payment of the maintenance fees and/or noncompliance with the Minimum Production Obligation);

The right of concession holders to sell mine production freely in world markets is established. Peru has become party to agreements with the World Bank’s Multilateral Investment Guarantee Agency and with the Overseas Private Investment Corporation.

4.5.5 Taxation and Foreign Exchange Controls

A recent modification on the tax law approved by the government reduces the cooperate taxes starting in year 2015. The law progressively decreases the tax from 30% (applied until 2014) to 26% (2016 onward). The new law reduces the rate of corporate income tax and increase the tax rate on dividends as shown in the following Table 4.5-2:

Table 4.5-2 Corporate Income Tax


Fiscal Years	Corporate Income Tax	Dividends
2015 – 2016	28%	6.8%
2017 – 2018	27%	8%
2019 – forward	26%	9.3%

There are currently no restrictions on the ability of a company operating in Peru to transfer dividends, interest, royalties or foreign currency in to or out of Peru or to convert Peruvian currency into foreign currency.

Congress has approved a Temporary Net Assets Tax, which applies to companies subject to the General Income Tax Regime. Net assets are taxed at a rate of 0.4% on the value exceeding one million Peruvian soles (approximately $345,000). Taxpayers must file a tax return during the first 12 days of April and the amounts paid can be used as a credit against Income Tax. Companies which have not started productive operations or those that are in their first year of operation are exempt from the tax.


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The Tax Administration Superintendent is the entity empowered under the Peruvian Tax Code to collect federal government taxes. The Tax Administration Superintendent can enforce tax sanctions, which can result in fines, the confiscation of goods and vehicles, and the closing of a taxpayer’s offices.

4.5.6 Environmental Laws

Under Legislative Decree 1013, approved on May 14, 2008, the government created the Ministry of the Environment to coordinate all environmental matters at the executive level. Currently, the Ministry of the Environment is still being implemented and its areas of responsibility are being defined, but it has already assumed, and is likely to continue to assume further responsibilities currently held by other ministries and supervisory agencies.

Environmental Legal Framework Applied to Mining Activities

The “Environmental Regulations for the activities of Exploitation, Processing, Transport, Anciliary Works and Development of Mining and Metallurgic Activities”, are the controlling regulatory acts that establish, among others, the environmental requirements to conduct mining activities within the Peruvian territory.

Under this legal framework, the General Bureau of Environmental Affairs (“DGAAM”) of the Ministry of Energy and Mines (“MEM”) is the responsible governmental agency to approve the environmental studies required to undertake mining activities in Peru, while the Environmental Inspections and Auditing Bureau (OEFA) of the Ministry of the Environment is currently the agency responsible for the inspection and auditing of mining projects and operations in order to confirm compliance with environmental obligations and related commitments.

Mining Exploration Activities

In connection with the environmental aspects specifically related to the development of mining exploration projects, currently these are governed by the Regulations on Environmental Protection for the development of Mining Exploration Activities.

Pursuant to these regulations, depending on the scale and impact of the exploration activities to be conducted, mining exploration projects are classified into the following two categories:

Category I: Before conducting exploration activities under this category, title holders will submit a DIA and have it approved by the DGAAM.

Category II: In order to conduct exploration activities under this category, title holders will have an EIAsd approved by the DGAAM of the MEM plus the permits outlined below.

Notwithstanding the above, it should be noted that the approval of the corresponding environmental certificate does not grant the titleholder the right to start conducting exploration


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activities, given that, titleholders of mining concessions are also required to obtain all governmental consents and permits legally required and the right granted by the landowner to use the surface land required.

4.5.7 Mine Development, Exploitation and Processing Activities

Pursuant to the “Environmental Regulations for the activities of Exploitation, Processing, Transport, Anciliary Works and Development of Mining and Metallurgic Activities”, prior to conducting mining and processing activities, titleholders of mining concessions must have an EIA approved.

A titleholder with an existing EIA that has executed mining activities such as exploration, exploitation, processing, closing or related activities or related to, and / or built components or modifications made without modifying its EIA , must report it to the DGAAM and OEFA within 60 business days from the entry into force of the Decree Supreme 040-2014-EM.

4.5.8 Mine Closure and Site Remediation

Exploration Activities

Regarding environmental remediation of areas affected by mining exploration activities, the “Regulations on Environmental Protection for the Development of Mining Exploration Activities”, establishes that titleholders of mining exploration projects will do “progressive closure”, “final closure” and “post closure” programs as outlined in the corresponding environmental study. Any amendment of the closure measures or its terms requires the prior approval of the DGAAM of the MEM.

Mining Development, Exploitation and Processing

Prior to the start-up of mining activities, including mine development, exploitation and processing, titleholders are required to have a Mine Closure Plan, duly approved by the DGAAM of the MEM in order to be authorized to carry out such activities.

Peruvian legal framework covering Mine Closure Plans includes a number of financial and legal requirements intended to ensure the completion of the closure obligations by the titleholders of mining projects. In case of non-compliance, these financial and legal requirements allow the mining authority to seize the financial guarantees from titleholders and complete the Mine Closure Plans as approved, preventing mining environmental liabilities.

4.5.9 Worker Participation

Under Peruvian law, every company that generates income and has more than twenty employees on its payroll is obligated to grant a share of its profits to its workers. For mining companies, the percentage of this profit-sharing benefit is 8% of taxable income. This profit-sharing amount made


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available to each worker is limited to 18 times the worker’s monthly salary, based upon their salary at the close of the previous tax year.

4.5.10 Regulatory and Supervisory Bodies

The five primary agencies in Peru that regulate and supervise mining companies are the Ministry of Energy and Mines (“MEM”), the National Institute of Concessions and Mining Cadastral (“INGEMMET”), the Supervisory Entity for the Investment in Energy and Mining (“OSINERGMIN”), the Labour Ministry (“MINTRA”) and, as previously described, the recently created Environmental Inspections and Auditing Bureau (“OEFA”) of the Ministry of the Environment.

The MEM promotes the integral and sustainable development of mining activities, as well as regulates all the activities in the Energy and Mines sector.

The INGEMMET is the Government Entity in charge of granting mining concessions, which entitles the concession holder the right to explore and exploit the area in which boundaries such concessions are located.

OSINERGMIN and MINTRA oversee regulatory compliance with safety, job-related health, contractors, and mine development matters, while OEFA oversees regulatory compliance with environmental regulation, investigating and sanctioning the breach of any environmental obligation.

4.6 Risks that may affect access, title, or the right or ability to perform work

Natural resources exploration, development, production and processing involve a number of risks, many of which are beyond the Company's control. Project and business risk factors and discussion on these are included in the Company’s quarterly Management Discussion and Analysis and the Annual Information Forms that are filed on SEDAR, the following list is a summary of those. Without limiting the foregoing, such risks include:

Changes in the market price for mineral products, which have fluctuated widely in the past, affecting the future profitability of the Company’s operations and financial condition.
Community groups or non-governmental organizations may initiate or undertake actions that could delay or interrupt the Company’s activities. See Social and Community Issues below.
The Company has limited operating history and there can be no assurance of its continued ability to operate its projects profitably.
Mining is inherently dangerous and subject to conditions or elements beyond the Company’s control, which could have a material adverse effect on the Company’s business.


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Actual exploration, development, construction and other costs and economic returns may differ significantly from those the Company has anticipated and there are no assurances that any future development activities will result in profitable mining operations.
Increased competition could adversely affect the Company’s ability to attract necessary capital funding or acquire suitable producing properties or prospects for mineral exploration and development in the future.
The Company’s insurance coverage does not cover all of its potential losses, liabilities and damage related to its business and certain risks are uninsured or uninsurable.
The Company depends heavily on limited mineral properties, and there can be no guarantee that the Company will successfully acquire other commercially mineable properties.
The Company’s activities are subject to environmental laws and regulations that may increase the cost of doing business or restrict operations.
The Company requires numerous permits in order to conduct exploration, development or mining activities and delays in obtaining, or a failure to obtain, such permits or failure to comply with the terms of any such permits that have been obtained could have a material adverse effect on the Company.
Exploration, development and mining activities on land within Peru generally require both ownership of mining concessions and ownership of or a leasehold interest over surface lands (“surface rights”).
The Company constantly seeks to expand its activities and may experience delays in obtaining surface rights or may not be able to acquire surface rights because of unwillingness by the owner of such rights to transfer ownership or the right to use at a reasonable cost or in a timely manner.
The Company may experience difficulty in attracting and retaining qualified management to meet the needs of its anticipated growth, and the failure to manage the Company’s growth effectively could have a material adverse effect on its business and financial condition.
Insofar as certain directors and officers of the Company hold similar positions with other mineral resource companies, conflicts may arise between the obligations of these directors and officers to the Company and to such other mineral resource companies.
Title to the Company’s mineral properties may be subject to prior unregistered agreements, transfers or claims or defects.
The Company’s business is subject to potential political, social and economic instability in the countries in which it operates.
Changes in taxation legislation or regulations in the countries in which the Company operates could have a material adverse effect on the Company’s business and financial condition.
The Company has no dividend payment policy and does not intend to pay any dividends in the foreseeable future.


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5	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Project Access

The project can be accessed via a 165 km national roadway from the coastal city of Trujillo directly east towards Huamachuco, passing through Chiran, Shorey/Quiruvilca and the Lagunas Norte project (Barrick Gold Corporation). The road is paved / sealed all the way. An air strip is also present at Huamachuco, a town of approximately 35,000 people located 21 km from La Arena that accommodates small airplanes. A private airstrip is also present at the nearby Lagunas Norte Mine operated by Barrick Gold Corporation.

5.2 Physiography and Climate

The topography in the project area is relatively smooth with undulating hills. Elevations vary between 3,000 and 3,600 meters above sea level. In general, the slopes are stable with grades varying between 16º and 27º, and the land is covered with vegetation typical of the area.

On the northern and southern flanks of the deposit, localized unstable areas exist where landslides have occurred during previous rainy seasons.

In Peru, the temperature normally varies according to the elevation, approximately 0.8°C per 100 m of elevation change. Average annual temperature data recorded from the La Arena meteorological station in 2013 is 10.6ºC. The maximum recorded temperature is 22.6°C and the minimum is 0.4ºC.

Historically, total average annual rainfall has been estimated in 1124 mm/annum and the average total annual evaporation rate in 733 mm/a. The average relative humidity varies monthly between 77 and 88%.

Maximum precipitation usually occurs during the months of October through to March while the months of June to September are the driest. The maximum daily precipitation recorded to date at the La Arena site is 245.6 mm and occurred in February 2012 while minimum precipitation was recorded in July 1998 with a total of 0 mm.

5.3 Hydrology

In September 2012 Golder Associates completed a hydrological study for the proposed new tailings site area that could be applied at the La Arena site. The study included the review and analysis of 13 regional meteorological stations located near the site. Regionally there is no hydrometric station data that can be used to determine surface water flows and calibrate the information obtained from the


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meteorological stations. Flow measurements have only been taken periodically as part of the environmental baseline studies for the sulfide project.

The climatological conditions of the La Arena site corresponds to typical climatic conditions found in the northern Sierra of Peru, where the weather is mainly controlled by the ground elevation as well as the geographical location on the eastern side of the western cordillera; the precipitation annual regime, and local specific climatic conditions.

Wind speed and direction varies according to the season. From June to September the monthly average wind speed is 4.5 m/s with east direction. From October to May, monthly average wind speeds are in the order of 3.7 m/s and from east to west direction.

Average annual evaporation has been estimated at 733 mm, with maximum and minimum values of 1029 mm and 555 mm respectively. The evaporation rate for the site is also controlled by the precipitation regime, and the average evaporation rate is lower from December to April (36 to 48 mm) and higher from June to September (70 to 94 mm).

Some 82% of the annual rainfall occurs during a six month period, from October through to April. For the site, average annual precipitation has been estimated at 1124 mm. Total annual precipitation values were also estimated for dry and humid years associated with return periods from 5 to 100 years, as shown in the Table 5.3-1.

Table 5.3-1 La Arena Annual Precipitation Values


Hydrological year	Return period (years)	Total annual rainfall (mm)
Dry	100	777
	50	803
	25	837
	10	900
	5	968
Average		1124
Humid	5	1277
	10	1367
	25	1466
	50	1532
	100	1593

The hydrological study aimed to determine the Probable Maximum Flow (PMF) for the area in order to estimate the size of hydraulic infrastructure for the project, such as dams and drainage systems.


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5.4 Population Centres

According to the EIA Modification (2013), in the direct area of influence (DAI) there are 1428 inhabitants residing in three communities: La Arena, La Ramada and Peña Colorada, being the most populated La Arena (47%). Within the DAI the estimated annual growth rate is 0.8%. A little more than a half of the total population (52%) is female.

Population within the area is predominantly young, with 52% of the people under 20 years of age. These results follow the demographic pattern of the country’s rural population. Households tend to be small, averaging less than 5 persons per household.

The young population moves temporarily or permanently in search for educational services (45%) and job (28%), mainly to Huamachuco (43%) followed by the city of Trujillo (33%). The majority of emigrants are women (55%).

Immigration to the local area is lower than emigration. Between 36% and 43% of people that live in the area were born in a different place. The majority of those who now live in the local area come from surrounding rural communities.

5.5 Surface Rights

Approximately 1,600 ha of surface lands will be required for the gold oxide and copper-gold sulfide project, of which 1,200 ha have been acquired. The gold oxide project requires approximately 700 ha, all of which have been acquired. In addition, the Company has acquired 65 ha of surface rights necessary to build an electrical substation to provide grid power to the gold oxide and copper-gold sulfide projects.

100% of the surface rights to be acquired are owned by individuals. Title for such land should be registered in the Public Registry (SUNARP). However the Company estimates that about 90% of the individual titles are properly registered. The Company is assisting land owners with the registration process so that negotiation for the transfer of legal title may proceed.

5.6 Local Infrastructure and Services

All existing and current facilities are designed and constructed to support the oxide gold mining and extraction activities. All working areas of the mine are accessible by well-maintained dual lane gravel roads. The ongoing brownfields drilling and copper sulfide feasibility study work are supported by these facilities.

The ADR (adsorption, desorption and refining) processing plant was expanded to a production capacity of 36,000 t/d in 2012. All the required pumping facilities have been installed for both the


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barren and the pregnant solutions, and construction has been completed for the pregnant leach solution pond and major storm pond (fences, lighting, security hut and associated utilities).

The site was connected to the Peru power grid in October 2014. All facilities are connected to the internal 22.9 kV power network supplied from a substation on site that receives power from the national grid. There are generators on site for backup power.

The dump leach pad construction is on-going. The mine development plan requires approximately 15 – 20 ha of dump leach pad to be constructed over the next 3 years.

An independent analytical and assay laboratory and a metallurgical laboratory (column leach testing) are both operational on site.

An industrial water purification plant has been installed to treat 220 m³/hour to a suitable quality for discharge to the environment.

Other associated facilities constructed in the processing plant are a reagent warehouse, a workshop and offices.

Camp and offices have been constructed on site with facilities to house 550 people. Other site infrastructure constructed in 2012 includes core shed, warehouse, mining workshop and equipment wash-bay, mine entrance and reception facilities and a highway underpass to access the waste dump #2 from the two pits.

The offices all have phone and data connection via microwave link to the Peruvian telephone network (Claro) with a total available bandwidth of 12 Mb/sec. A backup capacity of 512 kb/sec is also available and both services are expandable. A 3G cellular phone service has been installed under contract with a major Peruvian service provider. This cell phone service is also available to the general public as a community service provided by La Arena S.A.

Two bores supply water for the processing plant, camp, workshop and other facilities. One is an 80 m deep bore located approximately 1 km from the site offices with a nominal continuous flow capacity of 5 L/s and the other is located to the north of Calaorco pit with a nominal flow of 10 L/s. Sewage and waste water management facilities are installed.

Due to an increase of the oxide reserves on this report, the leach pad capacity was extended above the current waste dump area located to the south east of the Calaorco Pit. The locations and areas for waste dumps, tailings storage, dump leach pads, processing plant and other infrastructure are discussed further in Section 18 Project Infrastructure.


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5.7 Seismicity

In October 2012, Golder Associates completed a probabilistic and deterministic seismic hazard study for the site. The principal conclusions and recommendations are;

The La Arena project is located in the Peruvian Andes region with moderate to high seismic hazards, controlled by strong seismic sources associated with tectonic subduction zones and its relative location with the Peru-Chile trench. Historically earthquakes with a magnitude of M 8.0 and M 9.0 have occurred in the central and northern Peruvian cordillera.
Seismic hazard curves obtained for the site shows ground maximum horizontal acceleration values (PGA) of 0.28g, 0.37g and 0.52g, for 475, 975 and 2,475 years of return period respectively.
It has been estimated that sources of seismic hazard that controls the seismic parameters for the site are generated from moderate to strong earthquakes with magnitudes of M > 7.5 that are produced in the subduction zone associated with the superior and lower Nazca inter-plates at a distance approximately of 100 to 130 km from the site.
The Credible Maximum Control Earthquake (CMCE) has been estimated in magnitude M 8.0 at a distance of 104 km from the seismic source.
The Bureau of Reclamation (USA-Department of Interior) and the International Commission of Large Dams (ICOLD) defines the CMCE as a seismic parameter to consider for the design and validation of critical facilities and structures, such as tailings dams and waste dumps. The CMCE corresponds to a maximum horizontal ground acceleration PGA = 0.42g.


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6	HISTORY

6.1 Ownership History

The deposit was first discovered by Cambior geologists in December 1994. Cambior staked a claim for mining concessions of 1,800 ha over the deposit in January 1995. A further 70,000 ha of mining concessions were claimed in 1996, most of which have been allowed to lapse or have been sold. The mining concessions making up the La Arena Project passed to Iamgold following its acquisition of Cambior.

Rio Alto entered into an option and earn-in agreement with Iamgold Quebec Management Inc. in June 2009 which provided it with an option to acquire 100% of La Arena S.A., the Peruvian company that owns the La Arena Project, upon payment of $47.6 million cash, subject to certain adjustments and the completion of expenditure commitments.

On February 9 2011, Rio Alto announced that it had exercised its option and acquired 100% of the La Arena gold-copper project upon payment of the exercise price of $ 49.0 million.

6.2 Previous Mineral Resources

This section refers to the Mineral Resources statements reported for the La Arena deposit since 2010. Previous Mineral Resources estimates conducted by Cambior and Iamgold from October 1997 to February 2007 are discussed in the July 31, 2010 Technical Report.

6.2.1 Coffey Mining 2010

The Mineral Resource estimate by Iamgold was completed in August 2007 and was reviewed and validated by Coffey Mining in 2008. In 2010, the Mineral Resource was revised by Coffey Mining based on updated metal prices and pit optimization parameters. The Coffey Mining 2010 Mineral Resource is summarized in Table 6.2-1. Resources were confined within an pit shell based on $1,050/oz Au and $12/oz Ag for copper-poor mineralization largely in oxide sandstone (Cu < 300ppm) and a shell based on $3.00/lb Cu and $1,050/oz Au for copper-rich mineralization largely in primary and secondary porphyry.

Table 6.2-1 Coffey Mining Mineral Resource (as at July 31st 2010)


Material	Cut-	Category	Tonnage	Au Grade	Cu Grade	Ag Grade	Au	Cu	Ag
Material	off	Category	(Mt)	(g/t)	(%)	(g/t)	(‘000 oz)	(Mlb)	(‘000 oz)
Oxides	0.11g/t	Indicated	79.6	0.41	0.01	0.08	1,050		204
Oxides	Au	Inferred	9.2	0.19	0.01	0.29	57		86
Porphyry	0.1%	Indicated	225	0.27	0.35		1,932	1,722
Porphyry	Cu	Inferred	178	0.21	0.3		1,216	1,171


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6.2.2 Andes Mining Services (AMS) 2011

A Mineral Resource estimate was completed by AMS in September 2011 and is summarized in Table 6.2-2 and Table 6.2-3. Resources were confined within a pit shell based on $1,600 / oz for Au and $3.00 / lb for Cu. No credits were applied for Ag or Mo to derive the copper equivalent (CuEq) cut-off grade.

Au-oxide mineralization interpretations were created at a 0.10 g/t cut-off. Intrusive oxide above the cut-off grade criteria was included in the oxide resource, as this material was being mined as leach feed at the time of preparation of the Resource. The cut-off grade was 0.18% CuEq for the sulfide Resource. Cu-Au sulfide interpretations were based solely on geology with one domain for sandstone and another for intrusive.

Table 6.2-2 Mineral Resource – Oxide Total (as at September 30th 2011)


Resources⁽¹⁾	Tonnage	Au	Cu	Ag	Mo	Au
Resources⁽¹⁾	(Mt)	(g/t)	(%)	(ppm)	(ppm)	(‘000 oz)
Measured	10.3	0.67	0.01	0.6	8.3	221
Indicated	90.4	0.43	0.02	0.5	11.7	1,263
Measure + Indicated	100.7	0.46	0.02	0.5	11.4	1,484
Inferred	10.4	0.27	0.01	0.5	13.1	90

⁽¹⁾Au-Oxides resources reported at 0.10 g/t Au cut-off.

Table 6.2-3 Mineral Resource – Sulfide Total (as at September 30th 2011)


Resource⁽²⁾	Tonnage	Au	Cu	CuEq	Ag	Mo	Au	Cu
Resource⁽²⁾	(Mt)	(g/t)	(%)	(%)	(ppm)	(ppm)	(‘000 Oz)	(Mlb)
Indicated	312.7	0.24	0.29	0.48	0.7	42.9	2,422	2,007
Inferred	319.7	0.20	0.30	0.46	0.6	46.1	2,075	2,134

⁽²⁾Sulfide mineral resources reported at 0.18% CuEq cut-off.

6.2.3 Andes Mining Services (AMS) 2013

A major update on the Mineral Resource was conducted by AMS in 2012 and released in the Technical Report in January 2013. The updated resource was for both the oxide and sulfide component of the deposit which incorporates a full reinterpretation of the geology and mineralization, with the inclusion of significant additional drill data.

The Mineral Resources were reported within an optimized pit shell using metal prices of $1,800 / oz for Au and $3.50/lb for Cu. The cut-off grade for Au oxide Resources was 0.10 g/t Au and 0.13% copper equivalent (CuEq = Cu + Au x 0.396) for the sulfide Resource. No credits have been used for Ag or Mo to derive the CuEq cut-off grade.


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A summary of the Mineral Resources as per January 2013 is presented in Table 6.2-4 and Table 6.2-5.

Table 6.2-4 Mineral Resource – Oxide Total (as at January 1st 2013)


Resources⁽¹⁾	Tonnage	Au	Cu	Ag	Mo	Au
Resources⁽¹⁾	(Mt)	(g/t)	(%)	(ppm)	(ppm)	(‘000 Oz)
Measured	6	0.45	0.01	0.5	5.6	87
Indicated	116	0.42	0.01	0.5	4.2	1,571
Measured and Indicated	122	0.42	0.01	0.5	4.2	1,658
Inferred	5.4	0.37	0.01	0.3	2.7	65

⁽¹⁾Au-Oxides resources reported at 0.10 g/t Au cut-off.

Table 6.2-5 Mineral Resource – Sulfide Total (as at January 1st 2013)


Resource⁽²⁾	Tonnage	Au	Cu	CuEq	Ag	Mo	Au	Cu
Resource⁽²⁾	(Mt)	(g/t)	(%)	(%)	(ppm)	(ppm)	(‘000 Oz)	(Mlbs)
Indicated	561.7	0.21	0.3	0.39	0.4	42.9	3,829	3,745
Inferred	32.5	0.11	0.19	0.24	0.4	50.2	116	137

⁽²⁾Sulfide mineral resources reported at 0.13% CuEq cut-off.

6.2.4 Mineros Consultores SAC (MICSAC) 2014

During the year 2013, there was a RC drill program in the oxide domain (197 RC holes for 12,329 m) and a colluvium drill program (141 RC holes for 2,456m). New domains were created for the oxide portion in the Mineral Resource Update in January 2014. The most important new domain included was the oxide intrusive based on metallurgical test work completed in 2013.

The oxide domains were similar to the 2013 resource model parameters. The Table 6.2-6 shows the Mineral Resource estimates as at January 1st 2014.

The cut-off grade is 0.07 g/t Au for the Au oxide Resource. The major reason for the drop in cutoff grade from 2013 (0.10 g/t Au) is the change in power supply from diesel to grid power and resultant drop in power price, effective in late 2014.

Table 6.2-6 Mineral Resource – Oxide Total (as at January 1st 2014)


Resources⁽¹⁾	Tonnage	Au	Cu	Ag	Mo	Au
Resources⁽¹⁾	(Mt)	(g/t)	(%)	(ppm)	(ppm)	(‘000 oz)
Measured	2	0.43	0.04	0.5	8.4	28
Indicated	98.2	0.41	0.04	0.5	8.5	1,299
Measured and Indicated	100.2	0.41	0.04	0.5	8.5	1,327
Inferred	0.3	0.2	0.01	0.4	5.7	2

⁽¹⁾Au-Oxides resources reported at 0.07 g/t Au cut-off.


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There was not additional drilling or update to the sulfide resource.

6.3 Previous Mineral Reserves

This section refers to the Mineral Reserves statements reported from La Arena deposit since 2010 onward. Previous Mineral Reserves estimates conducted by Iamgold are discussed in the July 31, 2010 Technical Report.

6.3.1 Coffey Mining 2010

Mineral Reserves were updated by Coffey Mining and detailed in the July 31, 2010 Technical Report. All key inputs for both the Rio Alto gold oxide feasibility study work and the previous Iamgold PFS work were reviewed by Coffey Mining.

Rio Alto planned to proceed with a staged approach to the project, commencing mining and processing for the gold ore dump leach and once this is operational expand the project by mining and processing the copper ore. The key optimization input parameters used are shown in Table 6.3-1.

Table 6.3-1 Coffey Mining Pit Optimization Parameters 2010


Parameter		Dump Leach	Mill
Market Price		$950 per ounce Au / $2.30 per lb Cu
Mining cost	Sediment	$1.74 ore and waste	$1.74 ore and waste
($/t mined)	Porphyry	$1.82 ore and waste	$1.82 ore and waste*
Processing Cost ($/t Ore) G & A Cost		$1.55	$4.77
Processing Cost ($/t Ore) G & A Cost		$0.72**	$0.95
Mill Recovery	Au	80%	40%
Mill Recovery	Cu	0%	88%
Slope Angles Royalty		38º and 45º 1.70%

* Note that the mining cost was increased by $0.03/t for every 12m bench mined below elevation 3328mRL
** Note the G&A cost assumed an ore processing rate of 8.6Mt/y when Whittle work was done.

These mineral reserves were estimated using the following cut-off grades:

For oxide ore with Cu<300ppm (dump leach feed) 0.11 Au g/t.
For oxides with Cu>300ppm, secondary and primary sediments and porphyry 0.13% Cu.

The Coffey Mining 2010 Mineral Reserve estimate is summarized in Table 6.3-2.


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Table 6.3-2 Coffey Mining Mineral Reserve 2010


Ore Type	Oxide Ore			Secondary Ore			Primary Ore			All Ore
Ore Type	Mt	g Au/t	%Cu	Mt	g Au/t	%Cu	Mt	g Au/t	%Cu	Mt	g Au/t	Oz Au	%Cu	000’s lbs Cu
Gold Oxide Pit Design
Sediments	57.4	0.44								57.4	0.44	821,000
Sulfide Pit Shell (excluding Oxide Pit)
Sediments	2	0.57	0.11	0.1	0.34	0.32	0.1	0.81	0.6	2.1	0.58	39,000	0.14	7,000
Porphyry	13.1	0.3	0.2	13.2	0.36	0.52	160.1	0.28	0.38	185.2	0.29	1,709,000	0.38	1,567,000
Total Shell	15.1	0.34	0.19	13.3	0.36	0.52	160.2	0.28	0.38	187.3	0.29	1,748,000	0.38	1,574,000

6.3.2 Kirk Mining 2013

Mineral Reserves were updated by Kirk Mining Consultants and detailed in the 1st January 2013 Technical Report. The update on Mineral Reserve was conducted by Kirk Mining Consultants based on the Mineral Resources estimated by AMS (Table 6.2-4 and Table 6.2-5) and published together in a NI 43-101 Technical Report with an effective data of the 1^stJanuary 2013.

Oxide and sulfide Mineral Reserves were estimated from within a pit design. The pit design was based on pit optimization of the measured and indicated mineral resources. The economic assumptions and other parameters used by Kirk Mining to undertake pit optimization for both the oxide Au deposits and sulfide Cu-Au deposits are presented in Table 6.3-3.

Table 6.3-3 Kirk Mining Pit Optimization Parameters 2013


Pit Optimization Parameters 2013
Market Conditions	Value
Gold price per ounce	$1,400
Payable proportion of gold produced	99.90%
Copper price per pound	$3.00
Payable proportion of copper produced	96.50%
Minimum government royalty	1% of revenue
Mill Recovery	Value
Mining recovery of ore	98%
Overall pit slopes	34 to 39^o
Gold processing recovery (dump leach)	85%
Gold processing recovery (Sulfide plant to concentrate)	35%
Copper grade-recovery formula	Average 88%
Costs	Value
Mining cost per tonne of oxide ore (plus depth increment)	$2.38
Mining cost per tonne of sulfide ore (plus depth increment)	$2.44
Mining cost per tonne of waste (plus depth increment)	$2.50


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Processing cost per tonne of oxide ore (including pad expansion)	$2.06
Processing cost per tonne of sulfide ore (including tails dam lifts)	$3.99
Concentrate shipping and selling cost per tonne	$160
General and administration costs per tonne of ore	$2.45

The average, rounded up cut-off grade equates to approximately 0.12 g/t Au for the oxides and 0.15 % CuEq for sulfides. The Kirk Mining 2013 Mineral Reserve estimate is summarized in Table 6.3-4.

Table 6.3-4 Mineral Reserve – Oxide and Sulfide (as at 1st January 2013)


	Material	Classification	Oxide Ore		Sulfide Ore			Metal Mined
	Material	Classification	Tonnage (Mt)	Au (g/t)	Tonnage (Mt)	Au (g/t)	Cu (%)	Au (‘000 Oz)	Cu (‘000 lbs)
Likely oxide pit	Sediments⁽¹⁾	Proven	5.6	0.47	-	-	-	84	-
Likely oxide pit	Sediments⁽¹⁾	Probable	47.9	0.52	-	-	-	803	-
Final pit excluding oxide pit	Sediments⁽¹⁾	Proven	-	-	-	-	-	-	-
	Sediments⁽¹⁾	Probable	8	0.39	-	-	-	100	-
	Porphyry⁽²⁾	Proven	-	-	0.1	0.32	0.29	1	942
	Porphyry⁽²⁾	Probable	-	-	268.7	0.24	0.33	2,091	1,945,929
Pit Design	All	Proven + Probable	61.5	0.5	268.9	0.24	0.33	3,080	1,946,872

(1) Au-Oxides Reserves reported at 0.12 g/t Au cut-off. 98% Mining Recovery applied.

(2) Sulfide mineral reserves reported at 0.15% CuEq cut-off. 98% Mining Recovery applied.

The oxides reserves were increased in tonnes, grade and contained gold from the July 2010 Report by 7%, 14% and 20% respectively. The sulfide reserves were significantly increased (44%) compared to July 2010 due to the resource drilling program in 2012 and subsequent conversion of inferred resources to indicated resources. Metal expected to be mined within the porphyry pit design has increased by 20% for Au and 24% for Cu.

6.3.3 Mining Plus 2014

The oxide Intrusive material was added to the mineral reserves in 2014 after a series of tests performed on this material concluded that this oxide intrusive material could be processed when blended with sandstone. Table 6.3-5 shows the economic assumptions and technical parameters used in the pit optimization.


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Table 6.3-5 Pit Optimization Parameters for January 2014 report


Pit Optimization Parameters for Oxide Mineral Reserves
Mining Parameters	Units	Value
Mining Dilution Factor	factor	1.05
Mining Recovery Factor	factor	0.98
Mining Cost Sediments (direct & indirect)	$/t	2.99
Mining Cost for Intrusive	$/t	3.14
Processing Parameters	Units	Value
Ore processing rate	Mt/y	13.0
Processing Cost Sediments	$/t	1.53
Processing Cost Intrusive	$/t	1.65
General & Administration Cost	$/t	1.69
Gold leaching recovery intrusive	%	82
Gold leaching recovery sediment	%	85
Economics Assumptions	Units	Value
Gold price	$/oz	1,200
Payable proportion of gold produced	%	99.9
Gold Sell Cost	$/oz	12.37
Royalties	%	1

Given the above parameters, Mining Plus estimated the cut-off grade at 0.07 g/t Au for the Sediments and 0.1 g/t Au for the Intrusive material. The likely higher cost of processing intrusive material (because of the rehandling cost) was reflected in a higher cut-off grade.

Colluvium was not included in the estimates due to restrictions with infrastructure already in place.

The Table 6.3-6 presents the Mineral Reserve Estimates as January 01^st, 2014.


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Table 6.3-6 Mineral Reserve Estimates (as at December 31^st, 2013)


Classification	Material	Tonnage	Au	Cu	Ag	Au
Classification	Type	(DMT)	g/t	%	g/t	(´000 oz)
Proven	Sediments	1.4	0.45	0.01	0.44	20
Proven	Intrusive	0.2	0.38	0.26	0.34	3
Proven Stockpiled	LG stockpile	1.2	0.23	0.004	0.81	9
Total Proven	Total	2.8	0.35	0.03	0.59	32
Probable	Sediments	56.9	0.47	0.01	0.46	853
Probable	Intrusive	16.5	0.32	0.14	0.37	172
Total Probable	Total	73.4	0.43	0.04	0.43	1,025
Proven and Probable	Sediments	58.2	0.47	0.01	0.48	873
Proven and Probable	Intrusive	16.8	0.32	0.14	0.39	175
Proven Stockpile	LG stockpile	1.2	0.23	0	0.81	9
Total Proven and Probable⁽¹⁾	Total	76.2	0.43	0.04	0.47	1,056

(1) Au-Oxides Reserves reported at 0.07 g/t Au (Sediments) and 0.1 g/t Au (Intrusive).

(2) 98% Mining Recovery and 5% Mining dilution applied. Tonnage in Millions of In-situ Dry Metric Tonnes

The updated oxide mineral reserve estimate outlined in Table 6.3-6 is summarized as follows:

76.2 million tonnes grading 0.43 g/t gold in the proven and probable categories containing 1,056,000 ounces of gold ounces, which after mining 261,232 ounces during the January 1, 2013 to January 1, 2014 period represents an increase of 430,232 ounces (+49%) from the oxide mineral reserve estimate in the January 2013 Report (or a 19% increase from 887,000 ounces at January 1, 2013 to 1,056,000 ounces at January 1, 2014).

As the sulfide Mineral Resource remained unchanged, the sulfide Reserve was not updated in the technical report of January 2014.

6.4 Production

The historical mine production is tabulated in the Table 6.4-1. The mine production is the rock mined from the pit and do not include stockpile ore rehandle. The mine production has been reported until the effective day of this report (December 31^st, 2014)


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Table 6.4-1 Mine Production for 2014 from Oxide Gold deposits (as Mined)


Year	Ore Mined			Waste	Total Tonnes
Year	tonnes	Au (g/t)	Ozs	tonnes	tonnes
2011	3,663,752	0.879	103,547	4,182,371	7,846,123
2012	8,266,964	0.820	217,128	12,953,447	21,220,411
2013	13,811,137	0.603	268,223	22,997,357	36,808,494
2014	15,274,666	0.520	256,375	17,332,132	32,606,798
Total	41,016,519	0.640	845,273	57,465,307	98,481,826

A summary of processing for the project to date is presented in Table 6.4-2. Ore dumped in the leach pad may differ from actual mined ore tonnes in Table 6.4-1 due to the rehandle ore from the stockpile to the leach pad.

Table 6.4-2 Historical Ore Processed and Metallurgic Recovery


	Ore	Head Au	Ounces Au	Ounces Au	Recovery
Year	Dumped	Grade	Dumped	Poured
	(tonnes)	(%)	(oz t)	(oz t)	(%)
2011	2,466,882	1.01	80,452	51,145	77.0%
2012	7,964,954	0.84	214,090	201,733	86.8%
2013	13,148,713	0.62	261,232	215,395	85.6%
2014	16,232,916	0.50	263,940	222,492	86.1%
Sub-Total 2012-2014	37,346,583	0.62	739,262	639,620	86.1%
Total 2011 - 2014	39,813,465	0.64	819,714	690,765	85.2%


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7	GEOLOGICAL SETTING AND MINERALISATION

7.1 Regional Geology

The La Arena Deposit is located on the eastern flank of the Andean Western Cordillera in northern Peru. The area is underlain by sediments of the Mesozoic West Peruvian Basin which were folded and faulted during the Cenozoic deformation.

The regional stratigraphy is dominated at outcrop by the folded Upper Jurassic (Chicama Formation) to the Lower Cretaceous (Goyllarisquizga Group), which are mainly siliciclastic sediments, with lesser amounts of younger Lower-to-Upper-Cretaceous carbonate sediments occupying the cores of synclines. West of La Arena, the Cretaceous sediments are unconformably overlain by the Cenozoic volcanics of the Calipuy Group. The regional stratigraphical column is summarised in Table 7.1-1 and a plan of the regional geology is shown in Figure 7.1-1.

Table 7.1-1 Regional Stratigraphic Column of La Arena and Surrounding Areas


Erathem	System	Series	Group	Formation	Extrusive Lithology	Intrusive Gold Lithology Mineral’n Abbr
Cenozoic	Quaternary	Recent		Alluvial, Fluvial	Q-al/Q-fl
	Quaternary	Pleistocene		Glacial, Lacustrine	Q-gl/Q-la
	Neogene		Calipuy	Pn-ca	P-da	AC
	Paleogene		Calipuy	Pn-ca	P-and	AC
Mesozoic	Cretaceous	Upper		Yumagual	Ks-yu
		Lower		Pariatambo	Ki-pa
				Chulec	Ki-chu
				Inca	Ki-In
			Goyllarisquizga	Farrat	Ki-fa

				Carhuaz	Ki-ca	S
				Santa	Ki-sa
				Chimu	Ki-chi	AC, ET, LA, LV, SR
				Oyón	Ki-o
	Jurassic	Upper		Chicama	Js-ch
after Reyes R. L, 1980 and Navarro et. al. 2010). Gold Mineralization: AC: Lagunas Norte, ET: El Toro, LA: La Arena, LV: La Virgen, S: Shahuindo, SR: Santa Rosa

From oldest to youngest, the regional stratigraphy is described as follows:


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Figure 7.1-1 Regional Geology of La Arena


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Palaeozoic (and Precambrian): Constitute basement rocks to the east of La Arena along the River Marañon and the Eastern Cordillera. They are not exposed at La Arena or in the immediately surrounding area.

Mesozoic: The oldest outcropping rocks in the region belong to the Upper Jurassic Chicama Formation and consist of soft, laminated marine black shales with thin sandstone intercalations.

These pass upwards into the Lower Cretaceous shallow marine siliciclastic Goyllarisquizga Group, the lowest unit of which, the Oyon Formation, consists of fine-to-medium-grained sandstone and thinly-bedded shale, with some coal seams. Overlying the Oyon Formation are thickly-bedded, medium grained quartzitic sandstones of the Chimu Formation which constitutes the principal host rock for gold mineralization at Lagunas Norte, El Toro, La Arena, La Virgin and Santa Rosa. The remainder of the Goyllarisquisga Group (Santa, Carhuaz and Farrat formations) consists of generally finer grained siliciclastic units with interbedded minor carbonates. The Carhuaz Formation provides the host for gold mineralization at Shahuindo.

Overlying the Goyllarisquisga Group sediments are Lower-Cretaceous shallow marine carbonates of the Inca, Chulec, Pariatambo formations and the Upper Cretaceous Yumagual Formation.

The Mesozoic sediments were folded and faulted towards the end of the Cretaceous by the early stages of the developing Andean Orogeny.

Cenozoic: Calipuy Group, cordilleran arc volcanics unconformably overlie the folded and faulted Mesozoic strata south and west of La Arena. These sub-aerial volcanics are associated with Upper Miocene sub-volcanic intrusive bodies of andesitic to dacitic composition. The Calipuy volcanics are mainly tuffs with agglomerate horizons at the base, and inter-bedded with andesitic lavas. They constitute the host rock for high sulfidation, low sulfidation and polymetallic mineralization at Lagunas Norte, Tres Cruces and Quiruvilca respectively.

To the west of the area shown in Figure 7.1.1, the Coastal Batholith is emplaced in volcano-sedimentary strata of the Mesozoic Western Peruvian Trough, time equivalents of the rocks described above. Cenozoic intrusive rocks, including granodiorites, diorites and quartz–feldspar porphyries, are intruded as isolated stocks into both the Mesozoic sedimentary sequence and the overlying Calipuy volcanics. The age of those intrusions vary from c.a. 23 to 25 Ma. One of these intrusions hosts the porphyry-style mineralization at La Arena.

The main structural features of the region are associated with the Jurassic-Cretaceous sedimentary sequence and consist of a series of folds, reverse faults and over-thrusts trending generally NW-SE (see Figure 7.1-2.and Figure 7.2-1). Individual folds range up to 80 km in length and 5 km in width, and display various forms depending on the relative competency of the various stratigraphic levels. The highly competent sections of the Chimu Formation for example form structurally complex cores to the main anticlines, where they have resisted erosion better than the enclosing strata.


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Mesozoic sediments are affected by folds and reverse faults. Miocene intrusives were emplaced in the fold axes

Figure 7.1-2 Regional Cross Section - La Arena Project

The region is particularly well-endowed with mines and mineral occurrences varying from low-to-high sulfidation systems and from porphyry through polymetallic to epithermal deposits. Currently operating mines other than La Arena, include Quiruvilca (polymetallic Cu/Zn/Pb/Ag) and Lagunas Norte (Lagunas Norte), La Virgen and Santa Rosa (all epithermal Au).

7.2 Project Geology

The La Arena Project is located within a regional fold and thrust belt of predominantly Mesozoic sedimentary rocks. Sedimentary rocks in the project area have been intruded by intermediate-to-felsic porphyritic stocks which tend to occupy the cores of anticlinal structures as displayed in Figure 7.2-1.

Sedimentary rocks across the La Arena Project area consist of a lower, shallow-marine-to-deltaic, siliciclastic sequence followed by an upper, carbonate-dominated sequence, all of Lower Cretaceous age.

The oldest rocks exposed in the cores of anticlines are thinly bedded and laminated mudstones, minor siltstones and fine grained sandstones with occasional coal seams which make up the basal Lower Cretaceous Oyon Formation.


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Figure 7.2-1 Local Geology Plan - La Arena Project


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Overlying the Oyon Formation is the Goyllarisquizga Group (Chimu Formation). The Chimu Formation, is the principal host rock for epithermal gold at La Arena (and elsewhere in the region) has been sub-divided into the three members as shown in Figure 7.2-2 and described below (from oldest to youngest):

Transition Member (130 m) - consists of laminated fine-to-medium grained sandstones intercalated with siltstones and mudstones, and is a transitional facies between the more shaley Oyon Formation and the more sandy Lower Member of the Chimu Formation.
Lower Member (125 m) - consists of thickly bedded and compact medium-to-coarse grained sandstones which, due to their brittle nature, are fractured and often brecciated, and constitute the principal sedimentary host rock at La Arena. In addition to hosting the La Arena high-sulfidation Au mineralization, the Chimu Formation also hosts similar mineralization at Lagunas Norte, El Toro, La Virgin and Santa Rosa.
Upper Member (150 m) - consists of a mixed sequence of coarse-grained sandstones, laminated siltstones and carbonaceous mudstones.

Multiple intrusions of dacitic and andesitic feldspar porphyries have intruded the Cretaceous sedimentary sequence at La Arena (Figure 7.2-3).

The intrusive rocks vary from dacitic to andesitic composition. They are differentiated only by texture and composition. The early intrusion Feldspar Porphyry Dacitic One (FPD1) is generally barren. The second intrusion named FPD2 (previously named HA) is hosting most of the Cu-Au porphyry mineralization. The third intrusion stage is the FPD3 (previously named intra-mineral intrusion, HAI) is also associated with lower grade of Cu-Au mineralization. The final intrusive phase consists of barren Andesitic Dykes. Differentiating between individual intrusives is based principally on field and core observations.


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Figure 7.2-2 Lithostratigraphic Columns of the Chimu Formation


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Figure 7.2-3 Geology Section: Multiphase Intrusion Crosscutting the Sedimentary Rocks

In Table 7.2-1, the U-Pb age dating on zircons within individual intrusives, indicates overlapping dates at ~24.9 Ma (±0.4 a 0.7 Ma sigma errors). This would suggest that all three intrusives were emplaced in a short period of time (within 1-2 Ma intervals maximum).

Table 7.2-1 Age of Intrusive - La Arena Project


Intrusive	Avg. U-Pb Age	Max. age	Min. age
FPD3	24.7 ± 0.5	24.85 ± 0.49	24.51 ± 0.43
FPD2	24.48 ± 0.6	24.86 ± 0.58	24.1 ± 0.55
FPD1	25.12 ± 0.5	25.23 ± 0.39	24.89 ± 0.68

The following summary is presented for the three main intrusive phases identified at La Arena:

FPD1 (Feldspar Porphyry – Dacitic 1) is considered the first stage of intrusión (Figure 7.2-4). Textures are commonly porphyritic and sometimes phyric, with inequigranular subhedral plagioclase phenocrysts (≤1-4mm) embedded in a microcrystalline matrix with relicts of ferromagnesian minerals (amphiboles) with lots of pyrite in matrix and in veinlets; the pyrite veinlets are forming D veins of


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different stages crosscutting each other and forming pyrite stockwork. This intrusive lacks significant quartz stockwork (Cu-Au mineralization) and is locally lower grade near the contact with FPD2.

Figure 7.2-4 Intrusive Phase FPD1 (Feldspar Porphyry – Dacitic 1)

FPD2 (Feldspar Porphyry – Dacitic 2) is considered the second intrusive stage and is characterized by a porphyritic texture which is obliterated with remnant phenocrysts of plagioclases (1-3 mm) altered to clay (mainly sericite) (see Figure 7.2-5). In addition, strong stockwork (20-50%) with A, AB and B type veins (15-30 veins per metre) is consistent, with vein widths varying from <1 cm to 7 cm. In the potassic zone, the phenocrysts of plagioclases are more preserved together with primary biotite in small subhedral crystals with magnetite and calcite veinlets present in addition. Under microscope, the main components are: quartz II 15-50%; quartz I 15-43%; plagioclases 22-39%; biotite 1-20%; biotite II 10-15%; K feldspar 1-10%; sericite 10-54%, the accessory minerals are rutile 1-3%; hematitized rutile 1-3%; epidote 1%, chlorites 1%, hematite 1%, carbonates 1-5%.

Figure 7.2-5 Intrusive Phase FPD2 (Feldspar Porphyry – Dacitic 2)

FPD3 (Feldspar Porphyry – Dacitic 3); the third stage of intrusion which has at least three intrusion phases (Figure 7.2-6). Individual intrusion phases were identified by their textural features and by contact relationships. The texture of intrusions can vary from porphyritic to fine grained phyric


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texture. Principally, the intrusions are composed of subhedral plagioclase phenocrysts (30-40%) varying in size from ≤1mm to 4mm and subhedral biotite (1-10%) varying in size from ≤1mm to 3mm.

Figure 7.2-6 Intrusive Phase FPD3 (Feldspar Porphyry – Dacitic 3)

A weak stockwork of A and B quartz veins is present (1 to 8% intensity or 8 to 10 veins per metre approximately). The intensity of veins may increase slightly onto the contact with the FPD2 intrusion. Vein width can vary from <1cm to 2cm. Intrusions are characterized by potassic alteration, with moderate to strong magnetite in matrix and veins, along with chlorite and K feldspar. Phyllic alteration is characterized by quartz-sericite alteration, showing typically quartz fragments and some ductile A type veins along with late D type veins. Pyrite is common in very fine veins and matrix.

Under microscope, the main components are quartz II 10-15%; quartz I 1-37%; clays 5-51%; hematitized rutile 1-15%; sericite 7-50%; and accessory minerals: opaque 3-15%; K feldspar 1-20%; carbonates I y II 1-3%; hematite 1%; chlorites 7%, limonites 5%.

Late andesitic intrusions consisting of dykes and plugs FPA crosscut the earlier intrusions. In hand specimen, the texture is porphyritic, coarse phenocrysts of plagioclases (<1-6mm) subhedral and inequigranular plagioclases crystals, few pyroxenes/hornblendes with moderate chlorite in matrix. These late intrusions are barren and do not contain any economic mineralization.

Figure 7.2-7 Late Andesitic Intrusion


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The La Arena open pit currently in progress lies at the western margin of a FPD3-facies intrusion, where the latter forms a laccolith-like structure overlying an argillically-altered heterolithic breccia. The breccia is altered up to advanced argillic (quartz/alunite) facies, with an oxidized, porous matrix dominated by hematite, limonite and quartz. Remnant sulfides are also present.

One of the principal structural features of the project area is the La Arena anticline, the core of which hosts the mineralization-related porphyry intrusion. The strike of the anticlinal axis undergoes a deflection in the area immediately to the north of the current open pit (Figure 7.2-1). Regionally, fold axes trend generally NW-SE, but the La Arena anticline swings N-S for around 1,000 m, presumably influenced by the north-trending structures referred to previously and shown in Figure 7.1-2. This deflection, the porphyry intrusions and the mineralization are all considered to be inter-related.

Major faults within the Project area have strikes varying from NW-SE to N-S, mimicking the orientation of the fold axes and probably following the same controls. They are mainly reverse faults, probably syn-folding. Other mapped faults strike NE-SW to E-W, parallel to the main fold-related stresses, and these faults tend to be lesser structures displaying dilationary and tear movements.

In the current open pit the mineralization appears to be controlled by the interaction of three fault trends. The first corresponds broadly to the Andean Trend, NW-SE, with dips varying 50º to 70º to the NE. The second trend is N10ºE, dips sub-vertical and relative movement mainly dextral tear, and the third trend N40ºE, dips 70º to 80º to both SW and NE and has a sinistral component. The N40ºE fault trend cuts all the others, and appears to have acted as the principal feeder channels for mineralizing fluids, refer to the pit photo in Figure 7.2-8.


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Looking SW at the Calaorco Pit. The arrows are pointing out NE structures, which are filled by high grade hydrothermal breccias that are acting as feeders. The low angle argillized dykes, dashed lines, acted as mineral traps.

Figure 7.2-8 North East Trending Structures Outcropping in Calaorco Open Pit

7.3 Mineralization

The La Arena project area contains epithermal style gold mineralization in sandstone-hosted oxidized fractures and breccia, and porphyry Cu-Au (Mo) mineralization. Both styles of mineralization are probably linked because they likely emanate from the same source, namely residual magmatic activity related to an intrusive of intermediate composition.

The mineralization extends over a strike length of 2.2 km south-to-north, a width of 1.1km west-to-east and a 1,000m vertical range. Continuity of the mineralization is generally excellent, and improves with lower-grade cut-offs, which is a characteristic of this type of deposit. Further detail on mineralization is included in Section 8.

7.4 Structural Geology

The La Arena deposit lies within a regional flexion, which is characterized by the change in direction of fold axes which trend in general towards the Andean regional trend (NW-SE direction), however locally, the direction changes to a more N-S direction. This fault junction forms a dilational jog


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structure where the Cu-Au (Mo) porphyry was emplaced. To the western portion of the porphyry, lies the High Sulfidation Epithermal Au deposit hosted in sandstones of the Chimu formation. The location of this deposit is controlled by the intersection of NW-SE and NE-SW faults (Figure 7.5-1).

One of the principal structural features of the project area is the La Arena Anticline, the core of which hosts the mineralization-related porphyry intrusions. The strike of the anticlinal axis undergoes a deflection in the area immediately to the north of the current open pit (see Figure 7.2-2). Regionally, fold axes trend generally NW-SE, but the La Arena Anticline swings N-S for around 1,000 m, presumably influenced by the north-trending structures referred previously. This deflection, the porphyry intrusions (including mineralization), are all considered to be inter-related.

Four principal fault systems have been identified: the first two systems had compressive (NW-SE system) and dextral strike slip movements (N-S system), the third one represent extensional movements (NE-SW system) with normal and strike slip faults, while the fourth system has been reactivated by compressional movements (trust faults).

7.5 Hydrothermal Alteration

La Arena mineralization is related to linked deposits in the epithermal and porphyry environments, the former hosted by Chimu formation sandstones, and the latter by multiple intrusions with an age of ~25 Ma (Hedenquist; 2012).

At surface, distribution of mineral alteration clays are shown in Figure 7.5-2, where two major alteration patterns are illustrated which are dominated by, illite-pyrophyllite-muscovite and kaolinite (probably supergene) in the porphyry zone and silica-alunite-illite-dickite and supergene kaolinite in the epithermal high sulfidation zone. There are two NW oriented trends of pyrophyllite, on to the NE of Calaorco pit, and the other extending from the porphyry through Ethel pit and open to the NW; these two corridors parallel to the Andean trend were likely controlled by major structures. Conduits of hot muscovite-stable fluids have overprinted the porphyry and have cooled as they flowed to the NW along the structures (Hedenquist, 2012). In addition, there is a NE orientation of pyrophyllite in the NW end of the Calaorco pit, parallel to the major cross structure oriented to the NE.

The alteration distribution, both at surface and at depth (Figure 7.5-2 and Figure 7.5-3) is very consistent because of the strong phyllic alteration (quartz-sericite) overprints the prograde potassic alteration (secondary biotite-magnetite-k feldspar-chlorite), therefore, magnetite is completely destroyed; in addition, there is a later argillic overprint of illite-chlorite along structures deep into de porphyry. The transition from the margins of the porphyry deposit to the west, next to and within the epithermal deposit, is marked by pyrophyllite, particularly along NW structures; this is due cooling during the phyllic stage, from muscovite to pyrophyllite, with further cooling causing dickite to form. (Hedenquist, 2012).


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Figure 7.5-1 Structure and Mineralization Map - (Mineralization is plotted at 3,200 mRL.)


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Figure 7.5-2 Hydrothermal Alteration Map (at Surface) - La Arena Project


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Hydrothermal Alteration distribution in the section AA´. In the core and at depth K-alteration, envelop by phillic alteration and on top argillic alteration. It is the main characteristic from a porphyry Cu-Au deposit. To the SW (left), pervasive Advance Argilic alteration (alunite-dickite) and silica in the core, which is typically the distribution pattern of the epithermal Au-sediment hosted High Sulfidation system.

Figure 7.5-3 Hydrothermal Alteration Section - La Arena Project


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8	DEPOSIT TYPES

8.1 Introduction

The region is well endowed with mineral deposits in a variety of settings such as:

Porphyry (La Arena);
Polymetallic Au/Ag/Cu/Pb/Zn vein deposits such as Quiruvilca and Veca;
Epithermal gold, including both low and high sulfidation types, such as Lagunas Norte Mine at Lagunas Norte, Santa Rosa Mine, La Virgen Mine, La Arena Mine and the Shahuindo and Tres Cruces projects.

8.2 Deposit Types and Mineralization

La Arena hosts two types of mineralized deposits; one related to high-sulfidation epithermal Au, and the other related to porphyry environments Cu-Au (Mo). The former deposit is hosted by Chimu sandstones (Lower Cretaceous) while the latter by multiple intermediate intrusions with ages of about 24~25 Ma (Oligocene) as discussed previously in Section 7.2.

Both deposits are characterized by their typical alteration and mineralization occurrences as defined and described by Hedenquist, 1987 and Sillitoe, 2010. The epithermal deposit (currently being mined), is characterized by supergene oxidized high-sulfidation mineralization, which occurs in fractured sandstones and hydrothermal breccia zones. The porphyry deposit (located towards the east at lower elevation), is dominated by primary Cu sulfides along with Au and poor Mo.

8.3 High-Sulfidation Epithermal Au

Four separate zones of breccias containing anomalous gold have been recognized around the western and northern margins of the La Arena Porphyry. They are known as Calaorco, Ethel, Astrid and San Andrés.

Epithermal gold mineralization currently being mined in the Calaorco Open Pit occurs partly in the Calaorco Breccia (located at the contact between well-fractured Chimu quartz sandstones and the overlying intrusive), partly within the un-brecciated but still well fractured sandstones, and partly within the intrusive along the contact. Located to the north of the Calaorco Breccia and open pit, the Ethel Breccia is a similar but smaller oxidized epithermal gold deposit.

Au mineralization is both lithologically and structural controlled, and occurs principally in silicified fractured sandstones and locally in hydrothermal breccias. Structural control is mainly associated to the principle Andean orientation (NW-SE) and secondary to tensional fracturing, as well as to bedding


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planes. Tensional fracturing has acted as a principal fluid channel way, containing oxidized high sulfidation epithermal Au mineralization. Fine grained native gold is free in small proportions as electrum.

The Calaorco breccia lies parallel to the contact between the Chimu sandstones and the porphyry, with Chimu sandstone dipping gently towards the east, and the porphyry sitting on top (capping). Au Mineralization occurs within the Calaorco breccia and can be found to lie approximately ~700 m length (SE-NW) with a tendency to turn towards the north at depth. The width varies from 100 to 300 m from the contact between sandstone and porphyry. Gold mineralization is most pronounced within the oxide zone, which can extend more than 250 m depth beneath surface.

High grade zones of gold are directly controlled by the intersection of SW-NE faults, which transverse the mineralized trend and are oriented towards the NW-SE (e.g. Tilsa structure). The Tilsa structure has a strike length of approximately 300 m with a grade of 80-100 g/t Au and a variable true thickness of a few centimetres to 1 m. In this zone the Calaroco and Esperanza faults intersect, and form a high grade gold zone (≥1 g/t Au) extending towards the north up towards the Central Dyke. Beyond the Central Dyke, Au mineralization drops away slightly.

8.4 Porphyry Cu-Au (Mo) Deposits

Cu-Au mineralization is associated with phyllic (quartz-sericite) and potassic (secondary biotite -magnetite-k feldspar) alterations, which is dominated principally by pyrite, chalcopyrite, smaller amounts of bornite, covellite and chalcocite; and some molybdenite. Figure 8.4-1 displays the distribution, in section, of gold and copper values respectively.

Mineral zoning from surface downwards is typically no more than 40-50 m for the zone of secondary enrichment (cc + cv +/- copper oxides) and 10-40 m for the mixed zone (cc + cp +/- cv). The primary zone (cp +/- bn), which predominates at La Arena, is normally located at depths in excess of 100 m from surface.

The Cu-Au-(Mo) porphyry at La Arena comprises an elongated ore body 1400 m long (oriented NW-SE) by 200-400 m wide, associated with a stockwork in porphyritic andesite intrusive. Mineralization occurs as disseminations along hairline fractures as well as within larger veins. Mineralization extends down to 500 m, with the first 350 m providing the better Cu, Au and Mo grades. Sulfide mineralization consists of pyrite, chalcopyrite and molybdenite, with accessory pyrrhotite, sphalerite, galena, arsenopyrite, marcasite and rutile. In addition, very fine microscopic native gold has been observed (25 microns).

The FPD2 intrusion has the highest Cu-Au mineralization associated to phyllic (quartz-sericite) and potassic (secondary biotite, magnetite, K feldspar) with ranges from ≥0.5 to >1% Cu and ~0.5-1g/t Au respectively. Low Cu-Au mineralization is related to the intra-mineral FPD3 intrusion, which has ranges from 0.1 to <0.5% Cu and <0.2 to<0.5g/t Au.


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Cu-Au mineralization is controlled by local N-S faults and transected NW-SE Andean faults; the junction of these two systems generated a jog structure where main FDP2 and intra-mineral FPD3 have intruded.

Figure 8.4-1 Cu and Cu Distribution across the La Arena project (To the left (SW) is the epithermal gold mineralization and to the right (NE) is the porphyry Cu-Au mineralization.)


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9	EXPLORATION

9.1 La Arena Deposit

Previous exploration programs are described in Section 10.1.

In 2014, a total of 12,381 m (49 holes) of infill drilling in the Calaorco Open Pit was completed. Infill drilling (25 x 25 m grid) on the oxidized intrusive was also completed with a total of 9,058 m at 246 locations. Finally, three holes were drilled at the Astrid Deposit (648 m). The overall oxide domain drilling program for 2014 is 22,087 meters from 298 holes (see Figure 9.1-1). In addition, resource delineation drilling into the sulfide breccia Ethel zone was completed with a total of 4,487 m from 20 holes.

Figure 9.1-1 Drillhole program undertaken during 2014

9.2 Major and Regional Exploration Target

La Arena S.A. is controlling 43,140 hectares in the La Libertad region (see Figure 9.2-1). These properties include: La Arena, Charat, Cerpaquino, and Cachachi. The exploration criteria in order to stake those mining claims around La Arena were; their proximity to known mineral deposits, intersection of major lineaments, mineral-spectral anomalies (ASTER), stream sediments, air-mag anomalies, and lithological controls.


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Project generation programs were carried out at the Charat and Cachachi properties. Field reconnaissance programs were completed which included regional mapping and sampling in order to test the regional anomalies.

In addition to the La Arena project, the property includes several prospects that have been defined by a combination of soil geochemistry and exploration diamond drilling. These are Cerro Colorado, El Alizar Porphyry, Agua Blanca epithermal and porphyry occurrences, Pena Colorado and La Florida as shown in Figure 9.2-2.

Figure 9.2-1 Regional Exploration Targets - La Arena project


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Figure 9.2-2 Major Exploration Targets around the La Arena project


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10	DRILLING

10.1 Introduction

The principal methods used for exploration drilling at La Arena have been diamond core drilling (DC) and reverse circulation drilling (RC).

The deposit was relatively well drilled, with approximately 60,000 m of drilling, on a nominal spacing of 50 m in the sandstone and 65 m in the porphyry, from discovery in 1994 to 2007 with predominantly HQ and to a lesser degree NQ core.

Drilling since this time has focussed primarily on infilling the gold-oxide domain to a nominal 50 m x 25 m spacing and, to a lesser extent, infilling the sulfide domain to at least 50 m x 50m (predominantly in 2012).As of 31 December 2014, the deposit has 19,733 m of DC drilling and 114,281 m of RC drilling in the oxide domain. The sulfide domain has 121,858 m of DC drilling and 28,878 m of RC drilling.. During 2014 drilling consisted of 22,087 m (RC) into the oxide domain, and 4,487 m (DC) into the sulfide domain (targeting the sulfide breccia), as summarised in Table 10.1-1.

Table 10.1-1 Drilling Summary – La Arena Project


	Oxide Domain
		DC		RC		TOTAL
	Period	Metres	Holes	Metres	Holes	Metres	Holes
IAMGOLD	1996-2007	19,733	131	50	1	19,783	132
Rio Alto	2010-2013			92,144	857	92,144	857
Rio Alto	2014			22,087	298	22,087	298
Total		19,733	131	114,281	1,156	134,014	1,287
	Sulfide Domain
		DC		RC		TOTAL
	Period	Metres	Holes	Metres	Holes	Metres	Holes
IAMGOLD	1996-2007	36,891	199	1,136	10	38,027	209
Rio Alto	2010-2013	80,479	158	27,774	82	108,253	240
Rio Alto	2014	4,487	20	-	-	4,487	20
Total		121,858	377	28,910	92	150,768	469
	Total
		DC		RC		Total
	Period	Metres	Holes	Metres	Holes	Metres	Holes
IAMGOLD	1996-2007	56,625	330	1,186	11	57,811	341
Rio Alto	2010-2013	80,479	158	119,918	939	200,397	1,097
Rio Alto	2014	4,487	20	22,087	298	26,574	318
Total		141,591	508	143,191	1,248	284,782	1,756


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10.2 Drilling Procedures

Prior to 2007, DC holes were drilled by Sociedad Minera Cambior Peru S.A (SMCP) and RC holes were drilled by AK drilling. Most DC holes were drilled with HQ diameter until 1999 and about 40% of the holes were drilled NQ diameter from 1999 to 2005. The historical database does not clearly record core size. DC recoveries, in general, are very good, except where there are heavily oxidised zones. It is clear that in these areas there are wash outs and loss of fines from the core. RC drilling recoveries were noted as poor, during drilling, in general due to bad ground conditions and abundant underground water. There were only 11 RC holes drilled into the deposit until 2007.

The recent drilling programs commencing in 2010 were by AK drilling (RC) and Explomin del Peru (DC). DC drilling utilised HQ and to a lesser degree NQ bits with an average recovery achieved of 95%. RC drilling utilised 5¼” (133 mm) diameter bits with face sampling hammers and achieved an average recovery of approximately 90%.

10.3 Drilling Orientation

Drilling prior to 2008 was generally drilled from east to west at dips of between 60 to 70 degrees. Holes were targeted to perpendicularly intersect the expected main trend of global mineralization.

Mapping post 2008 has determined that a primary orientation of 040^ois also a major control for both gold and copper mineralisation. The majority of the infill drilling into the oxide gold domain post 2011 has been orientated orthogonal to this trend, and has likely contributed to an elevated Au grade in some areas. This has not been analysed in any detail due to the infill nature of the drilling.

The 2012 infill drilling into the sulfide domain has not had any preferential orientation. The intention of the program was to infill the drilling to a 50 m grid and drill orientation depended upon surface access and the major drilling gaps that needed to be infilled.

10.4 Surveying Procedures

10.4.1 Accuracy of Drillhole Collar Locations

Historical drillhole collars were surveyed by Eagle Mapping Ltd. using total station and differential GPS. Survey accuracy is reported as +/-0.5 m. Recent drillhole collars have been surveyed using a Total Station GPS.

10.4.2 Down-hole Surveying Procedures

Prior to the 2005 drilling campaign, holes were down-hole surveyed using acid test every 50 m. This method uses acid, in a glass test tube, with the acid etching the tube and indicating the inclination or dip of the hole. It is carried out by lowering the tube down the hole to the desired


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depth, for each reading. Magnetic azimuth readings are not obtained by this method. Also tropari survey measurements are noted in the drillhole logs. A tropari is a directional surveying instrument that gives inclination and magnetic azimuth and can be used in open holes or through rods 36 mm (1.40 inches) or larger. Accuracy to +/-0.5 degrees is claimed by the manufacturer.

After hole 172, down-hole surveys were collected with a SingleSmart Flexit tool with a reported accuracy of +/-0.2 degrees, recording both dip and azimuth. Real-time recording tools were used from 2007 onwards.

RC drilling in 2010/11 was not downhole surveyed (other than 5 RC holes) due to the lack of a non-magnetic downhole survey tool. All drilling post 2011 has been downhole surveyed with a non-magnetic down-hole Gyro tool.

Ian Dreyer considers the locations of the total data set of DC and RC holes have sufficient accuracy to make no material impact on the quality of the resource estimation, given the width and tenor of the mineralized zones encountered

10.5 Sterilization Drilling

A total of 48 RC holes were drilled between September and November 2009 to ensure planned gold oxide project infrastructure would not be placed in areas of potential economic mineralization. The holes were drilled to the south, east and north of the expected sulfide project pit limits to assess a planned waste dump to the south, planned gold oxide project infrastructure to the east and the planned gold oxide dump leach pad and ADR plant to the north.

In 2014 a small amount of sterilization drilling was completed at the proposed plant site for sulfide project. Also a small amount of drilling was completed in the La Ramada Valley should it be used in the future for any mining or civil purposes.


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11	SAMPLE PREPARATION, ANALYSES AND SECURITY

11.1 Sampling Method and Approach

11.1.1 Diamond Core Sampling

Core mark-up and sampling has been conventional and appropriate. Samples are generally 2 m long, except on geological contacts. Core has not been orientated for structural measurements, except for 18 holes drilled for the Phase II geotechnical program in 2012.

Sampling of core from exploration programs prior to 2011 consisted of chiselling the core in half. It has been noted that when the core had been split using the chisel method, the remaining half core was completely fractured, and that silicified core was generally not well split or sampled using this technique.

The standard procedure for core sampling from 2011 onwards consists of cutting the core lengthways, with a diamond saw, with half-core is sent for assay.

Diamond core samples are numbered and collected in individual plastic bags with sample tags inserted inside. Each sample batch is made up of approximately 73 samples, including 3 quality control blanks, 3 standards and 3 field duplicates. Each work order consists of a rice bag with samples along with an order list of which one copy is sent to the laboratory in Lima and another copy retained on site. Bags are closed with tie-wraps.

11.1.2 Reverse Circulation Sampling

RC samples are collected at 2 m intervals and quartered in riffle splitters. Sub-samples weigh approximately 6 kg and are collected in cloth-lined sample bags. The quality control insertion rate is identical to the DC procedure.

11.1.3 Logging

Diamond core is logged in detail for geological, structural and geotechnical information, including RQD and core recovery. Whole core is routinely photographed.

Diamond core and RC chip logging is conventional and appropriate. The DDH´s that pass through the sulfide Intrusive are being re-logged in more detail (vein density and vein type, per metre) to help to better domain the deposit. The hope is that this logging, redomaining and remodeling of the sulfides will convert previously uneconomic deeper sulfide resource into reserve, probably from an underground mining perspective.


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Core recovery has been recorded for all drillholes. Core recovery is generally 90-95% or higher and infrequently <80%. The lower recoveries occur mainly in the more weathered, upper parts of the deposit.

11.2 Sample Security

Reference material is retained and stored on site, including half-core, photographs generated by diamond drilling, duplicate pulps and residues of all submitted samples. All pulps are stored at the La Arena core shed.

11.3 Sample Preparation and Analysis

The sample preparation methods for the samples submitted prior to 2003 are not documented. Since 2003 the sample preparation methods have been constant as outlined below.

The flow sheet for drill core sample preparation and analysis is included as Figure 11.3-1. Samples are digitally weighed, dried to a maximum of 120ºC (for wet samples), crushed to 70% passing 2 mm (10 mesh), riffle split to 250 g, and pulverized to 85 % passing 75 μm (200 mesh). Furthermore, 50 g pulps are submitted for chemical analysis.

Chemical analysis at the primary laboratory (ALS Chemex since 2005) and the secondary laboratory (CIMM Peru) consisted of fire assay (FA) with atomic absorption spectrometry (AAS) finish, using 50 g sub-samples. Those samples that returned grades ≥ 5 g/t Au were analysed using gravimetric methods.

For Cu, Ag, Mo, Pb, Zn, As, Sb and Bi multi-acid (four) digestion AAS is used. Hg was analysed using cold vapour AAS. The primary laboratory has reverted to CERTIMIN (previously CIMM Peru) from 2010, with the secondary laboratory being ALS Chemex.


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Figure 11.3-1 Flow sheet for La Arena Core Sample Preparation and Analysis


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12	DATA VERIFICATION

12.1 Introduction

There was little or no routine QA/QC conducted prior to 2004 on the drillhole assays for this project. QA/QC since 2004 has generally been of a high standard. In general terms, the QA/QC in the field and in the laboratory is rigorous, and the results from the 2014 program confirm this.

12.2 Analytical Quality Control

There have been three phases of analytical quality control and quality assurance on the La Arena deposit. They are time bounded and are defined by: Pre 2004, 2004 to 2007 and 2010 onwards. There was no resource drilling completed between 2007 and September 2010 and hence no QAQC to report in this time period. The emphasis of this report is to present new data collected in 2014 as previous QA/QC results have been explained at length in previous technical reports.

12.2.1 2014 Quality Control

QA/QC recommendations outlined in the January 2014 Technical Report have been taken acted upon with respect to the increase in the frequency of control samples from around 2% in 2013 to around double this rate in 2014 (Table 12.2-1). The accuracy and precision displayed in the 2014 QA/QC data is acceptable for both RC and BH samples (from Figures 12.2-1 to 12.2-4), when examining all stages of the sample collection and laboratory analysis process. Blastholes have not been used during estimation in this resource update, although the QA/QC has been presented for completeness sake.

It is considered that the quality of the sampling in the key grade range of 0.05 g/t – 0.15 g/t Au is acceptable for resource interpretation purposes. This is being achieved by assaying down to ppb levels, rather than ppm levels. Blanks and standards show no signs of contamination or calibration/drift issues.

Table 12.2-1 Summary of Control Samples Submitted in 2014


Control Sample Type	RC	% of Total Samples Assayed
Total samples taken	6,900
Total QA/QC samples	843
Field Duplicate	216	3.1
Blank	327	4.7
Standards Submitted	300	4.3


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Figure 12.2-1 Oxide RC Field Duplicate Analysis – Au 2014


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Figure 12.2-2 Oxide RC Standard Analysis – Au - 2014


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Figure 12.2-3 Oxide Blank Analysis in RC sample stream – Au – 2014

Figure 12.2-4 Blasthole Field Duplicate Analysis – Au - 2014


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12.3 Bulk Densities

No further significant bulk density information has been collected in 2014. The bulk densities used for the resource are stated in Table 12.3-1. Those rock types that have no samples collected account for <0.01% of the total resource. Their bulk densities have been derived from similar local rock types or from first principals. The oxide intrusive measurements are clearly anomalous and have been replaced with a more appropriate value based on grab samples and first principals.

Table 12.3-1 Global Bulk Density Statistics


Lithology	Oxidation	Number of Samples	Bulk Density Average	Bulk Density in Resource Model
Sandstone	Oxide	591	2.55	2.52
Sandstone	Sulfide	-	-	2.60
Siltstone	Oxide	-	-	2.52
Siltstone	Sulfide	-	-	2.60
Slate	Oxide	-	-	2.20
Slate	Sulfide	28	2.48	2.40
Shale-Coal	Oxide	-	-	1.60
Shale-Coal	Sulfide	-	-	1.80
Intrusive	Oxide	118	2.55	2.32
Intrusive	Sulfide	1,610	2.49	2.49

12.4 Drillhole Database

The drillhole database is housed in a commercial quality Acquire database.

Hard copies of original paper drill logs, daily drill reports, core photos, assay results, and various ancillary logging features are stored on site at La Arena and are kept in good order.

The final laboratory paper assay reports are stored in Lima and a selection of 5 holes from the 2014 data set have been checked to the digital database with no errors noted.

12.5 Data Type Comparisons

La Arena has a project to date history of achieving higher grades than the initial Resource estimate and all subsequent resource estimates, prior to the January 2014 estimate. The Table 12.5-1shows the quarterly reconciliation reports for the January 2013 resource model against the actual ore mined.


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Table 12.5-1 Reconciliation of Resource Models 2013 to As-Mined Information


	As Mined (DMt-Truck)			Jan 2013 Model			Variance
Qtr/Yr	DMt ('000)	Au (g/t)	Ounces	DMt ('000)	Au (g/t)	Ounces	DMt	Au (g/t)	Ozs
2013/1	2,394	0.51	39,507	2,411	0.50	39,033	-1%	2%	1%
2013/2	3,035	0.65	63,055	2,420	0.59	46,019	20%	9%	27%
2013/3	4,996	0.58	93,542	4,586	0.47	69,684	8%	19%	26%
2013/4	4,237	0.57	78,195	3,170	0.58	59,598	25%	-2%	24%
2013T	14,662	0.58	274,299	12,586	0.53	214,334	14%	9%	22%

The reconciliation table between the financing Model from 2008 and actual ore mined has been also presented for the two periods prior to 2013 model.

Table 12.5-2 Reconciliation of Financing Models (2008) to As-Mined Data


	As Mined (DTm-Truck)			Financing Model (2008)			Variance
Qtr/Yr	DMt ('000)	Au(g/t)	Ounces	DMt ('000)	Au(g/t)	Ounces	DMt	Au (g/t)	Ozs
2012/1	1,506	1.23	59,445	2,830	0.65	58,707	-88%	47%	1%
2012/2	1,820	1.05	61,221	2,444	0.63	49,284	-34%	40%	19%
2012/3	2,369	0.59	44,737	3,136	0.45	45,564	-32%	24%	-2%
2012/4	2,572	0.63	51,725	3,786	0.43	52,093	-47%	32%	-1%
2012T	8,267	0.82	217,128	12,197	0.52	205,648	-48%	36%	5%
2011/2	551	0.54	9,538	2,529	0.39	31,798	-359%	28%	-233%
2011/3	1,227	0.63	24,948	1,227	0.59	23,162	0%	6%	7%
2011/4	1,886	1.14	69,061	2,590	0.82	68,702	-37%	28%	1%
2011T	3,664	0.88	103,547	6,346	0.61	123,662	-73%	31%	-19%

The causes of the under-estimate of grade in the 2008 model are:

The domaining style, which was based on lithology only, rather than standard industry practice of utilizing grade shells. This smooths estimates across cut-off grade boundaries.
The abundance of diamond drilling, with likely wash out of fine gold-bearing material along the heavily fractured areas of the sandstone-breccia domain.
The lack of information, due to the nominal 50m x 50m drill grid. A few larger gaps in the drill grid also exist, thus magnifying the problem locally.

A twin-hole program of DDH and RC pairs has never been completed to categorically prove whether a sample collection issue exists with DDH core. Drilling in 2014 highlighted a number of cases where the infill RC drilling returned higher grades than the nearby (+/- 10m apart) diamond drillholes.


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Given these instances, it was decided to compare RC and DDH data where composites were located within 10m of each other (Table 12.5-3) to see if a routine bias exists between RC and DDH data

The method used was to create a block model of 10m (X) x 10m (Y) x8m (Z) cells and corresponding 8m composite data, separated into RC and DDH data sets. Then a nearest neighbour interpolation was completed using a 10m (X) x 10m (Y) x 8m (Z) search with the nearest composite interpolated into the cell.

Table 12.5-3 Comparison of DDH and RC composite pairs (within 10m of each other)


Composite comparison	DDH	RC
Composite comparison	Au ppm	Au ppm
Number of composites	486	486
Minimum	0.005	0.005
Maximum	11.32	8.46
Mean	0.36	0.43
Standard Deviation	0.96	0.92
Variance	0.93	0.84
Coefficient of Variance	2.69	2.13

The Q-Q plots of models show that there is a clear systematic bias, at all grade ranges, where DDH estimates are systematically lower grade than RC estimates (Figure 12.5-1). This is likely indicative of a routine sampling bias from DDH data.

Figure 12.5-1 Q-Q Plot of RC and DDH composite pairs


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A comparison of blasthole to DDH and RC data pairs (within 10m of each other) suggests that there is a small information effect with the additional BH data (Figure 12.5-2).

Figure 12.5-2 Q-Q Plot of BH vs RC and BH vs DDH composite pairs

A final test was completed by constructing a resource model using only RC data. The reconciliation of this model to as-mined grades, over the life of the project, shows that it is a better estimator of achieved grades than models based on DDH + RC data (Table 12.5-4).

Table 12.5-4 Comparison of As-Mined Information to Resource Models with different data types


		AS-MINED (TRUCK			RESOURCE MODEL			AS-MINED/RESOURCE
		WEIGHTOMETER)			(DDH+RC)				(DDH+RC)
	Cut -	tonnes	Au	Ounces	tonnes	Au	ounces	tonnes	Au	ounces
YEAR	Off	(‘000)	(g/t)	(‘000)	(‘000)	(g/t)	(‘000)	(‘000)	(g/t)	(‘000)
2012	0.10	8,267	0.82	217.1	10,660	0.61	208.0	-29%	26%	4%
2013	0.13	14,662	0.58	274.3	12,048	0.52	202.8	18%	10%	26%
2014	0.10	7,259	0.52	121.9	6,800	0.42	92.3	6%	19%	24%
TOTAL		30,189	0.63	613.3	29,509	0.53	503.0	2%	16%	18%
		AS-MINED (TRUCK			RESOURCE MODEL (RC			AS-MINED/RESOURCE
		WEIGHTOMETER)				ONLY)		(RC ONLY)
	Cut -	Tonnes	Au	Ounces	Tonnes	Au	Ounces Tonnes		Au	Ounces
YEAR	Off	(‘000)	(g/t)	(‘000)	(‘000)	(g/t)	(‘000)	(‘000)	(g/t)	(‘000)
2012	0.10	8,267	0.82	217.1	10,505	0.70	236.0	-27%	15%	-9%
2013	0.13	14,662	0.58	274.3	12,532	0.55	221.5	15%	6%	19%
2014	0.10	7,259	0.52	121.9	6,802	0.47	103.6	6%	9%	15%
TOTAL		30,189	0.63	613.3	29,839	0.58	561.0	1%	7%	9%


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Therefore, only RC data has been used in this model update for the gold oxide domains. The background zones (outside the mineralized domains) and the sulfide intrusive use both DDH and RC data as there is insufficient RC data to generate reliable estimates with only DDH data in these areas of the resource.

12.6 Adequacy of Data

The historical data prior to 2004 has a lack of documented quality control. The new data presented is robust. In general, there are sufficient controls in place to ensure that the data collection is reliable and adequate for this resource estimate.

The approach and discipline of the QA/QC process has improved in 2014 as per the 2014 Technical Report recommendations, making for a more reliable data set for the resource estimation process.


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13	MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 Introduction

Mineralization on the La Arena property comprises an oxide deposit containing low grade gold - oxide mineralization and a contiguous sulfide deposit containing both primary and secondary copper mineralization in addition to gold.

The oxide deposit comprises sediments and oxide intrusive of which the sandstone material is currently being treated in a dump leach operation. Due to the presence of clay, the intrusive must be blended with sandstone for processing. Recent testing has been conducted to determine suitable blending ratios, from both an operating and metallurgical perspective.

The sulfide deposit is a porphyry Cu-Au type deposit. There is a recognised zone of hydrothermal alteration, in the outer zone and on the top of the argillic alteration system (kaolinite – illite), followed by phyllic alteration (quartz-sericite) and in the inner zone and on the bottom of the potassic alteration system (secondary biotite - magnetite-k feldspar), which is dominated principally by pyrite, chalcopyrite, smaller amounts of bornite, covellite and chalcocite; and some molybdenite.

13.2 Oxide Deposit

13.2.1 Mineralogy

The mineralized rock is classified in three types:

Oxidized Sandstone-breccia, hematite and goethite associated with fine-grained free Au, with a particle size from 20 to 30 microns, are present filling the matrix of the sandstone-breccia, which is a hydrothermal breccia.

Oxide Intrusive, which is on the top of the porphyry, supergene processes that have led to the formation of a leached zone extending to a depth of about 50 to 70 m from surface. The rock displays pervasive clay alteration (illite, and kaolinite), intruded by stock-work quartz veinlets associated with jarosite, hematite, goethite, chalcocite (rare), and free gold.

Colluvium, this deposit is a product of the erosion of the Calaorco Hill, where the oxidized sandstone-breccia material outcrops. Therefore, the mineralogy of this material is similar to the oxidized sandstone-breccia rock type.


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13.2.2 Leaching Performance of Sandstone Rock

The following were sourced from the historical data pertaining to the leaching of sandstone material. Both test work and plant data are presented.

13.2.3 Pre-Operations Test Program

A 2010 test work program was conducted at Heap Leaching Consulting S.A.C. (HLC) and supervised by La Arena to evaluate sandstone material from Ethel and Calaorco Pits. This program consisted of column leach tests designed to determine gold extraction and reagent usage with different particle size distribution.

Two composites were leached in columns that were 6 m high, with diameters ranging from 0.3 m to 1.2 m with 500 mg/L of NaCN and 45 days of irrigation (plus two days for additional PLS collection). Results are shown in Table 13.2-1.

Table 13.2-1 2010 Test work Program Results, Sandstone Composites


Test	Comp.	P80 mm	Head Au (g/t)	Head Cu (g/t)	CN- (mg/L)	Tailings Au (g/t)	Extraction at day 22 Au (%)	Extraction at day 47 Au (%)	Lime (kg/t)	NaC N (kg/t)
9.4.1	Calaorco	ROM	1.19	61.8	500	0.19	81.7	85.7	1.56	0.1
9.4.2	Calaorco	102	1.19	61.8	500	0.21	82.3	84.4	1.55	0.11
9.4.3	Calaorco	51	1.19	61.8	500	0.17	85.8	86.9	1.57	0.11
9.4.4	Calaorco	51	1.19	61.8	500	0.17	85.6	87	1.57	0.11
10.4.1	Ethel	ROM	0.49	36.2	500	0.03	95.4	95.6	0.84	0.08
10.4.2	Ethel	102	0.49	36.2	500	0.03	95.5	95.5	0.86	0.08
10.4.3	Ethel	51	0.49	36.2	500	0.03	95.5	95.5	0.86	0.08

13.2.3.1 Post-Operation Control Test Program

During the course of the oxide operation, the staff at the La Arena site has not typically undertaken control column leach tests. It has not been La Arena’s practice to track ore from resource or mining blocks to particular cells on the dump leach, nor to sample leach solutions to estimate recoveries from individual cells.

Two control tests have been undertaken, the first being in December 2012 and the second in August 2013. These leach tests were conducted in 0.75 m diameter and 6 m high columns on samples of sedimentary rock crushed to 100% less than 152 mm.

The test conducted in December 2012 yielded 87.4% gold leach extraction after approximately 45 days, and the one from August 2013 yielded 87% after approximately 25 days. These results are summarised in Table 13.2-2. The best indication available for the scale-up of the La Arena column


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leach test performance to dump leach recoveries is from a comparison of the gold recoveries from these two control tests with the dump leach data provided in Table 13.2-3.

Table 13.2-2 Tests Results from Column Tests during Dump Leach Operation


Test	Comp	Head Au (g/t)	Head Cu (g/t)	Leach Time d	CN- (mg/L)	Tailings Au (g/t)	Extraction. Au(%)	Lime (kg/t)	NaCN (kg/t)
12/12	Sandstone	0.78	57	45	100	0.1	87.4	0.7	0.1
08/13	Sandstone	0.64	290	25	150	No Result	87.4	1.1	0.1

13.2.3.2 Historical Performance in La Arena Sandstone Leach Pad Operation

Table 13.2-3 provides monthly data on the ore head grade, leaching duration, extraction rate, lime and sodium cyanide consumption for the oxide leach performance from the commencement of operations until December 2014.

The monthly results are consistent with the column leach test results (Table 13.2-2) on sandstone ore, which provides confirmation that the data from the site column tests can be scaled to plant operation.

Table 13.2-3 Historical Dump Leach Data


Period	Head Au (g/t)	CN- (mg/L)	Leach Time (d)	Extr. at Au (%)	Lime (kg/t)	NaCN (kg/t)
Apr 11	0.51	150	10	37.34	1.41	0.08
May 11	0.58	150	41	48.57	0.89	0.08
Jun 11	0.53	120	53	33.5	0.77	0.04
Jul 11	0.46	120	59	43.52	0.46	0.07
Aug 11	0.61	120	43	53.21	0.49	0.08
Sep 11	1.19	100	35	63.16	0.69	0.29
Oct 11	1.52	100	51	57.24	0.49	0.03
Nov 11	1.7	100	36	63.64	0.32	0.11
Dec 11	1.99	100	56	69.1	0.82	0.18
Jan 12	1.31	90 - 100	38	76.37	0.56	0.13
Feb 12	1.56	90 - 100	79	81.5	0.66	0.15
Mar 12	1.31	90 - 100	46	82.46	0.31	0.07
Apr 12	0.98	90 - 100	47	85.24	0.36	0.08
May 12	1.46	90 - 100	38	84.57	0.55	0.18
Jun 12	0.91	90 - 100	46	85.79	0.78	0.11
Jul 12	0.59	90 - 100	39	86.01	0.69	0.07
Aug 12	0.6	90 - 100	38	87.05	0.78	0.1
Sep 12	0.57	90 - 100	40	89.77	0.63	0.09
Oct 12	0.55	90 - 100	41	88.68	0.54	0.08
Nov 12	0.65	90 - 100	31	87.93	0.74	0.08
Dec 12	0.69	90 - 100	43	88.76	0.75	0.07
Jan 13	0.45	90 - 100	47	89.28	0.62	0.09


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Period	Head Au (g/t)	CN- (mg/L)	Leach Time (d)	Extr. at Au (%)	Lime (kg/t)	NaCN (kg/t)
Feb 13	0.52	90 - 100	44	88.91	0.56	0.1
Mar 13	0.58	90 - 100	60	89.55	0.74	0.12
Apr 13	0.65	90 - 100	41	89.26	0.7	0.13
May 13	0.63	90 - 100	57	87.48	0.66	0.09
Jun 13	0.67	90 - 100	43	86.42	0.83	0.08
Jul 13	0.55	90 - 100	69	84.29	0.56	0.05
Aug 13	0.59	90 - 100	43	83.74	0.8	0.07
Sep 13	0.75	90 - 100	54	83.28	0.99	0.09
Oct 13	0.98	90 - 100	44	83.4	0.72	0.14
Nov 13	0.63	90 - 100	54	83.9	0.78	0.09
Dec 13	0.46	90 - 100	44	85.5	0.88	0.1
Jan 14	0.50	100 -90	53	85.8	0.94	0.09
Feb 14	0.53	100 -90	53	86.5	0.65	0.11
Mar 14	0.55	100 -90	41	86.5	0.74	0.10
Apr 14	0.52	100 -90	48	86.3	0.81	0.09
May 14	0.48	100 -90	40	86.1	0.78	0.08
Jun 14	0.50	100 -90	70	86.3	0.84	0.08
Jul 14	0.49	100 -90	56	86.2	0.71	0.08
Aug 14	0.43	100 -90	48	86.5	0.61	0.06
Sep 14	0.49	100 -90	58	86.1	0.48	0.06
Oct 14	0.46	100 -90	62	86.0	0.59	0.07
Nov 14	0.58	130	65	85.0	0.54	0.09
Dec 14	0.58	120	57	86.9	0.47	0.13

Further details of the dump leach operation are included in Section 17.

13.2.4 Evaluation of Oxide Intrusive Leaching Properties

13.2.4.1 Test Work Description

During 2013, metallurgical test work was undertaken predominantly in La Arena’s on-site facilities, managed by CERTIMIN SA (Certimin). This test work assessed the gold oxide intrusive material’s leaching characteristics when blended with the sedimentary (sandstone). The test work focused on gold extraction, copper dissolution and cyanide consumption, but also noted solution breakthrough time (indicative of initial percolation rate).

Since all previous column leach testing on the blend of intrusive material with sedimentary rock was conducted by the “in-house” Certimin managed laboratory, during early 2014 quality assurance test work was initiated, duplicating two site column tests with two column tests conducted at SGS del Peru SAC (SGS) in Lima, to check the accuracy and repeatability of the site column tests.


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In late 2014 a pilot dump leach was conducted by the “in-house” Certimin managed laboratory. This test assessed the gold oxide intrusive material’s leaching characteristics when blended with the sedimentary (sandstone) at pilot scale dump leach.

Certimin is a certified ISO 9001:2008 company and provides laboratory services on site for mine and ADR samples, as well as offsite for exploration samples.

13.2.4.2 Sampling

Test programs were conducted on the oxide intrusive, colluvium materials and oxidized sandstone-breccia material using bulk samples from the surface. The locations of these samples are shown in Figure 13.2-1.

Figure 13.2-1 Location of Metallurgical Samples of 2013 and 2014 Programs

The 2013 test work conducted by Rio Alto used bulk samples from which composite samples were produced, having gold grades in the range 0.14 to 0.57 g/t Au; the average gold grade of the


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composites was 0.35 g/t Au. The copper grades in composites for test work were in the range 5 to 775 g/t Cu; the average copper grade was 409 g/t Cu.

The 2014 pilot dump leach conducted by Rio Alto used bulk samples from which composite samples were produced in a proportion of 2.3:1, 2.6:1, 4.2:1 and 4.8: (sandstone:intrusive), having head grades as showed in Table 13.2-4. The sample weights for these tests were 9.9 t, 24.9 t, 59.8 t and 67.6 t respectively.

Table 13.2-4 Pilot Dump Leach Head Assays


Test No.	Sandstone :	Au Assay	Ag Assay	Cu Assay
Test No.	Intrusive	(g/t)	(g/t)	(g/t)
1	2.3:1	0.46	0.21	170
2	2.6:1	0.46	0.21	170
3	4.2:1	0.30	0.62	215
4	4.8:1	0.35	0.58	147

The historical data of the current dump leach operation with sedimentary ore during 2011 and 2012 shows gold grades around 0.88 g/t, considerably higher than the future average reserve grades planned to be extracted for oxide intrusive.

13.2.4.3 Oxide Intrusive Test Work Programs

Three test work programs were completed on samples blended from oxidized sandstone-breccia material and oxide-intrusive material from the La Arena deposit. The laboratories used for the programs were:

2013 CERTIMIN S.A, La Arena Site for column test work and bottle roll tests,
SGS del Peru SAC (SGS)/Certimin, Lima for QA/QC column test work (Early 2014),
Certimin. La Arena pilot dump leach test work (Late 2014).

Test work at Certimin’s test facilities at the La Arena site was supervised by La Arena staff and concentrated on column leach gold extraction and reagent usage for different samples and blends of bulk material from the Ethel and Calaorco Pits. The column leach test reporting also noted breakthrough times. The test program consisted of 21 bottle roll leach tests and 22 column leach tests.

Subsequent test work in early 2014 at SGS in Lima and Certimin’s test facilities at the La Arena site focused on consistency and repeatability of column tests results for quality assurance (QA) purposes. The QA test program consisted of four samples split from the same composite blended sample, two tested by SGS in Lima and two tested by Certimin on site.


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The test work at Certimin’s test facilities at the La Arena site, conducted in late 2014, was also supervised by La Arena staff and concentrated on the pilot dump leach gold extraction and reagent usage for various blends of intrusive and sandstone material.

13.2.4.4 2013 Oxide Intrusive Program

The 2013 test work program was conducted in a number of phases. These consisted of both bottle roll and column leach tests to determine gold extraction and reagent usage with different rock types.

Bottle roll samples were crushed to 100% passing 1.5 mm and leached in a rolled bottle for 48 hours at pH 10 and cyanide strength ranging from 100 to 800 mg/L. As displayed in Figure 13.2-2, gold extraction ranged from 76% to 91%, sodium cyanide consumption ranged from 0.06 kg/t to 4.09 kg/t, with an average of 0.78 kg/t and lime consumption ranged from 0.6 kg/t to 2.43 kg/t. The atypically high values for sodium cyanide consumption, of 2.08 kg/t and 4.09 kg/t, correlated with high cyanide strengths of 400 and 800 mg/L, respectively. The remainder of the tests were performed with less than 200 mg/L of cyanide and the cyanide consumption ranged from 0.06 kg/t to 1.23 kg/t.

Figure 13.2-2 Bottle Roll Tests Results

Twelve column leach tests were successfully leached in open circuit, without barren solution recirculation. The following comments are based on the results presented in Table 13.2-5:

The colluvium composite gave a gold extraction of 91.5%, with reagent consumptions of 0.14 kg/t of sodium cyanide and 0.9 kg/t of lime.
The oxide intrusive from Ethel pit resulted in poor percolation performance and column leach irrigation was halted on the fourth day.


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The oxide intrusive composites 1 and 2 had low percolation rates but still resulted in gold extraction of 89.0% at 400 mg/L cyanide, with 0.92 kg/t of sodium cyanide and 2.7 kg/t of lime consumption in the successfully leached column.
Five leach column tests were performed with blend composites at different proportions of oxide intrusive and sandstone. The columns used were 2 m high by 0.2 m diameter and the samples were crushed to 100% passing 25 mm. All composite blends were successfully leached and gold extraction ranged from 88.2% to 91.2% at 100 – 150 mg/L cyanide, with total sodium cyanide consumption ranging from 0.23 kg/t to 0.35 kg/t and lime consumption from 1.6 kg/t to 2.2 kg/t.
The blended composite with 51% intrusive from Ethel Pit and 49% sandstone was leached successfully. The column used was 6.09 m high by 0.76 m diameter and the sample was crushed to 100% passing 152 mm and 100 mg/L of cyanide strength. Gold extraction was 74.1% and sodium cyanide consumption was 0.09 kg/t and lime consumption was 1.0 kg/t.
Copper head grade affected neither sodium cyanide consumption nor gold extraction in the samples tested, but no sample with more than 800 g/t of copper has been tested.
Eight column tests with composites of 100% oxide intrusive rock were not leached successfully and were stopped because the column showed percolation problems during irrigation.

Table 13.2-5 Column Tests Results


Composite	Head	Head	Tails	Extrac.	Lime	NaCN	CN	Success
Composite	Au g/t	Cu g/t	Au g/t	Au %	kg/t	kg/t	mg/L	Success
Run of Mine Material, Column 6.09 m high x 1.15 m diameter
Intrusive (Ethel Pit)	0.66	538	-	-	-	-	-	NO
Material Crushed 100% 152 mm, Column 6.09 m high x 0.76 m diameter
Intrusive (Ethel Pit)	0.66	538	-	-	-	-	-	NO
51% Intrusive (Ethel Pit)
	0.36	289	0.1	74.1	1	0.09	100	YES
/49%ST
Material Crushed 100% passing 76 mm, Column 1.98 m high x 0.20 m diameter
Intrusive (Ethel Pit)	0.66	538	-	34.8	1.4	0.05	100	NO
Intrusive (Ethel Pit)	0.66	538	-	28.7	1.4	0.02	100	NO
Intrusive (Ethel Pit)	0.66	538	-	41.4	1.4	0.04	100	NO
Intrusive (Ethel Pit)	0.66	538	-	45.6	1.4	0.05	100	NO
Intrusive (Ethel Pit)	0.66	538	-	49.9	1.5	0.06	100	NO
Material Crushed 100% passing 102 mm, Column 6.09 m high x 0.30 m diameter
Colluvium	0.14	5	0.01	91.5	0.9	0.13	150	YES
Material Crushed 100% passing 38 mm, Column 2.00 m high x 0.15 m diameter
Intrusive Composite 1, 2	0.46	562	0.08	77.6	2.7	0.41	200	NO
Intrusive Composite 1, 2	0.46	562	0.05	89	2.7	0.92	400	YES
Intrusive Composite 3	0.44	769	0.06	87.6	2.7	0.42	200	YES
Intrusive Composite 3	0.44	769	0.05	88.6	2.7	0.76	400	YES
Intrusive Composite 4	0.57	775	0.06	89.1	2.7	0.41	200	YES


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Composite	Head	Head	Tails	Extrac.	Lime	NaCN	CN	Success
Composite	Au g/t	Cu g/t	Au g/t	Au %	kg/t	kg/t	mg/L	Success
Intrusive Composite 4	0.57	775	0.05	91.2	2.7	0.80	400	YES
Material Crushed 100% passing 25 mm, Column 2.00 m high x 0.20 m diameter
20% Intrusive (Calaorco)/80%ST	0.24	182	0.03	88.2	1.6	0.12	100	YES
30% Intrusive (Calaorco)/70%ST	0.28	319	0.03	88.2	1.9	0.13	100	YES
40% Intrusive (Calaorco)/60%ST	0.33	620	0.04	87.4	2.2	0.16	100	YES
30% Intrusive (Ethel Pit)/70%ST	0.27	279	0.02	91.2	1.9	0.19	150	YES
40% Intrusive (Ethel Pit)/60%ST	0.33	325	0.03	90.4	2.1	0.18	150	YES
ST=Sandstone

13.2.4.5 Early 2014 Oxide Intrusive Program

The quality control test work was conducted in two column tests at the Certimin Site Laboratory, with replication of these tests conducted at SGS Lima. The column leach tests were designed to determine repeatability of gold extraction and reagent usage, with each composite sample split and sent to both laboratories.

The column test samples were crushed to 100% passing 25 mm and leached in a column 2.0 m high by 0.15 m diameter. The sample was 33.3% oxide intrusive rock and 66.7% sandstone. It was intended that both 2014 Certimin and SGS tests were undertaken using 150 mg/L cyanide solution. However, SGS misunderstood this requirement to be 150 mg/L sodium cyanide, equivalent to 80 mg/L cyanide. Column leach tests results are presented in Table 13.2-6 and the following comments summarise the outcomes:

There were no percolation problems for either the columns at SGS or at the Certimin site laboratory.
Despite the different cyanide addition rates, the test results did not demonstrate any material difference in gold extraction.
Cyanide consumption did not appear to be a function of gold extraction rate, however it did appear to be related to the initial strength of the cyanide solution.
Final gold extraction was similar for all four columns tested, ranging from 86.4% to 87.1%.
Lime consumption ranged from 1.53 kg/t to 1.56 kg/t.
Sodium cyanide consumption was low in the range of 0.10 kg/t (80 mg/L cyanide) to 0.16 kg/t (150 mg/L cyanide). Higher cyanide strengths resulted in higher sodium cyanide consumption.
The kinetic curves of Figure 13.2-3 and Figure 13.2-4 showed similar behaviour for the extraction of both gold and copper.


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The percolation of the column test at site was slower than the column tests at SGS, taking approximately 2 days longer to break through (Figure 13.2-3 and Figure 13.2-4).

Table 13.2-6 Certimin and SGS Column Tests Results


Composite	Head	Head	CN-	Tails	Extrac.	Lime	NaCN
Composite	Au g/t	Cu g/t	mg/L	Au g/t	Au %	kg/t	kg/t
Site Column C-18	0.44	339	150	0.06	86.2	1.56	0.15
Site Column C-19	0.44	339	150	0.05	87.1	1.57	0.17
SGS Column 01	0.49	349	80	0.07	86.5	1.53	0.10
SGS Column 02	0.49	349	80	0.07	86.4	1.53	0.10

Figure 13.2-3 Gold Extraction Curve Kinetics for Column Tests


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Figure 13.2-4 Copper Extraction Curve Kinetics for Column Tests

13.2.4.6 Late 2014 Oxide Intrusive program

Due to the percolation problems in the dump leach test for the 2:1 blend of intrusive and sandstone, Rio Alto conducted additional tests to explore alternative blending ratios. Pilot dump leach pads at ratios 2.3:1, 2.6:1, 4.2:1 and 4.8:1 were constructed for testing.

Four pilot dump leach tests were performed on the composite material blended at different proportions: 2.3:1, 2.6:1, 4.2:1 and 4.8:1 named dump leach test No. 1, 2, 3 and 4, respectively (Table 13.2-7). Industrial dump leach conditions were applied to all pilot dump leach tests. The irrigation solution was 100 mg/L CN^-and pH 11. A front end loader was used to mix and homogenize blended ROM material and the irrigation time was in a range of 51 to 60 days. All dump leach tests were 8 metres high. The following were observed during the test:

Dump leach tests No. 1 and 2 had percolations problems but completed the whole irrigation cycle of 54 days. Irrigation rates were decreased to 5.9 and 6.9 L/m²/h due to the low percolation rates.
Dump leach tests No. 3 and 4 did not show any percolation problems and completed irrigation cycles of 51 and 60 days, respectively. Irrigation rates were kept at 10 L/m²/h.
Gold extraction after 51 to 60 days of leaching was 80.3% to 83.6%.
Sodium cyanide consumption ranged from 0.09 to 0.10 kg/t for all tests.
Lime consumption ranged from 0.9 to 1.0 kg/t. The irrigation solution was pH 11 dropping to pH 9.8 to 11 after passing through the dump.
The percolation rate in dump leach test No. 4 was moderately slow, taking approximately three days to break through. Tests No. 1, 2 and 3 broke through in one day.


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The copper tenors of the pregnant solution were 8.3, 25, 27 and 37.2 mg/L for dump leach tests No. 1, 2, 3 and 4, respectively.
Figure 13.2-5 shows that after the tested leach periods, gold extraction had not yet plateaued, suggesting that additional leaching time would be beneficial for increased gold recovery.

Table 13.2-7 Late 2014 Pilot Dump Leach Results


Test No.	Sandstone : Intrusive	Irrigation Rate	Irrigation Time (d)	Head	Head	CN-	Tails	Final Extrac.	Extrac. 51 d	Lime	NaCN
Test No.	Sandstone : Intrusive	Irrigation Rate	Irrigation Time (d)	Au g/t	Cu g/t	mg/L	Au g/t	Au%	Au%	kg/t	kg/t
1	2.3:1	5.9	55	0.46	170	100	0.095	80.3	79.6	0.91	0.09
2	2.6:1	6.9	55	0.46	170	100	0.086	82.0	81.1	1.00	0.10
3	4.2:1	10	51	0.30	215	100	0.064	80.3	80.3	0.90	0.09
4	4.8:1	10	60	0.35	147	100	0.058	83.6	78.6	0.90	0.10

Figure 13.2-5 Kinetic curve for gold extraction in pilot dump leach test

13.2.5Dump Leach Results for Economic Modelling

Pilot dump leach test results on the various blend composites with differing proportions of intrusive material indicate that between 79.6% and 81.1% of the gold extracted in the dump leach after 51 days of irrigation reports to the adsorption plant.

In order to predict intrusive gold leach extraction, the sandstone gold extraction was compared to the rates achieved for the composite blends to determine an average intrusive gold extraction rate. This recovery is valid for a blend proportion ranging from 4:1 to 3:1. Blend ratios lower than this showed evidence of poor percolation and are unlikely to achieve adequate leaching performance.


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No correction factor was introduced to allow for any change in intrusive gold extraction for full scale operation because pilot dump leach tests were carried out in cells of the same dimensions of the industrial dump leach.

Determination of a gold extraction model for the sandstone material was developed from an analysis of the historical full scale recovery data as summarized in Table 13.2-3. The data used was from 2012 and 2013 years only, after ramp up of the leaching operation and showing steady stage operation.

The mathematical model that has been developed to describe gold leach extraction in sandstone as a function of head grade is a simple cubic best fit equation as follows:

Gold recovery (%) = 22.126 x (gold head grade)3 minus 61.477 x (gold head grade)2 plus 45.372 x (gold head grade) plus 77.377.

This equation is applicable for gold head grades between 0.29 g/t and 1.56 g/t. The fit of the equation is illustrated in Figure 13.2-6.

Figure 13.2-6 Sandstone gold extraction model

Final gold recovery has been estimated as 1% less than dump leach gold extraction due to minor gold losses during pregnant liquor transport, carbon adsorption, carbon losses during transport to desorption, fine carbon losses due to attritioning and gold losses in smelt slag.

Lime and sodium cyanide consumptions were calculated for intrusive rock in a similar manner to the calculation of the gold extraction rate. The bases for this calculation were the historical data from the industrial dump leach and the pilot dump leach tests results for the blended material at


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different proportions. As showed in Table 13.2-8 the average lime and cyanide consumption are 1.78 kg/t and 0.062 kg/t, respectively for intrusive material only when it is blended is a proportion rate between 3:1 and 4:1 (sandstone : intrusive).

Typical lime and cyanide consumptions are 0.675 kg/t and 0.098 kg/t, respectively, for sandstone leaching and lime and cyanide consumption are 0.928 kg/t and 0.090 kg/t, respectively, for the blend leaching.

Table 13.2-8 Reagents Consumption Calculated for Each Rock Type


		Blend NaCN	Calc.	Sandstone	Blend	Calc.	Sandstone
		Cons. at 51	Intrusive	NaCN Cons.	Lime	Intrusive	Lime
Test	Blend	d (kg/t)	NaCN Cons.	(kg/t)	Cons.	Lime	Cons.
			(kg/t)		(kg/t)	Cons.	(kg/t)
						(kg/t)
1	2.3:1	0.090	0.071	0.098	0.910	1.455	0.675
2	2.6:1	0.090	0.069	0.098	1.000	1.842	0.675
3	4.2:1	0.090	0.057	0.098	0.900	1.834	0.675
4	4.8:1	0.090	0.052	0.098	0.900	1.976	0.675
Avg.		0.090	0.062	0.098	0.928	1.777	0.675

13.3 Sulfide Deposit

13.3.1Historical Test Programs

13.3.1.1 Stage 1 Testing

Stage 1 metallurgical test work on La Arena rock samples was managed by Cambior Inc. This was carried out by SGS Lakefield, Canada, in December 2006. The main elements of this work were.

Three composites were prepared and designed as: mixed rock, primary average grade and primary high grade composites. Information on the construction of these composites is not available for this test work program.
Copper head grades varied between 0.55% to 0.99% and gold head grades varied between 0.31 g/t to 0.73 g/t.
Physical testing included: Bond rod and ball mill work indices.
Metallurgical testing included: QEMSCAN, reflected light microscopy, rougher flotation tests for optimization, batch flotation tests for optimization, locked cycle tests and cyanide leaching on different flotation streams.
Bond ball mill work indices (BBWi) ranged from 6.9 to 8.1 kWh/t for the samples tested and the Bond rod mill work index (BRWi) was 7.1 kWh/t for the sample tested.
After reagents scheme and grind size optimization work, the flotation conditions selected were as follows:


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	o	primary grind size P80 of 90 µm,
	o	regrind size P80 of 35 µm,
	o	27.5 g/t of Aero 5100, 10 g/t of stove oil, 2.5 g/t pine oil,
	o	a natural pH in the rougher flotation and a pH of 11.5 in the cleaners.

The samples tested were characterised as having a high mass pull to rougher concentrate. Therefore a cleaner stage before regrinding was introduced to reduce the mass pull to the rougher concentrate, followed by three cleaner stages.
Locked cycle test results were 82.3% to 84.3% of copper recovery with a copper concentrate grade around 28.5% to 30.4%. Gold recovery ranged from 34.2% to 35.9% with a gold concentrate grade of 7.2 to 8.3 g/t.
Head grades of the samples used for locked cycle test were 0.5% copper and 0.3 g/t gold.

13.3.1.2 Stage 2 Testing

Stage 2 metallurgical test work was managed by Cambior Inc. This work was carried out by SGS Lakefield, in Canada in February 2007. The main elements were:

Thirty composites were prepared and designed as: variability composites from 1 to 30, master, north, south and south deep.
Copper grades varied between 0.37% to 0.79% and gold grades varied between 0.16 g/t to 0.83 g/t.
Physical testing included: Bond rod and ball mill work indices, abrasion index and SAG power index (SPI).
BBWi ranged from 5.2 to 9.7 kWh/t, BRWi ranged from 5.7 to 7.5 kWh/t, the Bond abrasion indices ranged from 0.037 to 0.093 and the SPI ranged from 7.5 to 18.4 minutes.
Metallurgical testing included variability testwork, using optimized conditions developed during Stage 1 Testing. All tests were open circuit batch cleaner tests.
Flotation standard conditions were:


	o	a primary grind size P80 of 90 µm,
	o	a regrind size P80 of 35 µm,
	o	27.5 g/t of Aero 5100, 10 g/t of stove oil, 2.5 g/t pine oil,
	o	a natural pH in the rougher flotation and a pH of 11.5 in the cleaners.

Copper recovery from the batch flotation of the variability samples ranged from 91.3% to 46.4% with an average of 84.9%.
Copper grade in final concentrate ranged from 6.77% to 30.9%, with an average of 20.8%.
Gold recovery ranged from 20.5% to 59.6% with an average of 40.1%.


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13.3.1.3 Stage 3 Testing

The stage 3 metallurgical test work on La Arena rock samples was managed by MQes Inc. This was carried out by ALS Metallurgy Kamloops, Canada in August 2012. The main elements of this work were:

Twenty seven composites were designed and prepared based on rock types, zones of the deposit and copper grade.
Copper grades varied between 0.13% and 0.44% and gold grades varied between 0.07 g/t and 0.26.
Metallurgical testing included: quantitative mineralogy, rougher flotation tests, batch cleaner flotation tests, locked cycles tests, variability batch cleaner test and cyanide leaching on different flotation streams.
After reagents scheme, pH and grind size optimization work, the flotation conditions selected were as follows:


	o	primary grind size P80 of 75 µm,
	o	regrind size P80 of 25 µm,
	o	6 g/t of Aerophine 3418A in rougher flotation and 34 g/t in cleaner flotation,
	o	15 g/t of sodium cyanide in cleaner flotation,
	o	A pH of 10 in the roughers and 11.5 in the cleaners,
	o	A rougher flotation residence time of 8 min and cleaner residence time of 22 min.

Copper recovery from the batch flotation of the variability samples ranged from 24.5% to 90.5% with an average of 63.4%.
Copper grade from the batch flotation in final concentrate ranged from 7.4% to 32.7%, with an average of 25.0%.
Gold recovery from the batch flotation ranged from 5.5% to 46.1% with an average of 21.9%.
Copper recovery from the locked cycle tests ranged from 72.3% to 93.7% with an average of 83.8%
Copper concentrate grade from the locked cycle tests ranged from 20.7% to 30.5%, with an average of 26.6%
Gold recovery from the locked cycle tests ranged from 28.5% to 74.7% with an average of 45.4%.
The head grade of the samples used for the locked cycle tests ranged from 0.12% to 0.47% copper.


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13.3.1.4 Stage 4 Testing

Stage 4 metallurgical test work on La Arena rock samples was managed by MQes Inc. This was carried out by ALS Metallurgy in Kamloops, Canada during April 2013. The main elements of this work were:

Forty four composites were designed and prepared considering rock types, zone of the deposit and copper grade.
Copper grades varied between 0.09% and 0.85% and gold grades varied between 0.04 g/t and 0.84 g/t.
Physical testing included: Bond ball mill work index, abrasion index, JK Drop Weight test and SAG mill comminution (SMC) tests.
BBWi ranged from 7.0 to 13.7 kWh/t with an average of 9.7 kWh/t; Bond abrasion indices ranged from 0.001 to 0.074, with an average of 0.022; and the Axb ranged from 62.7 to 328.2, with a 50th percentile value of 197.
Metallurgical testing included: quantitative mineralogy, variability flotation tests, locked cycles tests, rheology on tailings and flocculant screening for concentrate and tailings.
All flotation tests were conducted at the optimized conditions developed during the stage 3 testing.
Copper recovery from the batch flotation of the variability samples ranged from 8.6% to 88.9% with an average of 69.2%.
Copper grade from the batch flotation in final concentrate ranged from 1.0% to 42.1%, with an average of 27.6%
Gold recovery from the batch flotation ranged from 2.6% to 54.3% with an average of 26.3%
Copper recovery from the locked cycle tests ranged from 64.7% to 94.2 % with an average of 83.8%
Copper concentrate grade from the locked cycle tests ranged from 22.1% to 32.9%, with an average of 27.2%
Gold recovery from the locked cycle tests ranged from 16.9% to 51.6% with an average of 35.0%.
The head grade of the samples used for the locked cycle tests ranged from 0.11% to 0.48% copper.

13.3.1.5 Stage 5 Testing

Stage 5 metallurgical test work on La Arena rock samples was managed by Mr Pedro Martinez, a consultant to Rio Alto. This work was carried out by ALS Metallurgy in Kamloops, Canada, during May 2013. The main elements of this work were:


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One composite sourced from stage 4 test work labelled as: M1A Centre HA High.
The copper grade was 0.54% and the gold grade 0.23 g/t.
Metallurgical testing include: rougher flotation tests for optimization, cleaner batch flotation tests for optimization.
After reagents scheme, pH and grind size optimization work, the flotation conditions selected were as follows:


	o	primary grind size P80 of 150 µm,
	o	regrind size P80 of 25 µm,
	o	40 g/t of Aero 5100 and 40 g/t of Aero 3302 in rougher flotation,
	o	26 g/t of Aero 5100 in cleaner flotation,
	o	A pH of 11.5 in both roughers and cleaners,
	o	A rougher flotation residence time of 12 min and cleaner flotation residence time of 22 min.

The cleaner batch tests achieved copper recoveries ranging from 76.9% to 82.7%, with a copper concentrate grade ranging from 29.7% to 30.2% and a gold recovery ranging from 15% to 25.8%.

13.3.1.6 Stage 6 Test Work Description

This section describes the results from the latest round of metallurgical test work, referred to as Stage 6, undertaken predominantly in ALS Metallurgy Kamloops – Canada during 2013/2014. This test work assessed the copper sulfide porphyry material’s comminution and flotation characteristics. The work focused on copper recovery, copper concentrate grade and optimum parameters for processing.

13.3.1.7 Mineralogy

The mineralized rock is classified into three types:

Argillic Alteration: The alteration assemblage is kaolinite – illite, the sulfide minerals are pyrite – chalcopyrite – chalcocite – covelite.

Phyllic Alteration:The alteration assemblage is sericite – quartz, the sulfide minerals are pyrite – chalcopyrite - molybdenite.

Potassic Alteration:The alteration assemblage is secondary biotite – K feldespars – magnetite, the sulfide minerals are chalcopyrite – bornite - molybdenite.


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13.3.1.8 Metallurgical Sampling

The test work program at ALS Metallurgy – Kamloops was performed on the following composites summarized in Table 13.2-3.

Table 13.3-1 Stage 6 Test Work Composites


Composite	Alteration	Test	Composite	Alteration	Test
Name	Type	Test	Name	Type	Test
ARC	Argillic	Flotation	VC-06	Phyllic	SMC / Flotation
PHC	Phyllic	Flotation	VC-07	Phyllic	SMC / Flotation
KC	Potassic	Flotation	VC-08	Phyllic	SMC / Flotation
ARC-DW	Argillic	Drop Weight / SMC	VC-09	Phyllic	SMC / Flotation
PHC-DW	Phyllic	Drop Weight / SMC	VC-10	Phyllic	SMC / Flotation
KC-DW	Potassic	Drop Weight / SMC	VC-11	Potassic	SMC / Flotation
VC-01	Argillic	SMC / Flotation	VC-12	Argillic	SMC / Flotation
VC-02	Argillic	SMC / Flotation	VC-13	Phyllic	SMC / Flotation
VC-03	Argillic	SMC / Flotation	VC-14	Argillic	SMC / Flotation
VC-04	Argillic	SMC / Flotation	VC-15	Potassic	Flotation
VC-05	Phyllic	SMC / Flotation

The 6 composites comprised ARC, PHC, KC, ARC-DW, PHC-DW and KC-DW and the variability test work samples comprised VC-01, to VC-15. Figure 13.3-1 illustrates the metallurgical drill hole location in North pit shell.


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Figure 13.3-1 Metallurgical drill hole location superimposed on the North pit shell

Details of the drill intercepts used for composite construction can be found in Table 13.3-2.

Table 13.3-2 Stage 6 Testing Detailed Composites Construction



Composite	DDH	From (m)	To (m)	Composite	DDH	From (m)	To (m)

ARC	LA-D12-M1A	90	92	KC-DW	LA-D12-M4D	386	388
	LA-D12-M1A	154	156		LA-D12-M4D	393.4	394.5
	LA-D12-M1A	164	166		LA-D12-M4D	410	412.8
	LA-D12-M1A	268	270		LA-D12-M4D	472	474
	LA-D12-M1A	370	372		LA-D12-M4D	481	482.5
	LA-D12-M1A	406	408		LA-D12-M4D	484	486
	LA-D12-M1A	454	456	VC-01	LA-D12-M1A	330	334
	LA-D12-M1A	462	464	VC-01	LA-D12-M1A	338	348
	LA-D12-M16	324	328	VC-02	LA-D12-M16	200	214
	LA-D13-M01	100	102	VC-03	LA-D13-M01	48	57.7


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Composite	DDH	From (m)	To (m)	Composite	DDH	From (m)	To (m)

	LA-D13-M02	190	192	VC-04	LA-D13-M01	115	130
PHC	LA-D12-M16	246	248	VC-05	LA-D12-M4D	336	338
	LA-D12-M16	302	304		LA-D12-M4D	340	344.2
	LA-D12-M4D	286	290		LA-D12-M4D	346	349.9
	LA-D12-M4D	312	314		LA-D12-M4D	352	354
	LA-D13-M02	102	112	VC-06	LA-D12-M16	254	262
	LA-D13-M02	162	166		LA-D12-M16	264	268
KC	LA-D12-M4D	380	384		LA-D12-M16	270	272
	LA-D12-M4D	389.15	392	VC-07	LA-D12-M17	104	118
	LA-D12-M4D	394.5	396	VC-08	LA-D13-M01	74	87.3
	LA-D12-M4D	402.5	406	VC-09	LA-D13-M02	78	92
	LA-D12-M4D	412.8	414	VC-10	LA-D13-M02	138	152
	LA-D12-M4D	426	249.5	VC-11	LA-D12-M4D	447	452
	LA-D12-M4D	431	432		LA-D12-M4D	454	460
	LA-D12-M4D	434	436		LA-D12-M4D	462	466.5
	LA-D12-M4D	440	442	VC-12	LA-D12-M16	170	186
	LA-D12-M4D	486	488	VC-13	LA-D12-M4D	248	254
ARC-DW	LA-D12-M1A	88	90	VC-13	LA-D12-M4D	256	260
	LA-D12-M1A	196	198	VC-14	LA-D13-M02	168	180
	LA-D13-M01	98	100	VC-15	LA-D12-M4D	302	304
	LA-D13-M01	152	158		LA-D12-M4D	384	386
PHC-DW	LA-D12-M16	248	250		LA-D12-M4D	392	393.4
	LA-D12-M16	300	302		LA-D12-M4D	406	410
	LA-D13-M02	98	102		LA-D12-M4D	414	417
	LA-D13-M02	112	114		LA-D12-M4D	469.5	470.5
	LA-D13-M02	166	168		LA-D12-M4D	474	481
					LA-D12-M4D	482.5	484
					LA-D12-M4D	488	490

Domain Composites

The composites (ARC, PHC, KC, ARC-DW, PHC-DW and KC-DW) were selected to spatially represent the distribution, ore grade, and type of alteration in the ore zones.


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ARC composite was prepared from 11 intervals of 4 drill holes (LA-D12-M1A, LA-D12-M16, LA-D13-M01 and LA-D13-M02).

PHC composite was prepared from 6 intervals of 3 drill holes (LA-D12-M16, LA-D12-M4D and LA-D13-M02).

KC composite was prepared from 10 intervals of 1 drill holes (LA-D12-M4D).

ARC-DW composite was prepared from 4 intervals of 2 drill holes (LA-D12-M1A and LA-D13-M01).

PHC-DW composite was prepared from 5 intervals of 2 drill holes (LA-D12-M16 and LA-D13-M02).

KC-DW composite was prepared from 6 intervals of 1 drill holes (LA-D12-M4D).

Table 13.3-3 shows the head assays of the composites.

Table 13.3-3 Composite Head Assays


Element	Units	ARC	PHC	KC	ARC-DW	PHC-DW	KC-DW
Cu	%	0.28	0.27	0.36	0.14	0.31	0.36
Au	g/t	0.15	0.18	0.18	--	--	--
Fe	%	6.15	6.75	5.45	--	--	--
S	%	6.40	6.72	3.76	5.96	7.05	3.31

Variability Composites

15 Variability composites were constructed from samples selected from across the zone of mineralization to be representative of each alteration in a copper grade range. This selection was made to test the metallurgical response of these samples, in order to develop head grade recovery correlations for the mine block model.

13.3.1.9 Comminution Test Program

JK Drop Weight/SMC and abrasion tests were conducted on the ARC-DW, PHC-DW and KC-DW composites. Bond rod and ball mill work index tests were conducted on ARC, PHC and KC composites. SMC, abrasion, Bond rod and ball mill work index tests were conducted on all variability composites. The Axb values for argillic and phyllic alterations samples ranged from 90 to 304, which is classified as soft to very soft material. The percentile 50^thvalue was 158 for argillic and phyllic samples. The potassic alteration samples Axb ranged from 30 to 32, which is classified as competent material.


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Bond rod and ball mill work indices for the argillic and phyllic indicate that low energy levels will be required for the comminution circuit. The potassic ores had a significantly higher Bond rod work index and Bond ball mill work index compared to both argillic and phyllic composite samples. The BRWi for the argillic and phyllic domain composites and variability composites ranged from 5.9 kWh/t to 10.5 kWh/t. Similarly the BBWi for the argillic and phyllic domain composites and variability composites ranged from 7.3 kWh/t to 12.7 kWh /t.

The Bond abrasion index also shows that the argillic and phyllic alterations have low values i.e. <0.1, while the potassic alteration showed an average abrasion index of 0.1.

13.3.1.10 Flotation Test Program

Flotation optimization studies were performed on ARC, PHC and KC composites. The optimized conditions were applied to the variability composites. In addition, copper flotation tailings were produced for dynamic thickening test and tailings rheology.

The scope and objectives of this program included:

Development of optimal process conditions on the ARC and PHC composites for the design criteria;
Production of bulk flotation tailings for dynamic thickening and rheology tests;
Test of variability samples of the rock types to obtain metallurgical response information by rock type and to develop head grade recovery correlation for the mine block model.

Rougher Flotation Density

High rougher mass recovery in rougher concentrate was identified in previous flotation programs (stages 4 and 5) with around 30% mass pull to rougher concentrate. In order to reduce mass in rougher concentrate a number of dispersant reagents were tested. However, it was found that a slurry density of 20% w/w gave the best results in terms of rougher concentrate mass reduction and metal recoveries. Rougher flotation optimization testing indicated that the best rougher flotation conditions were:

Primary grind size of 106 µm
pH of 10.5
Hostaflot 3403 as the primary copper collector
Aerophine 3418A as the secondary collector
Residence time of 15 minutes

Aged Sample Evaluation


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The above rougher flotation conditions were used on a series of tests conducted to determine whether there was any effect of the age of samples on flotation performance. The tests were conducted on samples of argillic and phyllic ores that were stored under ambient conditions for 36 days prior to flotation. The test results indicate that there is negligible impact on copper flotation recovery or grade from aging of the rock.

Variability Tests

Cleaner flotation tests were carried out on each of the fifteen variability composite samples at the optimum conditions of; grind size P80 of 106 µm, pH 10.5, regrind size P80 of 25 µm, Hostaflot 3403 as primary collector and Aerophine 3418A as secondary collector

Table 13.3-4 summarises flotation results for each of the variability composites.

Table 13.3-4 Cleaner Flotation Tests Results on Variability Composites


		Copper			Gold
Comp.	Alteration	Head	Recover	Con. (%)	Head	Recover	Con.
		(%)	y (%)	Con. (%)	(%)	y (%)	(g/t)
VC-01	Argillic	0.29	83.3	30.3	0.17	25.3	5.60
VC-02	Argillic	0.12	68.6	29.1	0.11	24.0	10.0
VC-03	Argillic	0.65	88.1	35.3	0.55	56.2	19.1
VC-04	Argillic	0.47	88.3	29.6	0.36	26.7	6.84
VC-12	Argillic	0.09	69.4	16.3	0.10	26.9	7.50
VC-14	Argillic	0.68	88.5	30.0	0.51	52.9	13.6
VC-05	Phyllic	0.31	72.0	30.1	0.15	37.0	7.59
VC-06	Phyllic	0.37	79.0	24.5	0.35	36.5	10.8
VC-07	Phyllic	0.80	89.9	28.9	0.63	30.7	7.78
VC-08	Phyllic	0.71	93.2	30.9	0.52	45.6	11.1
VC-09	Phyllic	0.31	79.2	24.4	0.21	33.3	6.99
VC-10	Phyllic	0.50	80.7	30.5	0.38	54.6	15.7
VC-13	Phyllic	0.27	67.2	31.3	0.11	19.2	3.55
VC-11	Potassic	0.39	85.6	31.0	0.29	47.9	13.2
VC-15	Potassic	0.37	79.7	30.6	0.20	44.2	9.18

Figure 13.3-2 shows copper recovery against copper head grades for all variability tests. This figure shows a consistent trend of increased copper recovery with copper head grade for all rock types. A trend line has been included for phyllic alteration (PH) and argillic alteration (AR) as well as cleaner test results from KM3526 test program. For the potassic alteration (K) no trend was included due to limited data available for this type of alteration.


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Figure 13.3-2 Comparative Copper Recovery Results in Variability Composites

Figure 13.3-2 shows that copper recoveries for argillic alteration samples are above those from previous test work, while the phyllic and potassic alteration samples performed similarly to the results obtained in KM3526.

Figure 13.3-3 showed no clear tendency between gold recovery and gold head grade for the three composites. Similar results were observed for the previous test program KM3526.

Figure 13.3-3 Gold Recovery Results Vs Gold Head Grade for All Variability Composites

Locked Cycle Tests


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Locked cycle flotation tests (Figure 13.3-4) using optimized conditions were carried out on each of the alteration type composites. Each locked cycle test consists of five cycles.

Figure 13.3-4 Locked Cycle Tests Flow Sheet

Figure 13.3-5 and Figure 13.3-6 illustrate locked cycle test results for all three composites. The copper recovery model currently used in the mine plan has also been included in Figure 13.3-5. The current copper recovery model represents previous and current test work results. Copper recovery model is described as follows:

If the copper head grade is less than 0.08%, then copper recovery is 0%.
If the copper head grade is greater than 0.45%, then copper recovery is 92.9%.
If the copper head grade is between 0.08% and 0.45% than copper recovery is 9.4547 x ln (%copper head grade) plus 99.79%.


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Figure 13.3-5 Locked Cycle Test Cu Head Grade Vs. Cu Recovery

Figure 13.3-6 shows gold head grades against gold recoveries for all three composites. The potassic alteration achieves better gold recoveries compared to the ARC and PHC composites. Gold recoveries for ARC and PHC samples are similar to those obtained in the KM3526 test work program.

Figure 13.3-6 Locked Cycle Test Au Head Grade versus. Au Recovery

The main differences between this latest round of test work KM3991 and the results from the previous test work programs is that this work has simplified the flow sheet (no rougher cleaner stage) with feed dilution to 20%, used a coarser grind size (106 vs. 75 µHi m), lowered rougher


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and cleaner pH (10.5), and uses less expensive reagents in the rougher circuit (Hostafloat 3403 vs. Aerophine 3418A).

Figure 13.3-7 illustrates variability test results for argillic and phyllic rock types. The gold recovery model has also been included in the figure and is described as follows:

If the gold head grade is less than 0.10 g/t, then gold recovery is 0%.
If the gold head grade is greater than 0.55 g/t, then gold recovery is 45.5%.
If the gold head grade is between 0.10 g/t and 0.55 g/t then gold recovery is 12.862 x ln (g/t gold head grade) plus 53.205%.

Figure 13.3-7 Variability Test Au Head Grade Vs. Au Recovery

Table 13.3-5 shows copper concentrate assays from the locked cycle tests for each of the metallurgical composite. The salient points from the concentrate assay are:

Copper grades in concentrate are well over the target of 22%.
The level of zinc in the ARC composite is above the typical penalty threshold (i.e. > 2%) and additional investigation should be undertaken to determine whether blending and management controls are required.
The level of mercury in the PHC composite is above the typical penalty threshold (i.e. > 10 g/t) and additional investigations should be undertaken to determine whether blending and management controls are required.


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Table 13.3-5 Copper Concentrate Assay – Main Elements


Element	ARC	PHC	KC
Aluminium (%)	0.30	0.36	0.37
Antimony (%)	0.044	0.030	0.017
Arsenic (%)	0.10	0.15	0.011
Bismuth (g/t)	37	34	18
Chlorine (g/t)	130	<50	<50
Cobalt (g/t)	24	46	62
Copper (%)	28.2	28.6	27.0
Fluorine (g/t)	170	170	120
Gold (g/t)	8.26	7.66	8.83
Lead (g/t)	1,748	336	296
Magnesium (%)	0.04	<0.01	0.05
Mercury (g/t)	5	12	1
Molybdenum (%)	0.012	0.015	0.42
Nickel (g/t)	58	62	58
Silver (%)	26	18	38
Zinc (%)	2.17	0.72	0.14


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14	MINERAL RESOURCE ESTIMATES

14.1 Introduction

The following section describes both the oxide and the sulfide resources for the La Arena deposit. An updated model for the oxide resource was created in September 2014.

The sulfide resource quoted in this report is the pre-existing January 2013 model. A small drilling program was completed in 2014 testing the sulfide breccia at the top of the sulfide domain, however this was completed after the sulfide Cu-Au study commenced and the results have no material effect on the sulfide project. There have been additions to the oxide resource during 2014 due to additional data received from the Reverse Circulation (RC) infill drill program undertaken in the Calaorco area. The major changes in the oxide resource model for 2014, as displayed in Figure 14.1-1 and Figure 14.1-2, are:

The addition of a moderate amount of oxide resource on the western side of the Calaorco deposit and at depth in Calaorco due to a large infill RC drill program.
The deepening of the Indicated oxide Resource by between 20 and 100m due to the additional data received from the drill program, and
The inclusion of more high-grade Tilsa Style domains in the Calaorco area due to acquiring additional resource data in 2014.

Estimation methods for Au oxide domains have not changed since 2013, being Localized Uniform Conditioning (LUC) for low grade oxide mineralized domains and Ordinary Kriging (OK) for the oxide background material and Tilsa structures.

Composites are 8m downhole, using a best fit approach. No capping has been applied to the composites due to the project to date positive grade reconciliation when compared to as-mined information.

Parent block sizes in the oxide domains have been increased from 5m (X) x 10m (Y) x 8m (Z) to 10m (X) x 20m (Y) x 8m (Z) as this better reflects the SMU achievable with the current mining fleet. LUC is the preferred estimation method for the low grade Au oxide mineralized domains, when compared to other estimators as this method honours local variability of grades within large low grade diffuse mineralized systems (Coupland, 2012), due to the non-linear method of estimating grades. Other linear estimators (ID, OK) tend to smooth estimates, particularly in very low grade ranges. This has been a problem in the past for the project, particularly in the first year of operation

Estimation methods for Cu-Au sulfides are a mixture of OK (for Cu) and Inverse Distance Cubed grade estimation (ID3) for all the other elements. Broad lithological domains are used (a sandstone and an intrusive domain) and hence the model is a bulk style approach. A re-logging program is


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underway with the aim of determining more discrete domains based on vein density and vein type. A large resource of moderate grade Cu-Au sits at depth below the open pit reserve, and this area has discrete zones of moderate grade Cu (0.5%) and Au (0.3 g/t) mineralization. The controls on the orientation of these zones are yet to be determined with confidence and it is hoped a re-logging program, currently underway, will assist in this process.

Figure 14.1-1 Plan Projection Low Grade Oxide-Gold Domains, Cu-Au Sulfide Project and Drilling


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Figure 14.1-2 Cross Section 9126375N Displaying 2014 (Dec2013) and 2015 (Sept 2014) Resource Model (Oxide Domain Only)

14.2 Database

The drillhole database utilised for the resource estimate for oxides is dated 6th September 2014. Four spreadsheets were exported from the Rio Alto Acquire database and were imported into Datamine for basic validation procedures. The database for the sulfides remains as per the January 2013 technical report.

Table 14.2-1 Database Files Used in Oxide Resource Estimate


Data File	Date	Size (kB) and Details
Assay.xls	06-09-14	25,769 kB
Collar.xls	01-09-14	139 kB
Survey.xls	01-09-14	1,001 kB
Lithology for Gis.xls	01-09-14	1,995 kB

Drillholes were checked for sample overlaps and gaps with no errors were noted. Assays were also checked for any extraordinary high grades, with no issues found, and the digital assays for 5 holes were checked against the final laboratory paper reports with no errors noted.


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14.3 Geological Modelling

14.3.1Geology

The sedimentary package is subdivided into sandstone, siltstone, slate and shale-coal. These units are modelled using Leapfrog Software with no limitation on thickness. The intrusive is also modelled using Leapfrog Software and this boundary is often a complex one (see Figure 14.3-1). The intrusive is a hard boundary for grade estimation for all elements.

Oxidation domains are determined from logging. The transition from oxide to sulfide domains is generally a visually dramatic change and therefore a transition domain does not exist in this resource model. However, all elements show a relatively smooth transition across this boundary and so it is a soft boundary for grade estimation purposes between oxide and sulfide domains.

Late stage Andesitic dykes have been included in the sulfide portion of the model. They have mainly been included for geometallurgical purposes. They have not been used to constrain grade estimates.

Alteration domains (Argillic, Advanced Argillic, Potassic and Phyllic) show weak to moderate correlation with Au and Cu mineralization and therefore are not used as constraints on grade estimation. They are used in the Cu-Au sulfide project for geometallurgical purposes.

Figure 14.3-1 Isometric View Looking North Displaying Intrusive/Sediment Boundary and Drillholes with Au Grades at La Arena


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14.3.2Gold Estimation Domains

The sulfide domains are separated into two broad lithological domains with a hard boundary separating the sedimentary package from the intrusive. The difference in tenor and style of mineralisation is marked between to two domains.

The oxide domains (see Figure 14.1-2) are separated into four large low grade domains (ZONECODES 100-400), a higher grade series of Tilsa (sub-vertical) structures (ZONECODE 500) and a lower-grade domain for the Astrid deposit (ZONECODE 600). The lower-grade domains use a nominal 0.05 g/t cutoff grade as a hard boundary, within which LUC is utilized to estimate the grade. The very low cut-off grade is chosen as it appears to be a natural break in the Au population and also it is below the current open pit operational cut-off grade of 0.10 g/t.

Tilsa style structures have a core (+/-1m) of very high grade but typically this is not well defined with resource drilling or blasthole drilling due to the combination of the thickness and orientation of these structures. Therefore, a lower nominal cut-off grade of 0.5 g/t is used to bulk up these structures into broader zones. This also facilitates more depth and strike length to the interpretations.

Both blasthole and resource drilling data are used to form the gold interpretation, which is completed in cross section only on irregularly spaced intervals, depending upon the drill spacing, at intervals of between 25m to 50m along strike. There is an apparent northern plunge to the higher grades within the Tilsa structures, and it is postulated that normal faulting may be causing this. Further drill testing is required to confirm this theory.


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Figure 14.3-2 Cross Section 916275N, Displaying Gold Domains


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14.3.3Copper Estimation Domains

Copper displays little zonation within the oxide domains and is therefore there are no discrete domains used for estimation purposes, other than a hard domain boundary between the sediments and the intrusive.

High grade copper within the intrusive shows some weak association with both potassic alteration and FPD2, although typically this is not clear cut, and high grades often transgress these boundaries by as much as 20-30m downhole. Therefore, no internal subdivision/domaining of the intrusive is used in grade modelling. A relogging program of the diamond drilling is underway focusing on vein density and vein type, in an effort to see if this can be used to domain higher grade Cu zones within the sulfide intrusive.

14.3.4Other Elements

Silver, Molybdenum and Arsenic show different trends between sedimentary and intrusive rock types, however no internal domaining is evident, and no clear trends between oxide and sulfide domains (particularly on the boundary) is evident.

Sulfur is separated into oxide and sulfide domains and also honors the sedimentary/intrusive hard boundary.

14.4 Sample Selection and Compositing

Only RC data was utilized for estimation of gold-oxide domains, with background oxide domains utilizing both RC and DDH data. The QP is satisfied that RC data generally provides a much better estimate of grade than DDH data, and that there is sufficient RC data within the gold-oxide domains to ensure reliable estimates for mine planning purposes. Sulfide domains utilized both RC and DDH data as RC data has a 300m depth limit and is therefore scant at depth. Blastholes were not used in any grade estimates.

Samples were composited to 8m lengths for gold oxide domains, compositing on a best fit length rather than a standard length ensuring no loss of sample length occurred due to any potential loss of residuals. Composites were then selected within each domain wireframe. Sulfide domains retain the 6m composite length used in the construction of the January 2013 Resource Model.

14.5 Basic Statistics

The Au oxide distributions are typically very well structured, with minimal outliers, due in part to the selected composite length. Grade capping for the Au oxide domains was not deemed appropriate due to the consistency of the data distributions, the large block sizes used in this resource model, and the positive grade reconciliation for the project to date.


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Domain 100 shows some instability at the top end of the Au log-probability plot, however the QP is confident that this is well constrained data in the centre of the Calaorco Open Pit and does not warrant cutting.

Cuts applied to the sulfide resource remain unchanged. Statistics from the January 2013 Resource Model data set for the sulfide domain are presented for completeness sake in Table 14.5-1.

Table 14.5-1 Basic Statistics Summary - Uncut Data


Domain	Mean	Variance	Std Dev	CV	Composites	Minimum Value	Maximum Value
Oxide High Grade Domains - All Data – Au (g/t)
100	0.23	0.26	0.51	2.18	3,681	0.002	23.68
200	0.27	0.04	0.19	0.72	2,177	0.008	1.82
300	0.34	0.22	0.47	1.41	1,136	0.002	5.17
400	0.17	0.10	0.32	1.84	1,090	0.002	4.36
500	1.17	3.14	1.78	1.51	1,103	0.002	27.97
600	0.07	0.01	0.12	1.74	1,167	0.005	1.13
Colluvium	0.2	0.07	0.27	1.35	1,272	0.025	2.56
Bulk Domains (Remainder Outside Oxide High Grade Domains) - Au (g/t)
Sandstone	0.06	0.0341	0.185	3.08	5,211	0.001	10.32
Intrusive	0.141	0.026	0.163	1.15	19,052	0.001	3.89
Bulk Domains (Remainder Outside Oxide High Grade Domains) - Cu (ppm)
Sandstone	171.8	398700	631.4	3.68	12,438	2.6	26,163
Intrusive	1,968	5,063,715	2,250	1.14	19,086	0.535	31,850
Bulk Domains (Remainder Outside Oxide High Grade Domains) - Ag
Sandstone	0.48	1.17	1.08	2.25	11,421	0.007	71.7
Intrusive	0.48	0.49	0.7	1.46	15,837	0.002	20.5
Bulk Domains (Remainder Outside Oxide High Grade Domains) - Mo
Sandstone	5.37	118.3	10.88	2.03	11,779	0.06	476.5
Intrusive	38.79	2,883	53.7	1.38	17,372	0.005	1,005.60
Bulk Domains (Remainder Outside Oxide High Grade Domains) - As
Sandstone	207.1	141,000	375.5	1.81	12,212	1.5	9,272
Intrusive	45.4	6,446	80.3	1.77	19,022	0.15	2,359.10


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Figure 14.5-1 Log Probability Plots of Au Composites in Gold Oxide Domains


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14.6 Variography

Gold semi-variograms for the oxide domains were updated to take into account the new RC data gathered in 2014. Declustered blastholes were used in conjunction with drillholes to assist in determining short scale ranges Gaussian transform semi-variograms were generated using Isatis software. These models were then back transformed to normal semi-variograms to obtain sills and ranges for the LUC estimation routines.

Two structure spherical models were used to model the back transformed semi-variograms. The semi-variogram orientations and anisotropies reflect obvious geological and visible data trends. Anisotropic semi-variograms were oriented with no plunge within the plane of the major axis and a variety of dips for the semi-major axis reflecting the known mineralization trends.

The gold oxide semi-variograms were of good to very good quality and generally had a well-defined nugget variance of between 10-30%. Fifty percent of the total variance was generally taken up within a range of between 25 and 50 m in the direction of the major search axis. Total ranges were between 120 m and 350 m. Tilsa domains (ZONECODE 500) display less geo-statistical structure than the lower grade domains, so the variography from domain 100 was used during estimation.

Figure 14.6-1 Directional Semi-variograms – Au –Oxide Domains 100 - 400


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Table 14.6-1 Semi-Variogram Models


DOMAIN	Name	Strike Orientation		Across Strike Orientation		Orthogonal Orientation		Relative Nugget	Sill 1	Range Structure 1 (m)			Sill 2	Range Structure 2 (m)
		Strike Orientation		Across Strike Orientation		Orthogonal Orientation		Relative Nugget	Sill 1	Range Structure 1 (m)			Sill 2	Range Structure 2 (m)
		Dip (°)	Azimuth (°)	Dip (°)	Azimuth (°)	Dip (°)	Azimuth (°)	(C₀%)	(C₁%)	Strike	Across Strike	Ortho	(C₂%)	Strike	Across Strike	Ortho
										Strike	Across Strike	Ortho		Strike	Across Strike	Ortho
										Axis	Axis	Axis		Axis	Axis	Axis
100 (Au)	Calaorco	0	150	-70	60	-20	240	12	51	56	56	34	37	340	250	100
200 (Au)	Ethel	0	150	-30	60	-60	240	20	26	80	80	50	54	180	200	88
300 (Au)	Ethel	0	170	-80	80	-10	260	7	63	50	46	26	30	100	140	50
400 (Au)	Calaorco	0	150	-60	60	-30	240	31	28	54	44	46	41	170	150	42
500 (Au)	Tilsa	0	150	-70	60	-20	240	12	51	56	56	34	37	340	250	100
600 (Au)	Astrid	0	125	-70	35	-20	215	8	38	70	62	22	54	170	124	68
Intrusive	Copper	0	140	-60	50	-30	230	18	44	80	80	80	38	620	460	200


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14.7 Block Modelling

The block model was generated using Isatis, Vulcan, and Datamine mining software. Variable parent block sizes were used for various components of the model which were then added together as the final model in Datamine.

The parent block size for the gold oxide mineralised domains is 10 mE x 20 mN x 8 mRL. The increase in the X and Y dimensions (from 5m x 10m) is a reflection of the real mining selectivity achievable with the current equipment and bench heights used at the mine..

The parent block size for the sulfide Intrusive domain remains at 10 mE x 20 mN x 6 mRL.

All wireframes were checked visually to ensure that there was adequate filling with blocks. All gold oxide domains were projected above the topographic surface to ensure that there were no edge effects in volume filling and then they were cut with the surface topography.

The block model parameters are shown in Table 14.7-1. Each block was characterized by a series of attributes as described in the Table 14.7-2.

Table 14.7-1 Block Model Parameters


	East	North	Elevation
Minimum Coordinates	814,500	9,125,200	2,200
Maximum Coordinates	817,500	9,128,200	4,600
Oxide Mineralized Domains
Parent Block size (m)	10	20	8
Minimum Sub-Block Size (m)	0.25	2.5	0.5
Sulfide Mineralized Domain
Parent Block size (m)	10	20	6
Minimum Sub-Block Size (m)	2.5	5	1


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Table 14.7-2 Datamine Block Model Attributes List


Attribute	Type	Description
IJK	Numeric	Parent Cell Identifier
XC	Numeric	Centroid of cell easting
YC	Numeric	Centroid of cell northing
ZC	Numeric	Centroid of cell RL
XINC	Numeric	Cell easting dimension
YINC	Numeric	Cell northing dimension
ZINC	Numeric	Cell RL dimension
ROCK	Numeric	0=Colluvium;1=Sandstone, 2=Intrusive; 3=Siltstone; 4=Slate; 5=Shale-Coal
RESCODE	Numeric	1=Measured, 2=Indicated, 3=Inferred,4=Unclassified
DENSITY	Numeric	Bulk Density
ZONECODE	Numeric	Oxide Mineralized Gold Domains: 100,200,300,400,500,600
OXIDE	Numeric	1=Sulfide; 2=Oxide
ALT	Numeric	Sulfide Alteration;20=Argillic; 30= Advanced Argillic; 50=Potassic; 60=Phyllic
PAF	Numeric	0=Non-Acid forming material;1=Potentially Acid forming material
AU	Numeric	Au (g/t) grade
CU	Numeric	Cu (ppm) grade
AG	Numeric	Ag (ppm) grade
MO	Numeric	Mo (ppm) grade
AS	Numeric	As (ppm) grade
S_PCT	Numeric	S (%) grade
DEFS_PCT	Numeric	Default Sulfur grade assigned to unestimated cells
UCS	Numeric	Unconfined Compressive Strength (UCS)
DEFUCS	Numeric	Default UCS values assigned to unestimated cells
RMR	Numeric	Rock Mass Rating (RMR)
DEFRMR	Numeric	Default RMR values assigned to unestimated cells
NS_Element	Numeric	Number of composites used in grade interpolation
PASS_Element	Numeric	Interpolation Pass
DIST_Element	Numeric	Distance to the nearest composite
VAR_Element	Numeric	Kriging Variance for estimate

14.8 Grade Estimation

Bulk gold oxide mineralized domains (Domains 100-400) were estimated using localised uniform conditioning (LUC) due to the diffuse nature of the grades within these bulk domains. Tilsa Structures (Domain 500), Astrid (Domain 600) and the Colluvium domain (ROCK=0) were estimated using OK methods as these are more discrete zones with little or no mining selectivity possible within an open pit mining environment.


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All boundaries used for the Au estimate within the gold oxide mineralised domains are hard boundaries except for the oxidation domains.

Gold oxide mineralised domains (100-400) used a single search strategy set to the range of the Semi-variogram for each domain. Negative kriging weights have been utilised and have been re-set to zero where appropriate. Sulfide Cu estimates are completed with OK methods. All other elements are estimated by ID3 methods.

All other domains/element combinations use a 3 pass strategy for estimation. The search strategy used in the model is as follows:

First pass searches used a maximum anisotropic range of 100 m for all grade variables.
If a block was not estimated in the first pass, a second pass search utilized a maximum range of twice the initial search radius.
If a block was not estimated in the second pass, a third pass search utilized a maximum range of five times the initial search radius.
The orientation of the search axes was identical to the Semi-variogram model orientations where appropriate. Where Semi-variograms were poor, the orientation used was that of the dominant geology.
Octant based searching was utilised in the first two estimation passes. A minimum of 4 octants needed to be estimated and each octant required the use of at least 2 composites to obtain an estimate.
All estimates were into parent cells and these estimates were discretised down to 5m (X) x 5m (Y) x 2.0m (Z).

Table 14.8-1summarises the search parameters used for resource estimation.


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Table 14.8-1 Search Neighbourhood Parameters Used for Resource Model Estimation


		Search Ellipse Ranges			Search Ellipse Orientation			First Pass		Second Pass			Third Pass			Max. No. of
Domain and	Variable		Semi-			Semi-		Min. No.	Max. No.	Search	Min. No.	Max. No.	Search	Min. No.	Max. No.	Comps
ZONECODE		Major	Major	Minor	Major	Major	Minor	of Comps	of Comps	Volume	of Comps	of Comps	Volume	of Comps	of Comps	From Any
		Axis	Axis	Axis	Axis	Axis	Axis	Used	Used	Factor	Used	Used	Factor	Used	Used	Drillhole
Domain 100	Au	340	250	100	0^oto 150^o	-70^oto 060^o	-20^oto 240^o	6	24	-	-	-	-	-	-	2
Domain 200	Au	200	240	100	0^oto 150^o	-30^oto 060^o	-60^oto 240^o	6	24	-	-	-	-	-	-	2
Domain 300	Au	150	200	100	0^oto 170^o	-80^oto 080^o	-20^oto 260^o	6	24	-	-	-	-	-	-	2
Domain 400	Au	200	150	80	0^oto 140^o	-60^oto 050^o	-30^oto 230^o	6	24	-	-	-	-	-	-	2
Domain 500	Au	75	70	30	0^oto 150^o	-75^oto 060^o	-25^oto 240^o	8	24	2	6	20	4	4	12	2
Domain 600	Au	75	60	30	0^oto 125^o	-70^oto 035^o	-20^o’215^o	8	24	2	6	20	4	4	12	2
Background Oxide	Au, Cu, Ag, Mo, As,S	100	90	40	0^oto 140^o	-70^oto 050^o	-20^oto 230^o	10	24	2	8	20	5	4	16	3
Sulfide	Au, Cu, Ag, Mo, As,S	100	90	40	0^oto 140^o	-70^oto 050^o	-20^oto 230^o	10	24	2	8	20	5	4	16	3


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14.9 Model Validation

The gold oxide block model has been validated by:

Comparing the estimates to nearest neighbour (NN) and inverse distance (ID3) estimates, by domain and by northing panels). There is good correlation between all estimation methods (Figure 14.9-1).

Figure 14.9-1 Swath Plot – Gold Oxide Domain 100

Visually comparing estimates to composite grades in cross section In general, there is a good visual correlation of composite grades and block model grades where there is sufficient data density to make this kind of comparison (Figure 14.9-2).


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Figure 14.9-2 Cross Section 9126200N Displaying Resource Model Cells and Drillholes

The key test for the model is to reconcile it to the project to date as-mined production (Table 14.9-1). Both tonnage and grade are conservative based on the 2014 as-mined production. This is a reliable outcome for mine planning purposes. Typically the data spacing at 25m x 25m is still too wide to adequately define some smaller zones, particularly in the western portion of the Calaorco Open Pit area, thus accounting for the tonnage under-estimation. The higher grades in as-mined production are probably attributable to an information effect.

Table 14.9-1 Reconciliation of Resource Model


		AS-MINED (TRUCK WEIGHTOMETER)			RESOURCE MODEL (RC ONLY)			(DIFF%)AS-MIN/RES
	CUT-		Au			Au			Au
YEAR	OFF	Tonnes	(g/t)	OZ	Tonnes	(g/t)	OZ	Tonnes	(g/t)	OZ
2012	0.10	8,266,964	0.82	217,128	10,505,451	0.70	236,008	-27%	15%	-9%
2013	0.13	14,662,323	0.58	274,299	12,531,891	0.55	221,455	15%	6%	19%
2014*	0.10	7,259,443	0.52	121,855	6,802,144	0.47	103,578	6%	9%	15%
PROJECT TO DATE		30,188,730	0.63	613,282	29,839,485	0.58	561,041	1%	7%	9%

* 2014 As-Mined is only from January to June inclusive


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14.10 Ancillary Fields

The Resource Model has been depleted to the December 31, 2014 topographic surface. The bulk densities used in this Resource Model, as displayed in Table 14.10-1, have not changed from the January 2014 Resource Model.

Table 14.10-1 Bulk Densities Used In Resource Model


Lithology	Oxidation	Bulk Density
Sandstone	Oxide	2.52
Sandstone	Sulfide	2.6
Siltstone	Oxide	2.52
Siltstone	Sulfide	2.6
Slate	Oxide	2.4
Slate	Sulfide	2.4
Shale-Coal	Oxide	1.8
Shale-Coal	Sulfide	1.8
Intrusive	Oxide	2.32
Intrusive	Sulfide	2.49

RMR and UCS fields have been estimated into the model, based on geotechnical core logging and 6m point load tests for the majority of the 2012 DDH program. RMR and UCS have been domained based on geology and geotechnical domains and have been estimated with ID3 methods.

A pyrite wireframe (>1%) is coded into the model with all material inside this wireframe flagged as potentially acid forming material (PAF).


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14.11 Resource Classification

The resource estimate has been classified as Measured, Indicated, and Inferred Mineral Resources based on the confidence of the input data, geological interpretation, and grade estimation. This is summarized in Table 14.11-1 as confidence levels of key criteria. An example of the style of classification adopted is displayed in Figure 14.11-1.

Table 14.11-1 Confidence Levels of Key Criteria


Items	Discussion	Confidence
Drilling Techniques	RC and DC – Good quality with good sample return. DDH removed from oxide resource.	High
Logging	Standard nomenclature adopted.	Moderate
Drill Sample Recovery	Good for RC and all diamond core.	Moderate - High
Sub-sampling Techniques and Sample Preparation	2m samples are reliable to adequately represent both styles of mineralization.	High
Quality of Assay Data	Recent data is reliable, based on QAQC results and observed and documented practices. Historical data set is of lower confidence.	Moderate - High
Verification of Sampling and Assaying	Assessment of sampling has been completed on site. Reconciliations are positive on grade project to date.	High
Location of Sampling Points	Survey of all collars conducted with DGPS by professional surveyors. Topographic surface is detailed. Downhole surveys of good quality; recent RC drilling has used a GYRO tool.	Moderate - High
Data Density and Distribution	Drilling on a notional 50m x 50m spacing for the Sulfide Domains consisting of RC and DC drilling to establish continuity, and is on a nominal 25m x 25m to 50m x 50m spacing for the majority of the oxide domains, except for at depth, in the inferred resources.	Moderate - High
Audits or Reviews	Logging and mapping checked on site.	Moderate - High
Database Integrity	Assay certificates checked.	High
Geological Interpretation	Mineralization interpretations are considered reliable for oxide domains. Sulfide domains require more detailed logging and interpretation for Cu and Au This work is underway.	Moderate-High
Estimation and Modelling Techniques	Uniform Conditioning is considered industry standard for deposits similar to this. Ordinary Kriging is industry standard method for bulk low grade Cu deposits that display zonation of grades.	High
Cut-off Grades	Reasonable cut-off grades applied for the proposed mining method. The theoretical cut-off grade of 0.07 g/t and the practical cut-off grade (used in the operations at present (0.10 g/t) are tabulated. Recent metallurgical testwork suggests that recovery of material at 0.07 g/t is ~ 60% and therefore is not economic.	High
Mining Factors or Assumptions	Parent block size for oxides has changed in line with the SMU used at the mine.	High
Metallurgical Factors or Assumptions	Current and project to date recovery is in line with expectations at 84%. Current blend of oxide intrusive to sandstone breccia is 1:4 for stability reasons. Sulfide metallurgy is robust.	Moderate - High
Tonnage Factors (Insitu Bulk Densities)	Sufficient bulk density work for global averages.	Medium

The Resource Statement has been prepared and reported in accordance with Canadian National Instrument 43-101, Standards of Disclosure for Mineral Projects of February 2001 (the Instrument) and the classifications adopted by CIM Council in December 2005.

The resource classification is also consistent with the Australasian Code for the Reporting of Mineral Resources and Ore Reserves of December 2012 (the Code) as prepared by the Joint Ore Reserves Committee (JORC) of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Mineral Council of Australia.


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Measured Resources are limited to the upper portions of the resource, that are close to the as-mined areas and have at least 25m x 25m (or closer) spaced drilling.

Indicated Resources have a drill spacing of between 25m x 25m and 50m x 50m depending upon the mineralization style and the confidence of the geological interpretation. Inferred Resources have drill spacing of nominally >50mx50m.

Figure 14.11-1 Cross Section – 9126250N –Resource Codes with Drillholes (+/- 25m Window)

14.12 Mineral Resource

The Mineral Resource for the La Arena project is tabulated in Table 14.12-1 to Table 14.12-4 and includes the area of Mineral Reserve. The Resource is separated into the three clear mineralization styles of oxide Resource being Sediments, Intrusive and Colluvium.

The oxide resource is reported within an optimized undiscounted cash flow pit shell using metal prices of $1,400 / oz for Au and updated cost parameters. The resource is quoted at a 0.07 g/t Au cut-off grade with no constraints on copper, as this can be blended in the open pit operation. The sulfide resource is reported within an optimized undiscounted cash flow pit shell using metal prices of $1,400 / oz for Au and $3.50 / lb Cu and updated cost parameters. The sulfide resource is quoted at a 0.12% g/t Cu cut-off grade.


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Table 14.12-1 Mineral Resource – Oxide Sediments - Inside Resource Pit Shell


Resource	Tonnes	Au	Cu	Ag	Mo	Au
Resource	(‘000 000 t)	(g/t)	(%)	(ppm)	(ppm)	(‘000 oz)
Measured	1.1	0.23	0.07	0.3	32.6	8
Indicated	100.8	0.38	0.01	0.5	4.1	1,234
Measured and Indicated	102.0	0.38	0.01	0.5	4.5	1,243
Inferred	2.2	0.34	0.01	0.4	2.9	24

Table 14.12-2 Mineral Resource – Oxide Intrusive - Inside Resource Pit Shell


Resource	Tonnes	Au	Cu	Ag	Mo	Au
Resource	(‘000 000 t)	(g/t)	(%)	(ppm)	(ppm)	(‘000 oz)
Measured	9.4	0.28	0.15	0.4	61.6	86
Indicated	19.7	0.22	0.06	0.7	9.7	137
Measured and Indicated	29.1	0.24	0.09	0.6	26.5	223
Inferred	0.3	0.14	0.01	0.1	2.1	1

Table 14.12-3 Mineral Resource – Oxide Colluvium - Inside Resource Oxide Pit Shell


Resource	Tonnes	Au	Cu	Ag	Mo	Au
Resource	(‘000 000 t)	(g/t)	(%)	(ppm)	(ppm)	(‘000 oz)
Measured	-	-	-	-	-	-
Indicated	2.6	0.34	0.01	0.2	2.5	28
Measured and Indicated	2.6	0.34	0.01	0.2	2.5	28

Table 14.12-4 Mineral Resource – Oxide Total - Inside Resource Oxide Pit Shell


Resource	Tonnes	Au	Cu	Ag	Mo	Au
Resource	(‘000 000 t)	(g/t)	(%)	(ppm)	(ppm)	(‘000 oz)
Measured	10.5	0.28	0.15	0.3	58.5	94
Indicated	123.1	0.35	0.02	0.5	5	1,399
Measured and Indicated	133.6	0.35	0.03	0.5	9.2	1,494
Inferred	2.5	0.32	0.01	0.3	2.8	25

Table 14.12-5 Mineral Resource – Sulfide Material - Inside Resource Sulfide Pit Shell


Resource	Tonnes	Au	Cu	Ag	Mo	Au	Cu
Resource	(‘000 000 t)	(g/t)	(%)	(ppm)	(ppm)	(‘000 oz)	(‘000 lbs)
Measured	-	-	-	-	-	-	-
Indicated	274.0	0.24	0.33	0.4	38.5	2,124	2,013,930
Measured and Indicated	274.0	0.24	0.33	0.4	38.5	2,124	2,013,930
Inferred	5.4	0.10	0.19	0.4	40.7	18	22,074

The author is unaware of any factors including environmental, permitting, legal, title, taxation, socioeconomic, marketing or political that may materially affect this resource statement.


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15	MINERAL RESERVE ESTIMATES

The Mineral Resources have been converted to Mineral Reserves based upon the following modifying factors:

Only Measured and Indicated Resources may be included
The Mineral Resources within an optimized pit limits are considered
Mining Dilution and Mining Recovery are applied
The mineralized rock is economically and technically feasible to extract

Each of these requirements was addressed in establishing the Mineral Reserves. The Mineral Reserves statement has been prepared according to the Canadian Institute of Mining, Metallurgy and Petroleum’s (CIM) standards. According to these standards, Mineral Resource classified as Measured and Indicated are reported as Proven and Probable Mineral Reserves respectively.

15.1 Types of Materials

There are two distinct mineralization types hosted at La Arena deposit: the oxide material and the sulfide material.

15.1.1Oxide Material:

This is the leachable oxidized material with economic Au grades related to a high-sulfidation epithermal system. This material is primarily located in the Calaorco pit which is currently in production. The production rate is approximately 36,000 t/day of ore to the leach pad. The oxide material is also present in the Tango 11 pit, North pit and South pit with higher Cu content. The oxide material is classified in two rock types:

	(1)	Sedimentary: Composed mainly of sandstone material; it is currently being processed with an average gold recovery of 87.5 %. It represents the principal source of leachable oxide material in the oxide Mineral Resource. It is located primarily in the Calaorco Pit and in minor quantities in the Tango 11 Pit.

	(2)	Intrusive: This is leachable material when blended with the sandstone rock. A revised blending ratio has been tested onsite with an average metallurgical recovery of 83.6 %. This material is located in the Calaorco pit, Tango 11 pit, North pit and South pit.

15.1.2Sulfide Material:

This is non-oxidized sulfide material contained within the porphyry Cu-Au deposit. The sulfide pit is on the East side of Calaorco pit and will provide sulfide feed material to the processing plant to produce a copper concentrate with a commercial gold content and low arsenic. A small zone of sulfide material has been found within the oxide pit limit containing grades grade above the cut-off


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that was successfully treated on the leach pad for gold recovery. Although various rock types and alteration styles have been described for the sulfide material, this material has been treated as one ore type for mine planning purposes and metallurgical recovery depending on the head grade only.

15.2 Assumptions and Parameters

The Mineral Reserves are constrained by a pit geometry that has been determined by technical and both cost and recovery economic inputs. The lists of assumptions used for the oxide pits are presented in Table 15.2-1.

Table 15.2-1 Pit Optimization Parameters for Oxide Mineral Reserves


Pit Optimization Parameters for Oxide Mineral Reserves
Mining Parameters	Units	Value
Mining Dilution Factor	factor	1.05
Mining Recovery Factor	factor	0.98
Mining Cost Sediments (direct & indirect)	$/t mined	2.08
Processing Parameters	Units	Value
Ore processing rate	Mt/y	16.7
Processing Cost Sediments	$/t leached	1.55
Processing Cost Intrusive	$/t leached	1.55
General & Administration Cost	$/t leached	1.22
Gold leaching recovery intrusive	%	83
Gold leaching recovery sediment	%	86
Economics Assumptions	Units	Value
Gold price	$/oz	1,200
Payable proportion of gold produced	%	99.9
Gold Sell Cost	$/oz	12.37
Royalties	%	1

The base mining cost of $2.08 /t includes the indirect and direct cost of mining in-situ rock. This cost was calculated using the monthly cost reports from site in the first eight months of 2014. Oxide Intrusive material will be mined with sandstone, and blended on the leach pad in order to minimize ore rehandling.

A projected mining cost for the sulfide portion of the project was provided by the current mine contractor using the best equipment configuration for the sulfide project.

The base processing cost for sediments was also extracted from Rio Alto’s monthly reports and includes the operating cost, plant maintenance and the current power cost. The processing cost for sediments is based on the ongoing processing operation. The power cost has decreased up to 65%


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from the previous reserve technical report due to the recent connection of the mine site to the national power grid. A preliminary cost for processing the sulfide material was estimated by Rio Alto at $4.7 /t which was later confirmed by the Pre-Feasibility Study produced by Ausenco.

The sale price of gold used for all optimization studies is $1,200 /oz. The copper sale price used for the sulfide project was estimated at $ 3.0 /lb Discussion on gold and copper sales price forecasts are presented in Section 19.

The assumptions and parameters used for the sulfide pit limits are tabulated in Table 15.2-2

Table 15.2-2 Pit Optimization Parameters for Sulfide Mineral Reserves


Pit Optimization Parameters for Sulfide Mineral Reserves
Mining Parameters	Units	Value
Mining Dilution Factor	factor	1.05
Mining Recovery Factor	factor	0.98
Mining Cost	$/t mined	1.92
Processing Parameters	Units	Value
Ore processing rate	Mt/y	6.57
Processing Cost Intrusive	$/t milled	4.61
General & Administration Cost	M $/ y	22.6
Process Copper Recovery	%	Avg: 91.1%, Range: 75.9 - 92.0
Process Gold Recovery	%	Avg: 38.9%, Range: 29.45 - 45.5
Economics Assumptions	Units	Value
Copper price	$/lb	3.0
Payable proportion of copper produced	%	96.5%
Copper Sell Cost	$/lb	0.37
Gold price	$/oz	1,200
Payable proportion of gold produced	%	88.6
Gold Sell Cost	$/oz	8.0
Royalties	%	1.0

The mining cost was developed based on first principles for each of the mining activities in the open pit mining operation. Hauling cost was adjusted to truly represent the haulage profile for the trucks on the sulfide pit.

The mining cost consists of $ 1.32 /t of direct cost and $ 0.60 /t of indirect cost. The direct cost covers the cost of operation and maintenance of the equipment working on site, including fuel, labour, consumables, spares, contractor’s fee and ownership. The indirect costs are the expenditures on administration, equipment mobilisation, supervision, light vehicles, dewatering, grade control and other ancillary activities related to the mine operation.


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Mine operation for the sulfide pit will be conducted on a contractor basis and it is planned to start at the completion of the current gold oxide mine operation.

15.3 Pit optimization

The pit optimization was conducted using the Whittle® software package with updated economic parameters. The optimized economic pit shells selected for the basis of open pit designs were created using this software. Whittle is a well-known commercial product that uses various geologic, mining, and economic inputs to determine the pit shell with the largest net value.

The optimization for the oxide and sulfide zones was independently run as the resource block models were completed in two different stages. As the oxide pit will be mined first, the blocks inside the optimum pit shells for the oxide were taken out of the model prior to the sulfide run.

For the oxide deposits, the pit shell with a revenue factor equal to 1.0 was selected. The pit design corresponds closely to the pit shell at this revenue factor. As the oxide mine life is only predicted to be for the next six years and most of the capital investment costs for the oxide mine are already sunk, a discounted pit value analysis was not considered.

The relationship between the capital required and the mine production figures are not always proportional and it depends on the layout and infrastructure required for the project. At times there can be a large “step change” in cost as the reserve increases. There might also be logistical constraints that limit the reserves.

The main components, constraints and contributions to the overall project CAPEX with respect to the Mineral Reserve for the sulfide pit can be defined as follows:

Rock fill for the tailings dam construction
Waste dump capacities
Property limits and land required for new infrastructure
Exisitng infrastructure on site that needs to be moved

The Figure 15.3-2 shows a “Pit by Pit Graph” analysis including the operational CAPEX required to continually expand the Mineral Reserves. Operational CAPEX considers, at this point, the Tailings Dam (TD) construction plan and the Waste Dump (WD) capital requirements.


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Figure 15.3-2 Pit by Pit Graph – Including Tailings Dam/Waste Dump Capital Expenditures

Three increments can be defined in the graph; the first increment (from pit 5 to pit 9) requires a low CAPEX since no extra tailings facilities will be required as the tailings are backfilled into the Calaorco Pit. The economic rock mined and waste reported in the second increment from pit 10 to pit 20 will require additional tailings and waste facilities; however, the resulting pit economic value of pit 20 is significantly higher than pit 10. Due to the CAPEX required to expand the waste dump facility and specially the tailings dam at La Ramada, the third increment does not add value to the project. For that reason, the second increment up to pit 20 was selected as the final pit shell. The final pit design corresponds closely to pit shell 20 at a revenue factor equal to 0.8.

15.4 Cut-off Grades

The cut-off grade was established to maximise the pit’s revenue. The minimum cut-off grade was derived from an existing reference equation listed below:


	COG (g/t) =	(Treatment Cost + G&A)
	COG (g/t) =	(Recovery) x (Price - Sell Cost) x (1 - Royalty)

The resulting cut-off grade was used to determine the Mineral Reserves shown in Table 15.4-1. The by-products recovered in the concentrate or in the doré bars were considered to have too little value contribution to be included in the cut-off grade calculation. The economic assumptions presented in Table 15.2-1and Table 15.2-2 from the previous section (pit optimization) were used to calculate the cut-off grades.


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Table 15.4-1 Minimum Cut-off Grade (COG) used to define the Mineral Reserves


Minimum Cut-off Grade used to define Mineral Reserves
Ore Type	Units	COG
Oxide Reserves	g/t Au	0.10
Sulfide Reserves	% Cu	0.18

As stated above, a COG of 0.18 % Cu was used to define the sulfide Mineral Reserve; however, Rio Alto raised the COG value used in the first production period in order to obtain higher head grades in the schedule and better cash flow as a result. This is a common industry practice in mine operations depending upon the grade distribution and stockpile capacity on site. The material between the raised COG and the marginal COG (0.18% Cu) is stockpiled for processing at a later time when mine production is lower or the capital costs have been retired. In the mine production schedule all the material sent to the stockpile is reclaimed for processing.

The raised COG for the sulfide project as the Table 15.4-2.

Table 15.4-2 Cut Off Grade (COG) for the Sulfide Mine Schedule


	Periods (Years)
	Y1	Y2	Y3	Y4	Y4-Y10
Cut off Grade Cu (%)	0.45	0.45	0.30	0.18	0.18

15.5 Mineral Reserve Statement

The resource estimates discussed in Section 14 were prepared using industry standard methods and appear to provide an acceptable representation of the deposit. The Qualified Person has reviewed the reported resources, production schedules, and cash flow analysis to determine if the resources meet the CIM Definition Standards for Mineral Resources and Mineral Reserves, to be classified as reserves. Based on this review, the assessment of the Measured and Indicated Mineral Resource occurring within the final pit design at the oxide and sulfide deposits can be classified as Proven and Probable Mineral Reserves.

The open pit Proven and Probable Reserves, include existing and future stockpiles scheduled for processing and inventory. As the sulfide material represents a new treatment process with a major facilities upgrade, the sulfide has been reported separately from the existing oxide operation.

The sulfide ore material will be sent to the crusher for processing in a new mill/flotation circuit while the oxide ore material will be sent to the leach pad for gold recovery by hydrometallurgical methods is already in place and operating. Stockpiles will be used to temporarily stock low grade sulfide ore material for later processing.


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15.5.1 Oxide Mineral Reserves:

The oxide Mineral Reserves estimated from the oxide project are shown in Table 15.5-1 as of December 31st 2014. The Mineral Reserves are reported as in-situ dry million tonnes and include 5% mining dilution and 98% mining recovery using a cut-off grade of 0.1 g/t gold as stated in Section 15.4 Cut-off Grades.

Table 15.5-1 Oxide Mineral Reserves – as December 31^st, 2014


Classification	Material	Tonnage	Au	Ag	Cu	Au
Classification	Type	(‘000 000 t)	g/t	g/t	%	(´000 oz)
Proven	Sediments	1.2	0.22	0.32	0.07	8.6
Proven	Intrusive	8.7	0.28	0.33	0.15	77.7
Proven Stockpiled	LG stockpile	0.3	0.24	0.33	0.14	2.3
Total Proven	Total	10.2	0.27	0.33	0.14	88.6
Probable	Sediments	80.9	0.42	0.42	0.01	1,085.2
Probable	Intrusive	12.3	0.27	0.84	0.06	105.7
Total Probable	Total	93.1	0.40	0.48	0.02	1,190.9
Proven and Probable	Sediments	82.1	0.41	0.42	0.01	1,093.8
Proven and Probable	Intrusive	21.0	0.27	0.63	0.10	183.4
Proven Stockpile	LG stockpile	0.3	0.24	0.33	0.14	2.3
Total Proven and Probable	Total	103.3	0.39	0.47	0.03	1,279.5

Intrusive ore hosted within the oxide zone cannot be separated as a different ore type for processing as it needs to be blended with Sediments. The projected gold recovery used for the Mineral Reserve is directly related to the material blend 4:1 and 3:1 as it was established in leach pile tests.

In addition to the oxide material inside the oxide pit, there is an additional 7.2 Mt of oxidized intrusive material within the sulfide pit. This is material that will become available for processing on the leach pad as the sulfide ore is mined, however, this material was not upgraded to Mineral Reserve at this time due to the lack of metallurgic testwork performed on this material to date.

The Mineral Reserve Statement contains the total minable reserve for the deposits described in Section 15.1. The Mineral Reserve passed an economic test conducted on the production schedule. The results of the economic analysis are shown in Section 22.

Colluvium was not included in the Mineral Reserve due to the financial and operational benefits of not moving the existing access road away from the deposits. However, the colluvium material inside the Calaorco Pit was included in the Mineral Reserves as Sediment. The colluvium deposit is


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a small shallow unconsolidated deposit of approximately 2.0 million tonnes grading 0.34 g/t gold and it is located immediately South-East of the main Calaorco Pit.

A small portion of the sandstone sulfide material (1.4 million tonnes) is contained within the Oxide Mineral Reserve. This material cannot be physically separated from the oxide ore and it presents similar leaching properties to the oxide ore.

15.5.2 Sulfide Mineral Reserves

The sulfide pit will be operated as a continuation at the end of the current gold oxide mine operation. As Rio Alto operates the oxide project, the assumptions used for the sulfide project were adapted to represent the operation of the porphyry pit.

Mineral Resource classified as Measured and Indicated are reported as Proven and Probable Mineral Reserves respectively. There were no Resources classified as Measure Resource within the sulfide pit limits. The Table 15.5-2 presents the Reserves Statement as of 31st of December 2014.

The tonnage and grades reported are calculated with a 98% mining recovery and include 5% dilution. The calculation of the cut-off grade equal to 0.18% Cu is presented in Section 15.4.

Table 15.5-2 Sulfide Zone Mineral Reserves – as December 31^st, 2014


Category	Ore type	Tonnage	Cu Grade	Au Grade	Cu Content	Au Content
Category	Ore type	'000 000 t	%	g/t	'000 lb	‘ 000 oz
Probable	Sulfide Ore	63,1	0.43	0.31	579,406	633.2

There are 6.4 million tonnes of sulfide material containing material at an average Cu grade of 0.41% at the top of the Ethel Pit, one of the oxide pits. This material was not upgraded to Mineral Reserves because of the likely oxidization of the sulfide minerals which will negatively affect the metallurgic recovery in the flotation cells. The period over which this material may be stacked is estimated to be from 3 or 4 years.

As presented in Section 15.3, the above sulfide Reserves Statement are constrained by the CAPEX required to increase the storage capacity of the tailings dam and waste dump facilities whereas the discounted value of continuing to add reserves beyond year 10, break-even the cost of elevating the tailings dam and the cost of preparing the land to open up new waste dump areas. Nevertheless, the potential to increase the sulfide mineral reserves will be determined by further studies regarding these two limiting factors.


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16	MINING METHODS

Rio Alto Mining has been operating the La Arena Mine since March, 2011. The Calaorco and Ethel pits have been the two open pit mines operated since mining commenced operation. Conventional drill, blast, shovel and dump truck is the mining method used under an alliance style contract signed with a local contractor. Mining is carried out on two 12 hour shifts, operating 7 days per week.

Mine reconciliations for the La Arena Mine show an overall positive trend. During 2014, the ounces production exceeded the plan for the year due to grade control defining more ore than defined in the resource model. The tonnage of ore mined exceeded the plan by 32 % as shown Table .

Table 16-1 Actual vs Planned Production in 2014


Actual vs. Planned Production in 2014
	Actual Tonnes	Au g/t	Planned Tonnes	Au g/t	Difference Tonnes	Au g/t
Ore mined	15,274,666	0.52	11,576,660	0.62	+ 3,698,006	- 0.1
Waste mined	17,332,139	-	18,885,439	-	- 1,553,300	-
Strip Ratio	1.13		1.63

16.1 Geotechnical

Rio Alto commissioned Piteau Engineering Latin America SAC (Piteau) to conduct the geotechnical analysis for the La Arena deposits. Piteau completed a geotechnical study for Calaorco Pit in August 2012 and several reviews have been conducted by George, Orr and Associates (Australia) Pty Ltd since then, with the latest site visit report dated in February 2014.

The geotechnical conditions of the Calaorco pit and the Ethel pit were considered to be adequate and the investigations conclude that there are no signs of impending multi-bench scale instability in the current pit walls. The report includes ongoing discussions to steepen the pit walls, comments on the geotechnical mapping protocols, pit monitoring, blasting best practices, and more.


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Table 16.1-2 outlines the slope wall parameters recommended by Piteau for each of the four oxide pits.

Table 16.1-2 Geotechnical Parameters for the Oxide Pits


Geotechnical Parameters for Design – Oxide Pit
Design Parameter	Calaorco - Sandstone	Calaorco - Intrusive	Ethel (Tango 11)	North and South Pit
Bearing	140° to 300°	300° to 140°	90° to 270°	All
Face Angle	75.0°	60.0°	50.0°	60.0°
Catch Berm (m)	12.2*	11.1*	3.5	10.0
Bench Height (m)	16.0*	16.0*	8.0	8.0
Inter-ramp Angle IRA	44°	38°	38°	29°
Overall Pit Slope Angle	41°	34°	34°	27.5°

* At double-8m-bench configuration

Piteau was also retained by Rio Alto to conduct a geotechnical study of the proposed sulfide pit. The study consisted of the training, supervision and geotechnical logging of a 32,000 m diamond drillhole program that commenced in October, 2011 and was completed in February, 2013. The database includes geomechanical and oriented core logging data, Point Load Index (PLI) laboratory testing results, and surface/bench face mapping and documentation data. This information was used to develop the design criteria for the sulfide Pit.

The average Rock Mass Ratings (RMRs) (Bieniawski, 1976) range from 29 to 56 for the sedimentary units (fair to poor rock mass quality), 40 to 53 for the intrusive units (fair) and 29 to 40 for the breccias (poor). Rock mass shear strengths were evaluated using the semi-empirical Hoek-Brown (2002) failure criterion, which can incorporate varying levels of disturbance to the rock mass caused by mining. The shear strength of discontinuities were evaluated based on the results of laboratory direct shear testing, indirect or subjective measurements of discontinuity characteristics, and engineering judgment based on Piteau’s experience in similar rock masses at other large open pit mining projects.

Slope stability analyses included kinematic assessments based on the structural mapping and core/borehole logging data, and subsequent deep-seated limit equilibrium analyses of the interramp and overall pit slopes to validate the kinematic bench and interramp designs and evaluate the influence of the rock mass competency, discrete fault zones, groundwater pressures and the possible influence of shear strength.


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The Table 16.1-3 tabulates the recommended slope designs for the sulfide Starter Pit based on assessments of kinematically possible failure modes.

Table 16.1-3 Slope Designs for the Sulfide Starter Pit Recommended by Piteau*

* Memorandum 3123 on November 14^th, 2013 – Summary of Bench, Interramp and Overall Slope Stability Analyses Results and Design Criteria for the Sulfide Starter Pit. Michael Scholz.

The configuration proposed by Piteau was used in the pit optimization process; however, this model was too complex for practical application in the pit design stage due to the distribution of domains. A simpler version outlined in Table 16.1-4 was proposed by Rio Alto for the geotechnical zoning for the pit design stage.

Table 16.1-4 Geotechnical Parameters for Sulfide Pit


Geotechnical Parameters for Designing – Sulfide Pit
Design Parameter	West Wall	East Wall
Bearing	150° to 35°	350° to 150°
Face Angle	60.0°	60.0°
Catch Berm (m)	10.5*	9.5*
Bench Height (m)	16.0*	16.0*
Inter-ramp Angle IRA	39.0	40.5
Overall Pit Slope Angle	34.9	36.8

* At double-8m-bench configuration


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Using the revised geotechnical recommendation proposed for the sulfide pit, a step-out berm of 30-m width was incorporated into the design at the 3272 level. This berm provides additional catchment capacity for loose material and improves the overall stability of the wall. The geotechnical zones including the oxide Calaorco Pit and sulfide Pit are presented in Figure 16.1-1.

Figure 16.1-1 Pit Layout and Geotechnical Zones

16.2 Hydrogeology and Hydrology

In January, 2014, Montgomery and Associates (M&A) delivered the results of the hydrogeological studies that were conducted during 2012 and 2013. From the report, the following observations represent the current conditions in the pit and the groundwater requirements for operation.

Neither the Calaorco nor Ethel Pits have reached the water table, there was no pit dewatering. In future years, after the base of the Calaorco Pit reaches the top of the saturated water zone, there will start to be an impact on water levels.


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Because the Calaorco Pit is in the part of the project with the highest elevation, the water table is deepest in wells in this area. In Wells 1001 and 1002, which are about 20 m apart and were drilled for an aquifer test northwest of the Calaorco Pit, the depth to the regional groundwater table is about 165 m. In Well 1003, south of the pit, the depth to the regional groundwater is about 245 m.
At the Calaorco Pit, groundwater inflow, precipitation, and surface water runoff is collected in a sump and is used for dust control and process water. Excess water is treated and discharged on to the leach pad
The model shows, in an average climate year, the current water balance is positive. The projected annual groundwater inflow rate for the Ethel pit is 0.03 L/s in 2012, decreasing to zero by 2013. The projected annual groundwater inflow rate for the Calaorco pit is 1.3 L/s in 2015, decreasing to zero by 2016 as a result of the dewatering wells starting in 2015.
The maximum projected annual groundwater inflow rate for the sulfide pit is 9.4 L/s, which stabilizes at a post closure equilibrium rate of 5.4 L/s by 2042.
When the sulfide pit starts mining, all runoff from mine facilities will be collected in the reclaimed water pond at the concentrator. Excess water will be stored in the Calaorco TSF, on top of the tailings. The water balance for this scenario was calculated for an average year.

16.3 Oxide Project Mine Layout

The open pit designs are based upon an optimized pit shell followed by a detailed design and development of phase plans for each of the pits. Geotechnical parameters and recommended pit slopesare outlined in Section 16.1.

The following design criteria were used for the open pits:

The designs have a minimum cutback width of 50 m.
24 m wide haul roads. 12 m single lane ramps with passing bays at berm levels on the last 3 benches.
10% maximum haul road grade.

The Calaorco pit is currently developing stages 2 and 3. The development of Stage 4 will commence in 2016. A new stage 5 has been added as the Mineral Reserve has increased. The cutbacks have been designed considering best ramp positions, pit access and geotechnical recommendations. The minimum elevation for the Calaorco pit is 3,148 masl and highest pit wall is 475.0 m on the West side.

The Tango 11 pit is a single pushback to the South from the existing Ethel pit. There is a residual oxide mineral resource under the Ethel Pit that has been sterilised by the location of Leach Pad 4. This mineralised material has not been included in the Mineral Reserves.


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The North and South Pits (see Figure 16.3) are scheduled within the Mineral Oxide Reserve because their economic content of intrusive oxide material has acceptable gold grades. These two pits are located within the sulfide pit limits; therefore they are the primary source of the high copper grade oxide intrusive material.

Calaorco Pit is currently in operation and it will continue to mine oxide mineralization until 2018, when the pushback on the South of Tango 11 will begin. North and South Pit will be the last oxide deposits to be mined. The North and South pits fall within the sulfide pit limits and can be considered part of the pre-stripping work for the sulfide project.

The colluvium material located on the South-East of the Calaorco Pit has not been considered within the pit limits as the infrastructure such as the public road already in place restricts mining in this area.

The waste dump and leach pad are located close to the pits to minimize hauling. The proposed mining infrastructure for the oxide project is located inside the footprint of the modified EIA, except for the new waste dump extension (phase 5). To the North are Leach Pads 2, 3 and 4 which will be in operation until the end of 2016 when the capacity exhausted. The waste dumped during 2015 and 2016 at waste dump phase 4 will act as the base platform for the next leach pad extension phases 5, 6 and 7 on the South. These new leaching areas will host the oxide ore material from 2017 onwards.

The waste dump extension (phase 5) on the west side has been designed to allocate the total waste rock moved in the production plan. This area will be included in next environmental assessment process for permitting in late 2015. Land purchasing of this area is currently ongoing. The waste dump phase 5 is expected to be neutral waste rock so it will also act as base platform for the waste rock produces in the sulfide pit which is higher in sulfur.

Rio Alto has commissioned independent studies on the proposed leach pad facilities and new waste dump area to “Anddes Consultores S.A“(Anddes). Early in 2015, as this report is prepared, Anddes is collecting samples from the test leach pad and will test them for mechanical characterization on the blended material 4:1 and 3:1. As the blend ratio (sandstone/intrusive) has increased to 3:1 and 4:1, the material is expected to shows similar or better stabilities conditions than the previous 2:1 blend.

The oxide mine site layout has been designed taking into consideration the future sulfide pit and infrastructure so no sterilization will occur in the upcoming projects. In Figure 16.3-1, the final layout for the oxide pit is presented. This layout is considered to be the initial layout for the sulfide project.


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Figure 16.3-1 Final Oxide Project Layout: Pit, Waste Dump and Leach Pad


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16.4 Sulfide Project Mine Layout

The mine layout is presented in Figure 16.4-1. The sulfide pit will be developed using a four stage production schedule. The selection of the stages was based on the best discounted value for the project and a balanced ore/waste production schedule. The sulfide pit will start mining after the oxide Mineral Reserves are depleted. Figure 16.4-2 shows the pit stages for the sulfide pit.

Figure 16.4-1 Layout: Sulfide Pit and Main Mine Facilities

Figure 16.4-2 Sulfide Open Pit Mining Stages


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16.5 Mining

Ongoing mining of the oxide deposits are by local mining and civil contractor (STRACON GyM) under an alliance agreement style contract.

Currently, the material is drilled and blasted on 8 m high benches using 155 mm or 171 mm diameter blastholes and a moderate powder factor (0.42 - 0.61 Kg/bcm). Loading of ore and waste is with diesel powered backhoes, face shovels and excavators into 92 t payload rigid frame dump trucks. The ore is then hauled to the dump leach pad and waste is hauled to the waste dump.

The operating experience in the oxide pit has shown a consistent drill penetration rate between ore and waste. In addition, the deeper benches have not shown any changes to the mechanical properties of the rock mass. The unconfined compressive strength (UCS) varies from 60 Mpa in high alteration areas up to 120 Mpa in the bedrock. The average penetration rate is 43 m/h which approximately corresponds to a UCS of 100 Mpa. The Table 16.5-1 shows the drilling parameters in the La Arena mine operation.

Table 16.5-1 Drilling Technical Parameters


	Units	Ore Rock	Waste Rock
Hole Diameter	Inches	6 1/8	6 1/8
Drilling Pattern	Burden/Spacing (m)	4.3 / 5.0	5.2 / 6.0
Bench Height	Meters	8.0	8.0
Sub Drilling	Meters	0.8	0.8
Re-Drilling	Meters	2.7%	2.7%
Penetration Rate	m/hr	43.0	43.0

An Unsensitised Gassable Bulk Emulsion Matrix is the primary explosive used for blasting. The emulsion matrix is shipped as an oxidizer and must be sensitized with a chemical gassing technology to become detonable prior to use. This product is being successfully used in the oxide pit producing an acceptable rock size distribution after blasting. The operators are already familiar with the product and it will be easily calibrated to the new conditions in the sulfide pit. The logistics for this product are already in place at the mine site. The blasting parameters are presented in Table 16.5-3Table 16.5-2.

Table 16.5-2 Blasting Technical Parameters for Sulfide Pit


Item	Units	Ore Rock	Waste Rock
Power Factor	Kg/bcm	0.61	0.42
Explosive	SAN-G (%)	100	100


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The mine is being developed using a conventional load and haul truck open pit mining method. Ore grade control has been considered in the mining method, so proposed mining will be conducted in 8 meter benches to minimize dilution and ore losses. The operative bench height is in the order of 10 meters due to the swell factor applied after the blasting. Table 16.5-3 shows the operative loading heights.

Table 16.5-3 Loading Technical Parameters for Sulfide Pit


Items	Units	Value
Nominal Bench Height	m	8.0
Operative Bench Height	m	9.0 – 10.0

La Arena mine operates under an over-trucking model which means that the production will be limited by the loading fleet, not the truck availability.

The running surface on the haulroads of 24 m wide has been designed using three times the truck width following international operational practices and Peruvian safety regulations. Single-lane haulroads of 12-m width were incorporated in the bottom three benches of the pit to maximise ore recovery. The designed haulroads include two 0.5 m drains on each side of the ramp and one additional safety berm.

Due to the lower permeability, oxide intrusive material must be blended with sandstone before it can be processed. A sandstone/intrusive ratio of 4 parts sandstone to 1 part oxide intrusive was determined in order to process the blended material on the leach pads. The blend will be achieved by blending the material on the pad. The oxide intrusive loads will be dumped close to the lift crest to be then “pushed” by a trackless dozer. The ore rock will roll down the lift slope blending it with previously dumped sandstone material.

As the sulfide pit is an extension of the current oxide pit, Rio Alto will continue with the same mining method, operational structure and machinery that are currently utilized for the oxide operation as it represents the best value for the project. Initial negotiations have started with the contractor already on-site to operate the future sulfide pit.

16.6 Mine Production Schedule

The mine production schedules for La Arena project are summarized in Table 16.6-1 for the oxide pits and in Table 16.6-2 for the sulfide pit.

The tonnes are reported in-situ dry tonnes after applying 2% ore losses and a 5% dilution factor in both tables. As stated earlier, the sulfide deposit will be mined consequently after the oxide reserve is exhausted.


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Table 16.6-1 Mine Production for La Arena Mine Oxide Pit


Oxide Pit Schedule for La Arena Mine
Year	Oxide Ore	Waste	Total	Grade Mined	Gold Mined
Year	('000 t)	('000 t)	('000 t)	Au g/t	Oz t
2015	15,095	18,751	33,847	0.543	263.7
2016	16,350	29,452	45,802	0.435	228.4
2017	21,637	26,278	47,915	0.323	224.4
2018	16,665	30,445	47,109	0.363	194.7
2019	17,555	28,721	46,275	0.339	191.6
2020	16,033	13,404	29,437	0.343	176.7
Total	103,335	147,051	250,386	0.385	1,279.5

Table 16.6-2 Mine Production (Mill Feed) for La Arena Mine Sulfide Pit


Sulfide Pit Schedule for La Arena Mine
Year	Direct Feed	Reclaimed Ore	Ore Feed	Cu Grade	Au Grade	Waste rock
	('000 t)	('000 t)	('000 t)	Cu %	Au g/t	'000 t
2021	5,679	-	5,679	0.60	0.51	7,024
2022	6,010	536	6,546	0.67	0.45	2,242
2023	6,238	308	6,546	0.44	0.35	3,611
2024	4,488	2,058	6,546	0.35	0.27	6,516
2025	6,546	-	6,546	0.48	0.38	3,616
2026	1,634	4,912	6,546	0.28	0.21	8,620
2027	5,746	800	6,546	0.34	0.25	10,628
2028	6,546	-	6,546	0.40	0.28	9,704
2029	5,097	1,449	6,546	0.36	0.24	1,567
2030	560	4,521	5,081	0.24	0.18	81
Total	48,545	14,584	63,129	0.43	0.31	53,610

The high grade copper material in the oxide zone is typically present above the proposed Cu-Au sulfide pits in the North and South oxide pits. Development of these pits commences around the middle of year 2018.

There is adequate sandstone material in the schedule to be mixed with oxide intrusive, at the time of mining and dumping on the pad. No extra cost has been allocated to rehandle material for blending purposes. However, in the actual operation, rehandling may be required at times to maintain a constant hourly blend of four parts sandstone to 1 part oxide intrusive. If so, the rehandling costs are an additional $0.49 /t.


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16.7 Mining Equipment

A fleet of drills, 170 t face shovels, 92 t payload trucks and associated load and haul support equipment purchased by the contractor for the gold oxide project are currently on site and working. The equipment list operating at Calaorco oxide Pit is shown in Table 16.7-1.

Table 16.7-1 Primary Mining Equipment Operating at Calaorco Oxide Pit (as of Dec 31^st2014)


Mine Equipment	Quantity
170 t Face Shovel	4
92 t Payload trucks	30
Drills Fleet	3
Dozers (Track and Wheel)	4

For the sulfide pit, the primary equipment fleet has been selected to achieve the required production rates. The mobile equipment selection was conducted by anticipating the fleet requirements that could be used in the sulfide pit operation. Therefore, a fleet of drills, face shovels, 92 t payload trucks and associated load and haul support equipment were found suitable for the project. Table 16.7-1 shows at each stage of pit development, the size of machinery that will be required. Only the primary equipment units are listed.

Figure 16.7-1 Primary Mining Equipment Units

A general overview of the ancillary equipment needed for the production stages of mining is outlined in Table 16.7-2


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Table 16.7-2 Ancillary Equipment Fleet Size


Period	21	22	23	24	25	26	27	28	26	27
Tractor CAT D8T	1	1	1	1	1	1	1	1	1	-
Tractor CAT 834 H	1	1	1	1	1	1	1	1	1	-
Grader CAT 16M	2	1	1	1	2	2	2	2	2	1
Water Truck	2	1	1	1	2	2	2	2	2	1
Fuel Truck	1	1	1	1	2	1	2	2	2	1
Lube Truck	1	1	1	1	2	1	2	2	2	1
Excavator CAT 374	1	1	1	1	1	1	1	1	1	1

Rio Alto and the contractor have succeeded in managing a flexible fleet size during the current operation. The contractor charges a fixed rate when the equipment is not utilized for short periods or alternatively the equipment is moved to other sites for longer periods.


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17	RECOVERY METHODS

17.1 Oxide Process Plant

17.1.1 Processing Flow Sheet – Dump Leach

Gold is recovered at the La Arena Phase I project via dump leaching. In 2014 ore was mined from two pits; Calaorco and Ethel. Mined material is trucked to pads where it is dumped to form lifts. The lifts are irrigated with sodium cyanide solution. As the solution passes through the lifts gold is dissolved. Pregnant solution discharges from the lifts and flows into a pregnant (gold enriched) solution pond.

Solution is pumped from the pond to an adsorption, desorption and refining (ADR) circuit where gold is recovered onto activated carbon. The carbon is stripped of gold to form a solution and the gold is extracted by the process of electrowinning to form a precipitate.

The precipitate is then dried and mercury evaporated off, mixed with fluxes and smelted to produce doré. The doré is weighed, sampled and shipped to a refinery. The refined gold is then sold.

Slag produced as part of the smelting process is crushed and any prills of gold are recovered and recycled for smelting. Stripped carbon is regenerated and recycled to the adsorption circuit. A flow sheet of the dump leaching operation is shown in Figure 17.1-1.

17.1.2 Dump Leach Process

The mineralized material is transported from the mine face to the leach pad using dump trucks. Lime is measured out at the moment the material is unloaded whilst the irrigation pads are formed. Spray irrigation systems are used (square 6 m by 6 m spacing). The height of the construction for every level of mineral is 8 m.

The barren solution percolates through the dump to the geomembrane and flows by gravity to the PLS Pond. The solution dissolves gold from the top bench of the dump and also any residual gold in the benches below before reaching the geomembrane.

The front-end engineering study considered irrigating the mineral for 60 days with an operational irrigation flow of 10 L/h/m²and a design parameter of 11 L/h/m². With these variables, the ADR Plant is able to process solutions from 355 m³/h to 1,500 m³/h, producing pregnant solutions with contents of gold reflecting the mine head grade.


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17.1.3 Process plant

For the treatment of the pregnant solution in the ADR Plant, two of the five adsorption circuits in existence are available for use. Each circuit consists of six columns containing 4 tonnes of activated carbon. Considering all columns there is a total treatment capacity of 1,500 m³/h.

The carbon from the adsorption tanks, containing 3 kg Au/t, is taken to the desorption plant where the gold is extracted from the carbon using a sodium hydroxide and sodium cyanide solution. This solution is then re-circulated to obtain a concentrated gold and silver solution. The gold and silver is recovered through electrowinning to obtain a precipitate that is dried in the press filters, which then goes through the smelting stage to obtain the doré bars.

The barren solution that flows from the adsorption circuit exit goes through a series of stationary meshes to collect any fine carbon. This solution is stored in barren solution tanks, where the strength of the sodium cyanide will be readjusted, anti-scalant added, and the pH adjusted. It is then pumped back to the leach pad to irrigate the mineralised material, using the barren solution pumps and pipes. All solution flow is in a closed circuit.


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Figure 17.1-1 Dump Leach Flow Sheet


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17.2 Sulfide Process Plant

17.2.1 General

The process plant design incorporates an open-air layout that minimizes overhead cranage requirements by maximizing accessibility for mobile cranes for maintenance. This layout, shown in Figure 17.2-1, has taken account of the site topography and limits imposed by the preliminary locations of the pit, stockpiles, and waste dumps.

The concentrator will use a conventional processing flow sheet and industry standard equipment. Concentrator operations will be monitored using a control system from a centrally located control room. Sampling and stream assay monitoring will be via an automated system linked to the control system.

The process plant and associated service facilities will process ROM ore delivered to the primary crusher to separate a copper/gold flotation concentrate from tailings. The proposed process encompasses crushing and grinding of the ROM ore, rougher flotation, regrinding and cleaner flotation. Concentrates will be dewatered on site prior to being trucked to the port for shipping to third-party smelters.


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Figure 17.2-1 La Arena Process Plant Layout


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17.2.2 Design Criteria Summary

The overall approach was to design a robust process plant that could handle a wide range of ore variability and operating conditions and deliver good value for capital. The key project and ore-specific criteria for the plant design and operating costs are provided in Table 17.2-1.

Table 17.2-1 Summary of the Process Plant Design Criteria


Criteria	Units	Design
Crusher Feed	t/d	18 000
Crusher Availability	%	70
Crusher Throughput	t/h	1,071
Mill Throughput	Mt/y	6.57
Mill/Flotation Availability	%	92
Mill Throughput	t/h	815
Bond Rod Mill Work Index Design	kWh/t	7.6
Bond Ball Mill Work Index Design	kWh/t	9.7
Axb		121.3
Abrasion Index		0.0202
Specific Gravity	t/m³	2.89
Primary Grind Cyclone Overflow Grind Size P80	µm	106
Regrind Cyclone Overflow Grind Size P80	µm	25
Cu Head Grade Average	%	0.37
Cu Head Grade Design	%	0.75
Cu Flotation Recovery for average grade	%	90
Au Flotation Recovery for average grade	%	37
Rougher Residence Time	min	30
Cleaner 1 Residence Time	min	12.5
Cleaner Scavenger Residence Time	min	7.5
Cleaner 2 Residence Time	min	10
Cleaner 3 Residence Time	min	7.5
Concentrate Filtration Rate	Kg/m²/h	300
Concentrate Thickening Flux	t/m²/h	0.15

17.2.3 Plant Design Basis

The key criteria selected for the plant design are:

A nominal plant treatment rate of 18 000 t/d (18 kt/d),
Design availability of 92% (after ramp-up), which equates to 8,059 operating hours per year, with standby equipment in critical areas
Sufficient plant design flexibility for treatment of all ore types included in the mining schedule at design throughput.

The selection of these parameters is discussed in more detail below.


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Throughput and Availability

Ausenco has nominated the selection of a 7.32 m (24 ft) diameter SAG mill and a 5.49 m (18 ft) diameter ball mill for this plant. This circuit is suitable to achieve a throughput of 18 kt/d for design competency ore. Ausenco has nominated an overall plant availability of 92% or 8,059 h/a. This is an industry standard for a large plant with low abrasivity ore. Benchmarking indicates that similar plants have consistently achieved this level.

Processing Strategy

Typically, the full range in variability of ore parameters such as hardness and head grade are considered during process design. However, the current mining schedule only including the argillic and phyllic alterations, so the design does not include any allowance to treat the potassic alteration (which is the most competent and hardest of the three rock types).

Head Grade

The plant is designed to treat various tonnages of ore with a maximum head grade of 0.75% Cu.

Flow sheet Development and Equipment Sizing

The sulfide process plant flow sheet design for the La Arena circuit was conceptually based on those of comparable flotation plants. Figure 17.2-2 shows a process flow sheet for the La Arena sulfide plant.

Details of the flow sheet design and selection of major equipment for the various options are discussed in the sections below.


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Figure 17.2-2 La Arena Sulfide Project – Sulfide Process Plant Flow Sheet


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17.2.4 Unit Process Selection

The process plant design is based on a flow sheet with unit process operations that are well proven in the minerals processing and sulfide flotation industries, and incorporate the following unit process operations:

ore from the open pit is crushed using a primary jaw crusher to a crushed product size of nominally 80% passing (P80) 150 mm (combined crusher product and fines bypass) and fed onto the stockpile feed conveyor
conical stockpile of crushed ore with a live capacity of 12 hours, with two apron feeders, each capable of feeding 110% of the full mill throughput
a 5.4 MW SAG mill, 7.32 m diameter (24 foot) with 4.88 m EGL (16 foot), with pebble return conveyor system
a 5.4 MW Ball mill, 5.49 m diameter (18 foot) with 9.75 m EGL (32 foot), in closed circuit with cyclone cluster, grinding to a product size P80 of nominally 106 µm
rougher concentrate regrinding stage in one of 3.0 MW Ball mill, 4.88 m diameter (16 foot) with 7.61 m EGL (25 foot), in closed circuit with regrind cyclone cluster, grinding to a product size P80 of nominally 25 µm
rougher flotation consisting of seven 300 m³forced air tank flotation cells to provide a total of 30 minutes of retention time
cleaner 1 and cleaner scavenger flotation consisting of ten 50 m³forced air tank flotation cells to provide a total of 20 minutes of combined retention time
cleaner 2 flotation stage consisting of three 50 m³forced air tank flotation cells to provide a total of 10 minutes of retention time
cleaner 3 flotation stage consisting of five, 8 m³trough-shaped flotation cells to provide a total of 10 minutes retention time
concentrate thickening in a 13 m diameter high-rate thickener
concentrate storage agitated tanks prior to pumping to the concentrate filter concentrate filtration at site in a horizontal plate and frame pressure filter tailings thickening in a 35 m diameter high-rate thickener
Tailings Management Facility (TMF) for process tailings deposition in the Calaorco pit raw water pipeline to site to provide water
process water and distribution system for reticulation of process water throughout the plant as required. Process water is supplied from water reclaimed from the TMF, waste dump 1 collection pond, waste dump 2 collection pond, noncontact water pond, storm event pond, contact water pond and process operations, with raw water used as make-up as required.
potable water is generated by treatment of raw water in a potable water plant at the process plant. Potable water is distributed to the plant.


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plant, instrument, and flotation air services and associated infrastructure

17.2.5 Comminution Circuit Sizing

Primary Crushing

Based on the design throughput and ore characteristics, a jaw crusher is considered the most suitable primary crusher for the duty. The primary crusher, 47" x 63" (1 200 mm x 1 600 mm), will be located at the bottom of the ROM bin after a vibrating grizzly. The vibrating grizzly will bypass the fines around the crusher and reduce the feed rate to the crusher sufficiently to allow this machine to operate at these throughputs.

Stockpile Feed Conveyor and Coarse Ore Stockpile

Crusher product will be reclaimed from the crusher discharge chute by the fixed-speed stockpile feed conveyor, which will transferred the ore to the mill stockpile.

The stockpile will provide a minimum of 12 hours live capacity at the design SAG mill feed rate of 18 kt/d, though higher throughputs will reduce this capacity. The total capacity of the stockpile is approximately 2.5 days of nominal SAG mill feed capacity. Two apron feeders have been selected to reclaim ore from the stockpile, each able to deliver 110% of the design mill feed rate.

Comminution Circuit

The major comminution design parameters used for this study are:

BRWi of 7.6 kWh/t based on the samples tested at ALS Kamloops during KM3526 and KM3991
BBWi of 9.7 kWh/t based on the samples tested at ALS Kamloops during KM3526 and KM3991
A x b value of 121.3 based on the samples tested at ALS Kamloops during KM3526 and KM3991
target grind size P80 of 106 µm, based on the ALS Kamloops flotation testwork during KM3991.

Flotation testwork and mineralogy have indicated that sulfide rock from La Arena is moderately fine-grained, and rougher flotation recovery is relatively low to grind sizes up to about 150 to 180 µm. In order to achieve the specified design metal recoveries, Ausenco has nominated a primary grind size target of P80 of 106 µm.

SAG and Ball Mill Design

The SAG mill feed weightometer will be installed on the SAG mill feed conveyor to provide feed rate data for control of the reclaim feeders. The reclaimed crushed ore will be fed at a controlled rate to the SAG mill. The SAG mill will be equipped with a wound rotor induction motor and variable-speed drive system.


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Discharge from the SAG mill will gravitate through a trommel, where oversize pebbles from the trommel (scats) will be recycled back onto the mill feed conveyor where, they will be reintroduced into the mill. Undersize from the SAG trommel will gravity flow into the cyclone feed hopper.

The mills discharge slurry will be pumped via dedicated cyclone feed pump to a mill cyclone cluster, operating in a closed-circuit configuration with the ball mill. Water is added to the cyclone feed hopper as needed to achieve the required cyclone overflow pulp density.

Cyclone underflow from the cluster will gravitate to the ball mill. Discharge from the ball mill will gravity flow back into the cyclone feed hopper for reclassification. Cyclone overflow will gravitate to the rougher flotation circuit.

The specific energy and mill sizing were determined using Ausenco’s in-house method for the major ore types and are shown in Table 17.2-2.

Table 17.2-2 Mill Design Criteria


Criteria	Units	Design
Throughput	t/h	815
Mill Type		SAG
Pinion Power Required	kW	3,512
Mill Speed	% Nc	62 - 80
Mill Diameter Inside Shell	m	7.32
Mill Length, EGL	m	4.88
Installed Motor Power	kW	5,400
Mill Type		Ball
Grind Size P₈₀	µm	106
Pinion Power Required	kW	4,723
Mill Speed	% Nc	75
Ball Charge Volume Maximum, operating	% vol	38
Mill Diameter Inside Shell	m	5.49
Mill Length, EGL	m	9.75
Installed Motor Power	kW	5,400

Installed ball mill power of 5,400 kW incorporates the allowances for drive train losses to determine the motor power from the pinion power, as well as a design contingency to account for the accuracy of the models, calculations, and testwork used to determine the expected average pinion power.

The installed motor power for the SAG mill incorporates similar allowances, as well as an additional contingency to allow adjustment in the mill operating conditions to handle ore variability. The minimum recommended motor size for the SAG mill is 4,500 kW. However, a 5,400 kW motor has been selected to match the ball mill motor.


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17.2.6 Flotation Circuit Design

Flotation test work has indicated that, while good rougher recovery can be achieved at a moderately coarse grind size, concentrate regrinding is required to achieve saleable concentrate grades.

The ores contain pyrite and clay minerals. The ores will require pyrite rejection conditions and fine grind, but should be easily amenable to conventional flotation separation to produce saleable concentrate grades.

Circuit Type and Size

The design flow sheet selected consists of copper rougher flotation, copper concentrate regrind, cleaner 1, cleaner scavenger, cleaner 2 and cleaner 3 flotation stages. Cleaner scavenger tailings will report directly to the final tailings thickener. Copper cleaner 3 concentrate will report to the copper concentrate thickener. The residence times and flotation parameters for the copper flotation circuit have generally been based on the testwork parameters obtained on the various La Arena ores. The testwork flotation and design residence times are summarized in Table 17.2-3.

Table 17.2-3 Summary of Copper Flotation Residence Times


Flotation Time	Locked Cycle Tests (min)	Scale Factor	Specified Design (min)	Actual Design (min)
Rougher	15	2.0	30	33.6
Cleaner 1	5	2.5	12.5	18.8
Cleaner Scavenger	3	2.5	7.5	14.2
Cleaner 2	4	2.5	10	60.8
Cleaner 3	3	2.5	7.5	38.6

Flotation Circuit Configuration

Cyclone overflow will gravitate to rougher feed box, where collector and frother will be added. The feed boxes will flow to the rougher flotation cells, which are connected in series. Seven 300 m³forced-air tank cells have been selected to provide the required residence time for the rougher flotation. Additional dosing points for frother and collector will be located at points along the rougher banks.

Rougher concentrate will be pumped to the regrind circuit through a hopper and pump. Rougher tailings will gravitate to the tailings thickener.

Regrind cyclone overflow and cleaner 2 tailings, will be mixed with flotation reagents (frother and collectors) in the cleaner 1 feed box before flowing into the cleaner 1 cells. The cleaner 1 flotation circuit consists of three banks of six cells, 50 m³tank flotation cells. The cleaner scavenger flotation circuit consists of an additional two banks of four 50 m³tank flotation cells for a total of ten cells.


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Concentrate from the cleaner 1 flotation cells will gravitate to the cleaner 1 concentrate hopper and pumped to the cleaner 2 feed pump. Tailings from the cleaner 1 flotation cells will gravitate to the cleaner scavenger flotation cells.

Concentrate from the cleaner scavenger cells will gravitate to the cleaner scavenger concentrate hopper, flowing to the copper cleaner scavenger concentrate pump. The copper cleaner scavenger concentrate pumps deliver concentrate slurry to the regrind cyclone feed hopper. Tailings from the cleaner scavenger cells will gravitate to the cleaner scavenger tailings hopper, and then flow by gravity to the plant tailings thickener for disposal.

The cleaner 2 cells will be fed with cleaner 1 concentrate and cleaner 3 tailings. The cleaner 2 cells will consist of three 50 m³forced air trough flotation cells.

Concentrate from the cleaner 2 cells will report to the cleaner 2 concentrate hopper, and will then be pumped by the cleaner 3 feed pump to the first cleaner 3 flotation cell. Tailings from the cleaner 2 cells will gravitate to the cleaner 2 tails pump. The cleaner 3 cells will consist of five 8 m³forced air trough flotation cells.

Concentrate from the cleaner 3 cells will report to the cleaner 3 concentrate hopper, and will then be pumped to the concentrate thickener feed box. Tailings from the cleaner 3 cells will gravitate to the cleaner 2 first cell.

17.2.7 Concentrate Regrind

The regrind circuit consists of a regrind cyclone cluster hopper (where rougher and cleaner scavenger concentrates are combined) and a regrind cyclone cluster in closed circuit with a regrind mill.

The regrind mill operates in closed circuit. Regrind mill discharge is combined with rougher and cleaner scavenger concentrates and pumped to the regrind cyclone cluster.

Based on a design rougher concentrate mass recovery of 190 t/h, the required regrind net pinion power is calculated to be 2,693 kW. For this power requirement, a single 4.88 m diameter by 7.61 m length with a 3,000 kW motor has been selected for this application.

Ausenco recommends that additional concentrate regrind testing should be conducted during the next phase of test work to confirm the target regrind size and the power requirements.

17.2.8 Concentrate Thickening and Filtration

The concentrate thickening and filtration circuit consists of a single 13 m diameter high rate thickener and one pressure filter.

The third cleaner concentrate is pumped to the concentrate thickener feed box via a cross cut sampler.


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Flocculant is added to the thickener feed stream to enhance settling. The concentrate thickener overflow reports to the process water tank. Concentrate solids settle and are collected at the underflow at a density of 60-65% w/w solids. The thickener underflow stream is pumped to an agitated filter feed tank by peristaltic pump.

The filter feed tank provides 24 hours of surge capacity allowing filter maintenance to be conducted without impacting on mill utilisation. The filter feed is pumped to the pressure filter to produce a filter cake of approximately 9.5% w/w moisture. The concentrate filter cake from the filter is discharged by gravity to a stockpile and stacked in the concentrate shed for on-site storage by FEL.

17.2.9 Concentrate Storage and Load Out

Concentrate filter cake is discharged by gravity to a stockpile. A front-end loader (FEL) is used to optimise concentrate storage within the covered building. The covered building provides storage capacity for up to 7 days at average production rates.

Concentrate is loaded by FEL into road trucks for transport from the mine site to the storage terminal at the port facility.

An automated truck wheel wash washes concentrate from the road trucks as they leave the concentrate storage building. Wash water is recycled and solids are recovered in a sump and pumped to the concentrate thickener.

Road trucks are weighed on a weighbridge located at the main security gate prior to leaving the mine site.

17.2.10 Tailings Thickening Disposal and Water Recovery

The tailings thickener consists of a 35 m diameter high rate thickener to thicken flotation tailings to approximately 55% w/w solids and recover process water prior to discharge to the tailings management facility.

The rougher and cleaner scavenger tailings streams gravitate to the tailings thickener collection box where they are sampled for metallurgical accounting and fed to the tailings thickener. Flocculant is added to the thickener feed to enhance settling.

Thickener overflow flows to the process water tank from where it is used in the grinding and flotation circuits. The thickened underflow is pumped to the tailings management facility (TMF) via a tailings pumps.

The recovered water in the TMF is reclaimed by pontoon-mounted pumps and pumped to the Process Water Pond for distribution throughout the plant.


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17.2.11 Reagents and Consumables

Major process reagents are received and stored on site. Dedicated mixing, storage and dosing facilities are provided for each reagent.

Lime

Lime is used to depress pyrite in the flotation circuit by increasing slurry pH. It also assists in the precipitation of soluble metals and for neutralisation of mine waste water stream.

Quicklime addition to SAG Mill: Powdered quicklime is added to the SAG mill. The system comprises of a storage silo and rotary feeder. Quicklime is delivered to site in 26 tonne bulk road tankers and unloaded pneumatically to a storage silo. Quicklime is transferred from the silo at a controlled rate via the rotary valve and fed to the SAG mill via the SAG mill feed conveyor.

Lime addition to Flotation: A quicklime slaking facility delivers milk-of-lime to the flotation circuit. The quicklime slaking system is a vendor package slaking system comprising storage silo, feeders and slaking mill. Quicklime is delivered to site in 26 tonne bulk road tankers and unloaded pneumatically to a storage silo at the lime slaking plant.

The quicklime is slaked in the mill with process water to produce a milk-of-lime slurry at 20% w/w solids. A lime circulating pump and pressurised ring main delivers lime slurry to dosing points in the grinding (cyclone overflow), flotation circuits and water treatment facility. Pinch valves are used to control lime addition at each dosing point.

Primary Collector (Hostaflot 3403)

Hostaflot 3403 is used as the primary copper mineral collector in the flotation circuit.

Hostaflot 3403 is delivered as a liquid in 1,000 L plastic composite intermediate bulk containers (IBC). IBCs are connected to a fixed manifold and the reagent is pumped to addition points via dedicated dosing pumps. A standpipe is connected to the manifold to monitor collector levels in the IBC. Storage capacity is one IBC.

Secondary Collector (Aerophine 3418A)

Aerophine 3418A is used as the secondary copper and gold minerals collector in the flotation circuit.

Aerophine 3418A is delivered as a liquid in 1,000 L plastic composite intermediate bulk containers (IBC) and distributed using a similar pump/manifold system to the Hostaflot 3403.

Frother (MIBC)

MIBC is used to provide a stable froth in the flotation cells.


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MIBC is delivered as a liquid in 1,000 L plastic composite intermediate bulk containers (IBC) and distributed using a similar pump/manifold system to the Hostaflot 3403.

Flocculant

Flocculant is used as a settling aid in the concentrate and flotation tailings thickeners.

The flocculant mixing system consists of a storage bin, screw feeder, blower, auto jet wet mixer, mixing tank, storage tank, and dosing pumps.

Flocculant is delivered as a dry powder in 800 kg bags. An electric hoist is used to load bags at the flocculant system. Dry powder is transferred to a heated hopper and blown into the wetting head to produce a 0.25% w/v flocculant solution. Flocculant is mixed in an agitated tank and transferred to a storage tank.

Dedicated dosing pumps deliver flocculant from the storage tank to the respective thickeners. Standby dosing pumps are provided at each system.

Miscellaneous Reagents

Additional reagents are required in the plant; however, these are expected to be used in relatively small quantities. These include chemicals for the potable water treatment plant and cooling water systems. Antiscalant and biocide are dosed to the cooling water system.

17.2.12 Water Services

Raw Water

Raw water is sourced from water bores. Raw water from the bores is pumped to the plant raw water tank. The plant raw water tank serves as a combined raw water / fire water tank with the lower section dedicated for fire water service and the remainder available for general use.

Raw water is used to supply the following services:

potable water treatment plant
crusher spray water
process plant equipment
slurry pumps gland seal
cooling water system
make-up water for the process water system.

Fire Water

The raw water tank contains a dedicated firewater reserve with a minimum capacity of 288 m³. The firewater pumping system comprising three fire water pumps (electric, jockey and diesel)


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provides failsafe supply of fire water to fire hydrants and hose reels via a dedicated fire water ring main.

Potable Water

Raw water (transferred from the plant raw water tank) is treated in a vendor-supplied water treatment facility to produce potable water for the process plant.

A potable water tank provides storage capacity of 48 hours at average consumption. Potable water is reticulated for general use in the plant and supplies the plant safety shower system.

Accumulators are used to supply potable water to the safety showers and eye-wash stations in the event of power failure.

Gland Water

Water for the gland water system is sourced from raw water tank. Raw water from the raw water tank is filtered in a vendor-supplied water filter to produce gland water for the process plant.

The gland water treatment consists of a duplex filter/strainer to remove particulate material. Gland water is distributed to the plant by two fixed speed pressure gland water pumps.

Cooling Water

Cooling water is used in the SAG mill and ball mill to cool oil lubrication systems.

Cooling water is supplied from a closed-loop evaporative cooling system. Water from the cooling water tank is pumped through the evaporative cooler to the heat exchangers in the grinding circuits. Warm water is returned to the cooling water tank and recirculated.

Inhibitor and biocide solutions are dosed to the evaporative cooling tower by dedicated dosing pumps.

Make-up water supply to the cooling water system is from the plant raw water tank.

Process Water

Process water is comprised of tailings thickener overflow, concentrate thickener overflow, treated mine water, TMF return water and contact pond water.

TMF return water is stored in the process water pond. Process Pond pumps transfer water from the storage pond to the process water tank. Excess water in the process water tank overflows back to the Process Water Pond.

Tailings thickener overflow (which includes treated mine water) and concentrate thickener overflow reports directly to the process water tank for immediate distribution and use. Process


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water pumps distribute process water to the grinding mills; flotation and regrind circuits; and lime slaking plant.

Mine Water Treatment

Mine water is used for process water make up in the plant. Water from this source is potentially acidic and requires treatment with lime prior to use in the plant.

The mine water treatment plant consists of two agitated mine water treatment tanks arranged in series. Lime slurry is added to the water treatment tanks to neutralize pH and precipitate heavy metals.

Treated mine water gravitates to the tailings thickener feed box. Precipitates formed in the mine water treatment tank settle with the tailings solids and are pumped to the TMF. The remaining solution reports with the overflow to the process water tank.

17.2.13 Air Services

Low pressure air for the flotation circuit is supplied by two blowers (operating as one duty and one standby). The rougher cells and cleaner cells operate at different air pressures. Air pressure to the pressure to the copper cleaner cells is reduced to the required pressure via an in-line pressure control valve.

Two separate air compressors (one duty and one standby) provide high pressure air for plant instruments and general service points. Compressed air is dried and filtered to instrument air quality prior to storage in plant air receivers and subsequent distribution.

The primary crusher is serviced by its own dedicated air compressor. The system is equipped with an air dryer and filter system.


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18	PROJECT INFRASTRUCTURE

As described in Section 5, the existing oxide gold Leach project includes an open pit, two waste dumps, a fully lined leach pad, a pregnant liquor solution pond, a major events pond (storm water catchment), an ADR processing plant and sundry facilities. Access to all these facilities is on dual lane gravel roads.

18.1 Roads

Access to site is from Trujillo on a national highway that is dual carriage bitumen. There are no tunnels on this route however there are 5 small bridges. Typical weight restriction on these bridges is 45 t unless specific reinforcement is designed and. This is the same route used to construct the Lagunas Norte project for Barrick. The La Arena site can also be accessed from Cajamarca on a bitumen and gravel dual carriageway highway.

All construction and operating supplies and materials for the gold oxide project have been transported from the coast via Trujillo to site by truck.

A road diversion has been planned to prevent the public from passing through the mining area and to open up options for waste dumps and other future mine infrastructure for the future operations. The Colluvium resource will also be available to be mined once the road has been diverted.

18.2 Accommodation

Existing accommodation has been constructed out of prefabricated pressed tin and foam sandwich panels. Materials and construction are suitable for the climatic conditions experienced on site. Individual rooms are set up for two people and are inter-joined by shower and toilet facilities shared between 4 people. Messing is provided in one location with an industrial kitchen capable of producing over 2,500 meals per day. Industrial laundry facilities are also installed with capacity to support a 600-person camp.

18.3 Offices, Workshops and Storage

The main offices and satellite office buildings have been constructed from the same material as the accommodation buildings. The workshop and warehouse are steel framed structures with sheet metal roofs. Offices inside the workshop and warehouse are constructed out of the same pressed tin foam sandwich panels as the accommodation blocks. All buildings have been designed with all appropriate storage, containment, drainage controls and are engineered for the storm and wind conditions prevailing on site. As part of the site infrastructure works, these buildings will be relocated to Raumate (camp and administration block) or near the new process plant (Truck work shop) to allow the site to be utilised for the future sulfide project plant.


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18.4 Laboratories

Two laboratories are operational at site; an assay/analytical laboratory and a metallurgical laboratory. The assay laboratory is analysing all process plant and run of mine samples. The metallurgical laboratory is designed to support the gold oxide project. The existing assay laboratory will be expanded and a new metallurgical (flotation) laboratory will be added to accommodate the needs of the sulfide project.

18.5 Fuel and Lubrication

Current operations are supported by a fuel facility with capacity of 120,000 US gallons (454.2 kL) located near the workshop and warehouse facilities. A major Peruvian fuel supplier, Primax, has been contracted to supply fuel for the gold oxide project and to manage the distribution of fuel from the site fuel facility. The fuel farm is built with sufficient spill (contingency) containment for 100% of the total tank capacity. Delivery of fuel is from the port of Salaverry in 9,000 gallon tanker trucks.

All lubricants are currently supplied under contract by Mobil. Delivery is by 200 L drums warehoused on site in purpose built containment yards.

18.6 Power Supply

The site power for the gold oxide project is supplied from the 220 KV national grid to a new substation recently built by La Arena in 2014 and commissioned on October 25^th, 2014.

Power is then distributed by an internal power distribution network supplying 22.9 kV to all facilities. Step down transformers are positioned at each significant installation including the PLS Pond, the offices, the workshop and the warehouse.

As part of the sulfide project, a new high voltage overhead line will be added to provide power to the sulfide plant HV switch yard. Additional LV overhead lines will reticulate power to the on-site water management infrastructure requirements for the sulfide project.

18.7 Water Supply

Water supply to the current operation is pumped from a fresh water bore located approximately 1 km from the office and camp buildings. The bore is 80 m deep with the pump currently positioned 50 m down the well. This supply is capable of a continuous flow of 5 L/s. A 150 mm (6”) HDPE pipe line is installed from the bore to an 80 m³holding tank positioned 200 m from the office buildings. From this tank water is distributed to the oxide processing plant, workshop, offices, camp, and kitchen via a potable water filtration system. The water quality is good and the pH is neutral.


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The water demand for the La Arena sulfide project will include water for drinking, sanitation and make-up water for mineral processing. The total demand is expected to be about 165 L/s. Most of this water will be used for mineral processing, mainly in the crusher and concentrator.

Sources of water with average annual flowrates, to be used in the concentrator, are as follows:

Tailings Storage Facilities Reclaim Water: 98.5 L/s
North Pit Dewatering: 11.9 L/s
Waste Dump 1: 1 L/s
Waste Dump 2: 12.3 L/s
Non Contact Water Pond: 32.7 L/s (dump leach rain coat collection water)
Groundwater well PW1, with a capacity of 5 L/s

Raw water requirements at the processing plant are 8.6 L/s. This water will be sourced from an existing ground water well (PW1), with a capacity of 18m³/h (5 L/s) and additional bores in the same region to ensure sufficient water to cover fresh water requirements at the plant.

Raw water at the concentrator plant will be used at the primary crusher, reclaim system and conveyor system for dust suppression, reagent preparation, filter wash water and make up water for the cooling system for the grinding mills.

Potential groundwater supplies are abundant. Both the unconsolidated alluvium near the Yamobamba River and the fractured sandstone of the Chimu formation have favourable aquifer characteristics. Water for the existing oxide phase operations is withdrawn from a water supply well in the alluvium and additional test wells have been constructed in both the sandstone and the alluvium. Results show that potential yields range from 5 to 20 L/s.

18.8 Explosives

A Peruvian explosive company Famesa is contracted to supply explosives to site for the gold oxide project. Famesa fabricates blasting accessories and provides a down-the-hole charging and a technical analysis and monitoring service. The blasting products are trucked to site from Trujillo.

A high explosives storage magazine has been constructed near waste dump #1. Detonators, fuses and detonating cord are stored in a bund protected 20’ container; boosters are stored in a separate bund protected 20’ container. Bulk emulsion is stored in 5 elevated silos each with a capacity of 60 t and positioned approximately 70 m from the class 1 explosives. Near the silos is a pad for storage of bulk ammonium nitrate.

The blasting agents are transported to the pit and pumped down the hole by a purpose built blasting truck The blasting truck can produce and deliver down-the-hole three different blasting products; ANFO, Heavy ANFO or gassed emulsion.


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18.9 Leach Pad Design

The design of the leach pad is based on conventional pad technology modified to accommodate the mountainous terrain as is common in Peru. The first two phases of the pad have been built, commissioned and are in operation. Part of Phase 3 is under construction and part is under leach. Phase 1 was designed by Ausenco Vector (Lima, Peru). Phases 2 and 3 were designed by AMEC (Lima, Peru). Two reviews of the design and construction were performed by RRD International (Reno, Nevada). Anddes Asociados SAC (Lima, Peru) performed an optimization and conceptual design of Phase 4, a revised seismic risk analysis of the Phase 4 and ultimate lift, and will be performing the detailed design in 2014.

The design of Phase 4 includes removal of unsuitable material, earth moving to create the required grades, installation of the subsurface drainage system (“subdrains”), construction of the compacted clay underliner, installation of the geomembrane liner, and installation of the overliner gravel and drainage piping system. Construction started in mid-2014 and the leach pad will be commissioned for operations in 2015.

Figure 18.9-1 Leach Pad Phase 3 and 4


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As of January 01st, 2015, the Phases 1 through 4 have a remaining capacity of 20.5 Mt. The new Phase 5, Phase 6 and Phase 7 leach pad facilities over the existing waste dump area will provide an additional 56.6 Mt of capacity. The phase 8 expansion will be the ultimate phase given the current reserves and the ultimate leach pad will provide storage for 108.4 Mt of ore. The northern slope of the phase 8 expansion and pad buttress fall outside the approved EIA. These structures will be included in the modification of the EIA to be presented in late 2015. The approval is expected within twelve months which is one year before phase 8 is required. Figure 18.9-2 shows the general arrangement of the new leach pad expansions.

Figure 18.9-2 Leach Pad Expansion - Phase 5, 6, 7 and 8

18.9.1 Drainage and Geomembrane Liner System

The leach pad subdrain system for Phase 4 will be similar to that used for Phases 1 through 3. These consist of a network of perforated, corrugated double-wall HDPE pipes in 100 to 450 mm diameter. These pipes will be installed in trenches and then backfilled with drainage gravel. The subdrain system has been designed to collect the subsurface seepage and direct that away from the containment system. These drains serve the dual purpose of early detection of any leakage through the pad liner system. In the lower portions of Phase 4 these drains will be connected to those from Phase 3. In the balance of Phase 4 they will be directed to the west edge of the pad where flow will be collected in a small, lined pond. This will allow monitoring of water quality before either discharging the flow to the environment or directing to the process circuit.


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The geomembrane liner of the leach pad will consist of 2.0 mm thick LLDPE, textured on the bottom side only. The geomembrane will be installed over a 300 mm thick compacted clay liner (CCL). The clay for construction of the CCL will principally be borrowed from locally identified sources (both within and nearby the construction area). Some material may be imported from other sources if needed. Geosynthetic clay liners (GCL) may be used to replace some of the CCL to speed construction especially in steeper areas.

18.9.2 Pregnant Solution Collection System

The PLS will be collected over the geomembrane liner by an overliner collection system. The overliner system will consist of a 500 mm thick gravel layer, except on steeper slopes where it will be either omitted or replaced with a geosynthetic drainage material. This gravel will be principally low grade ore from the Calorco pit. Within the overliner gravel will be a network of perforated, corrugated double-wall HDPE pipes of 100 to 450 mm diameter. Phase 4 zone will be divided into areas called 4A and 4B. Phase 4A will consist of most of the new pad area (above Phase 3) while Phase 4B will be the lowest portion of the new pad plus new ore placed over the previous phases. Drainage from the 4A area will flow along the central berm and exit the pad to 4 sediment ponds before being routed to the process plant. The drainage from 4B will be connected to Phase 3 and those solutions will comingle.

18.9.3 Operational requirements

The Phase 4 leach pad will extend to the south to the existing Phase 3 area (which is currently under simultaneous construction and operations). A section of the Phase 4 pad will be built over the Tango 11 push back, in the Tajo Ethel area. This will require scheduling of the push back sufficiently in advance of the completion of Phase 4 to allow the requisite ground preparation (slope contouring, some structural fill, and so forth).

There are no other special operating requirements for the expanded leach pad. As the pad grows larger the task of coordinating available ore stacking area with the mine production schedule becomes easier, and with Pad 4 there will be sufficient stacking area for any foreseeable mining plan until the end of the mine life.

18.9.4 Geotechnical Investigation

Each phase of the leach pad has been investigated for geotechnical properties both during design and while under construction. The characterization of the Phase 4 area and the borrow materials to be used in its construction were based on an investigation of those areas, the characterization of and the lessons learned from Phases 1, 2 and 3. Phases 3 and 4 have additional complications due to the presence of old waste dump slopes, rock quarries, haul roads, and some areas with organic matter.


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18.9.5 Dump Leach Stability

As part of the conceptual design of Phase 4, Anddes reviewed the seismic risk analysis report prepared by Golder Associates in order to define the design basis earthquake and it is corresponding peak ground acceleration (PGA). The risk assessment considered both probabilistic and deterministic methods. The resulting PGAs are summarized in Table 18.9-1. The leach pad, and thereby the lifts, will be founded on a combination of rock (in some areas) and stiff soil (in others). Thus, the mid-range attenuation law has been selected and the resulting PGA is 0.35g. For the purposes of pseudo-static analyses it is generally accepted (see US Army Corps of Engineers and other studies) that 50% of the maximum acceleration can be safely used in the analysis; thus 0.18g has been used for the pseudo-static analyses and 0.35g for the displacement analyses. Anddes found that acceptably high static factors of safety against slope instability and acceptable low displacements during earthquake events can be achieved with normal design considerations.

Table 18.9-1It is most common to use the 475-year return interval PGA for lifts analysis during the operating life of the facility. This return interval equates to a 10% risk of exceedence in a 50-year period. The maximum credible event (MCE) is the largest seismic event that can be reasonably expected given the regional geology and tectonic setting. That event was evaluated from the deterministic analysis. The MCE ranges from 0.38g for top of bedrock to 0.42g for top of soil profiles.

The leach pad, and thereby the lifts, will be founded on a combination of rock (in some areas) and stiff soil (in others). Thus, the mid-range attenuation law has been selected and the resulting PGA is 0.35g. For the purposes of pseudo-static analyses it is generally accepted (see US Army Corps of Engineers and other studies) that 50% of the maximum acceleration can be safely used in the analysis; thus 0.18g has been used for the pseudo-static analyses and 0.35g for the displacement analyses. Anddes found that acceptably high static factors of safety against slope instability and acceptable low displacements during earthquake events can be achieved with normal design considerations.

Table 18.9-1 Peak Ground Accelerations from Seismic Risk Analysis


Attenuation Law	IBC Site Classification (2006)	PGA (amax) for Return Interval (years)
Attenuation Law	IBC Site Classification (2006)	100	200	475	975	2500	MCE
Youngs et al Sadigh et al (rock surface)	B	0.14	0.19	0.26	0.32	0.43	0.38 (rock) 0.42 (soil)
Cismid+Sadigh et al (rock surface)	C	0.17	0.24	0.35	0.47	0.66
Youngs et al Sadigh et al (soil surface)	C-D	0.25	0.33	0.44	0.55	0.73

Notes:
Return intervals are taken from the probabilistic analyses
MCE = maximum credible earthquake from the deterministic analysis


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18.9.6 Access Road and Perimeter Diversion Channel

The design of all phases of the leach pad includes a 6 m wide all-season perimeter access road, except in a few areas were the terrain does not allow it is construction. A perimeter berm will be constructed where a road cannot be installed. The road will be used during construction for access of heavy equipment and deployment of the geomembrane (and, if used, GCL) rolls, and during operation for routine access to the lifts and the exposed areas of the leach pad. The access road will include a diversion ditch along its outboard side, and the road will be cantered towards that ditch. This ditch will collect surface water from the constructed slopes, road surface, and catchment basin outside the leach pad area and divert that flow away from the process containment. The diversion ditch and its outflows will be armored with rock and/or concrete where needed to protect against erosion.

18.10 Tailings Storage

Work has been carried out in 2014 on a variety of tailings storage facility (TSF) modifications. The current study incorporates sulpide tailings storage in the Calaorco pit after completion of oxide mining. Additional work to optimise the tailings disposal strategy is required as part of the next project phase.

18.10.1 Calaorco Pit Tailings Storage Facility

The La Arena project has been in operation since 2011. Phase I consists of exploiting the Calaorco mineral deposit, which is primarily a gold oxide resource. The exploitation process consists of open pit mining and dump leaching to extract the gold. During the current operation, this deposit is being mined, which will create a large open pit in the near future.

The sulfide project consists of exploiting a copper-gold sulfide deposit, which has a mine life of approximately 10 years. This resource is located just north of the Calaorco pit. As part of the Feasibility Study for the sulfide project, La Arena wanted to explore the option of tailings disposal in the Calaorco pit to reduce and defer capital expenditures for this project (refer to Figure 18.10-1).

La Arena hired Anddes Associates (Anddes) to develop the detail design for the disposal of 35 Mt of tailings into the Calaorco pit along with developing the tailings transport and discharge system, water reclaim system, and emergency spillway, which is based on the February 2014 ultimate pit configuration for Calaorco.


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Figure 18.10-1: Calaorco Pit Tailings Disposal

The design criteria utilized in the development of the tailings storage facility (TSF) is presented below (Error! Not a valid bookmark self-reference.) and served as the basis for the development of the detail engineering of the in pit tailings disposal and ancillary facilities.

Table 18.10-1 Design Criteria for Tailings Disposal in the Calaorco Pit


Description	Unit	Criteria
Mine Production Rate	t/day	18,000
Tailings Production Rate
Daily	t/day	17,737
Yearly	Mt/y	6.4
Tailings Properties
Tailings Specific Gravity	unitless	2.88
Tailing Gradation (D50)	µm	43.4
Percent Solids from Thickener	%	55
Tailings Density from Thickener	t/m³	1.56
Average In situ Tailings Dry Density	t/m³	1.55
Tailings Storage Facility
Tailings Capacity	Mm³	22.7
Tailings Storage Facility Life	year	5
Tailings Beach Slope	%	0.5
Tailings Slope Subaqueous	%	3
Maximum Tailings Deposition	masl	3,417
Hydraulic Design
Storm Event	Year	500
Tailings discharge System	type	Pipeline w/ spigot Boxes
Water Reclaim System	type	Barge w/ pumps


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The study was completed in May 2014. The final design included lined sections of the pit walls with geotextile or geomembrane, between 3325 to 3425 masl, to reduce the potential for migration of tailings through the highly fractured zones of bedrock in the upper walls beyond the pit limits (also potentially into the southern sulfide pit), a complex tailings pipeline and disposal system, a water reclaim system that included a float pump station, and an emergency spillway located at the north end of the pit. The bottom of the pit was at an elevation of 3250 metres and the filling of the Calaorco pit occurs over three phases that spanned five years and fill the pit to an elevation of 3417 masl.

The aerial tailing beach slopes are assumed to be approximately 0.5%, based on the rheology and the subaqueous tailings deposition are approximately 3%. Based on Anddes’s design, a Capital, Sustaining Capital, and Operating Costs were developed for the Calaorco in-pit tailings storage.

In parallel, Ausenco was carrying out the overall project feasibility study for the sulfide project to be completed at the end of 2014, which incorporated the Anddes study on Calaorco as well as the La Ramada TSF for tailings deposition from year 6 through to 10 of the life of the sulfide operation.

In late 2014, La Arena completed a new resource-drilling program and model, which increased the overall mine oxide resource and at the same time increased the Calaorco pit dimensions in both depth and width. For the 63 Mt case and the larger Calaorco pit, Ausenco was tasked with looking at the potential of eliminating the La Ramada tailings facility and placing all the tailings in the Calaorco pit. A revised tailings storage capacity curve verses elevation was developed to determine the capacity of the new Calaorco Pit (refer to Figure 18.10-2).

The new Calaorco pit design has the capacity to store approximately 77 Mt (51 m³) of tailings at an elevation of 3388 m (masl) at an average dry density of 1.55 t/m³, based on consolidation due to the depth of the tailings. However, the southern sulfide pit intersects the eastern pit wall of the Calaorco at an elevation of 3356 m (masl), thus reducing the storage capacity to approximately 53 Mt, which is short of the required capacity.

By repositioning the west pit wall of the Southern sulfide Pit 20 meters and offsetting the east pit wall 25 to 30 meters, the sulfide pit intersects the Calaorco Pit wall at an elevation of 3,372 m (masl), 16 vertical metres higher, increasing the storage capacity by an additional 12 Mt (refer to Figure 18 6). This increases into total potential storage capacity to over 65 Mt for Calaorco pit, thus eliminating the need for the La Ramada TSF for the 63 Mt case.


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Figure 18.10-2: Calaorco New Pit Verses Old Pit Configuration Storage Capacity

Figure 18.10-3: Calaorco Pit Modifications to Achieve 63 Mt Tailings Capacity

The changes in the economic value of the sulfide pit from these changes in pit geometry have been assessed as less than a couple of per cent.

Instead of undertaking a detailed re-design by Anddes, Ausenco looked at the effects of the new pit configuration on the original design. In addition, Ausenco also developed a new tailings conveyance system and water reclaim, which is discussed in other sections of this report.

The new pit configuration has increased the capacity to store 35 Mt tailings below the elevation of 3325 m (masl) without any liner system. Above 3325 liner systems consisting of geotextile, soil,


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and geomembrane are require to reduce the potential for migration of tailings through the highly fractured pit walls. Based on the Anddes investigation, and overlying it with the new Calaorco pit configuration, Ausenco has developed a new pit liner system.

At closure the pit walls of the oxide pit are non-acid generating, however the sulfide tailings have the potential to generate acid. Therefore, during the last year of operation, the tailings will be processed to remove the sulfides and then a capping system shall be installed over the tailings, which will create a barrier between the sulfide tailings and the air, thus reducing the potential for those tailings to become acidic.

18.11 Sulfide Waste Rock Storage Facility

18.11.1 Waste Rock Production

Approximately 52 Mt of waste rock will be produced during sulfide project, which comes from exploiting the copper-gold sulfide deposit over 10 years (refer to Figure 18.11-1).

The waste rock generated from the sulfide project consists of non-potential acid generating (PAG) material to PAG material. The non-PAG to low-PAG materials (approximately 24.4 Mt) will be utilized to encapsulate the PAG material (27.4 Mt) in the surface waste rock storage facility (WRSF). The waste rock production schedule is highly variable and in some years no waste is generated. However, this will not impact the design and operation of the waste rock storage facility.

Figure 18.11-1: Waste Rock Production Schedule

18.11.2 Waste Rock Storage Facility Design

Currently, La Arena is mining the gold oxide deposit and is presently depositing oxide waste rock in to a storage facility located approximately one kilometre east of the Calaorco Pit in Dump No 2. Due to limited space for waste rock storage, the plan is to place the sulfide waste rock on top of the oxide facility (refer to Figure 18.11-2).


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Figure 18.11-2: Sulfide Waste Rock Storage Facility.

The design criteria utilized in the development of the waste rock facility is presented below (Table 18.11-1) and served as the basis for the development of the WRSF and ancillary facilities. There is additional capacity for storage of the Non-PAG and Low-PAG material within the waste rock storage facility with some minor modifications, along with stacking higher.

Table 18.11-1: Design Criteria for Sulfide Waste Rock Storage Facility


Description	Unit	Criteria
Mine Waste Rock Production Rate
Yearly	t/year	varies
Waste Rock Properties
Waste Rock Specific Gravity	unitless	2.88
Waste Rock Dry Density	t/m³	1.78
Waste Rock PAG	Yes/no	yes
Waste Rock Storage Facility
Waste Rock Facility Life	years	10
Waste Rock Facility Capacity	tonnes	31.9
Waste Rock Facility Capacity	Mm³	5
Lift Height	m	10
Angle of Repose	H:1V	1.5
Global Slope Angles	H:1V	2.5
Number of Phases	#	2
WRSF Stability
Static Factor of Safety	unitless	1.3 (short term) 1.5 (long term)
Psuedostatic Factor of Safety	unitless	>1
Deformation Analysis	m	<1 if psuedostatic FS is <1
Maximum Tailings Deposition	masl	3,417
Hydraulic Design
Storm Event	Year	100


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Ausenco reviewed information from previous studies, such as civil, geotechnical, and hydraulic design for past waste rock storage facilities. As described above the sulfide WRSF will be placed on top of the oxide facility. Due to the high acid generating potential of the waste rock, the facility will have a geomembrane liner system and collection system to capture any effluent for treatment (if required) or released into the environment.

In order to provide a suitable foundation and proper drainage for the sulfide waste storage facility, La Arena will need to shape the top of the oxide dump during its stacking to provide gravity drainage for the sulfide facility to the east for collection, monitoring, and treatment (if required). The WRSF footprint for the sulfide project is approximately 607,000 m²and will be constructed in two phases (approximately 303,500 m²each).

The top surface of the oxide facility will be shaped to create a 2% slope to the east to ensuring adequate drainage slope to compensate settlement of the oxide facility from stacking sulfide waste on top. A compacted low permeability soil will be placed over the oxide waste to provide a bedding layer for the placement of 1.5 mm single sided texture LLDPE geomembrane liner. Then a collection system, consisting of a network of perforated HDPE pipes, will be placed on the geomembrane to facilitate drainage along with an overliner to protect the geomembrane from damage during stacking of the waste rock. The collection system ties into a conveyance system that bring any effluent captured to the base of the WRSF to a 5,000 m³lined pond where the water will be released to the environment (if it meets discharge requirements) or pumped to a treatment plant.

In addition, raincoats will be utilized to minimize the infiltration of rain into the facility to minimize potential acid generation. Once the facility is stacked to its final configuration a soil cover will be placed to reducing infiltration and ingress of air into the sulfide waste rock.

18.11.3 Waste Rock Disposal Sequence

The first lift will be 3 metres to prevent damage to the liner and then subsequent lift will be 10 metres along with 10 metre wide catch benches (refer to Figure 18.11-3). The waste rock construction shall progress from north to south, with waste rock being progressively dumped outwards from a number of active fronts. This method relies upon the inherent strength of the existing waste rock pile to prevent rotational failure of material through the foundations at the toe of the dump. A constant supply of good quality solid rock waste is therefore necessary.


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Figure 18.11-3: Typical Stacking Plan

18.12 Site Infrastructure for the Sulfide Project

18.12.1 Design Criteria

Infrastructure design for the sulfide project has been based upon information supplied to Ausenco by La Arena and Montgomery & Associates.

Water flows used in the design for the Off Plot facilities are based on the water balance provided by Montgomery and Associates, Technical Memorandum, "Revised Water Balance for the Starter Pit", issued on 22^ndMay 2014. The water balance is based on using Calaorco Pit as the Tailings Storage Facility (TSF) for the first five years of operation. Other mine facilities incorporated in the water balance are the North Starter Pit, Waste Dumps 1 & 2, and Leach Pad which include PLS and PGE ponds, and Non-contact water ponds.

The water balance model also incorporates the effect of using the Calaorco TSF to accumulate water for four months prior to the start of sulfide operations.

18.12.2 Utilities

18.12.3 Sewage Treatment

A packaged sewage treatment plant will be located at the process plant site. This plant will treat effluent from the plant, plant workshop and warehouse located at the plant, plant offices, chemical and metallurgical laboratory.

18.12.4 Emergency Generator

An Emergency Generator is considered in the design of the process plant to ensure supply of emergency power to process equipment during power outages. A diesel storage tank is provided to supply diesel to the generators. The diesel tank will be contained within an earthen bund. Road transport of diesel will be in conventional tanker trucks and transferred to the storage tank.


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18.12.5 On site Infrastructure

18.12.5.1 General

Allowances have been included in the cost estimates for surface water management for the plant site, and site access road, based on the principles of diverting as much clean water as possible around the site facilities. A run-off pond will be required downstream of the process plant site within the project area.

A plant site run-off pond (contact water pond) will be required downstream of the plant site. This will collect sediment from the ROM pad, plant site. The contact storage pond has a capacity of 10 000 m³.

18.12.5.2 Off plot Facilities

The off plot facilities consist of the following systems:

North Pit Mine Dewatering

This system transports mine surface water through a submersible pump from the North Pit Mine dewatering pond to the Mine Water Treatment Tanks located in the process plant.

Mine water is treated with lime at the process plant to neutralise acidic water and precipitate heavy metals that may be present in this water. Treated water reports to the Tailings thickener and used in the process.

Waste Dump 1 Pond

This system transports surface run-off water at the waste dump 1 through a submersible pump from the Waste Dump 1 water pond to the Mine Water Treatment Tanks located in the process plant.

Waste dump 1 water is treated with lime at the process plant to neutralise acidic water and precipitate heavy metals that may be present in this water. Treated water reports to the Tailings thickener and used in the process.

Waste Dump 2 Pond

This system transports surface run-off water at the waste dump 2 through a submersible pump from the Waste Dump 2 pond to the Mine Water Treatment Tanks located in the process plant.

Waste dump 2 water is treated with lime at the process pant to neutralise acidic water and precipitate heavy metals that may be present in this water. Treated water reports to the Tailings thickener and used in the process.

Calaorco Reclaim water


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This system transports TSF return water via two submersible pumps from the Calaorco TSF to the Process Water Tank located in the process plant. In addition a stand-by pump is considered in the design as tailings return water is critical for the continuous operation of the plant.

The TSF will be used as a water storage pond for the supply of process water to the plant.

La Ramada Reclaim water

This system transports TSF return water via two submersible pumps from La Ramada TSF to the Process Water Tank located in the process plant. In addition a stand-by pump has been considered in the design as tailings return water is critical for the continuous operation of the plant.

The TSF will be used as a water storage pond for the supply of process water to the plant.

PLS / PGE ponds

This system transports PLS / PGE pond water via a submersible pump from the Main Event Pond (PGE) to the Process Water Tank located in the process plant.

Non-contact pond water

This system transports non-contact water (from rain coats) via a submersible pump from the Non-contact pond to the Process Water Tank located in the process plant.

Bore water

This system transports bore water via an existing pump from well PW1 to the Raw Water Tank located in the process plant. Additional bore water sources are required to satisfy the demand of fresh water at the plant.


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19	MARKET STUDIES AND CONTRACTS

19.1 Gold Sales

RIO´s La Arena gold oxide mine produces gold in the form of doré bars. The doré bar historical weighted average gold grade has been 82% and has ranged between 72% and 88%. The historical weighted average silver grade has been 6% with the remaining 12% grade of doré bars consisting mostly of iron and copper.

The weight of the doré bar combined with the assay values (site assay and as well as from an independent laboratory for the purpose of comparing them to the refiner assays) allows the calculation of gold and silver contents and thus the overall value of each shipment.

Rio Alto refines its doré bars at Metalor Technologies S.A., (Metalor) refinery in Marin, Switzerland.

In order to accelerate the gold sales process, once doré bars are delivered to Metalor’s vault in Lima, Metalor credits 95% of the estimated gold content in the La Arena gold bars to the account of Rio Alto.

Once final doré bar assays have been agreed upon between Rio Alto and Metalor, the remaining gold and silver contents are credited to the account of Rio Alto.

Typical shipping and refining costs are approximately $ 4.50 per ounce of gold refined, based on annual 220,000 payable gold oz and a gold price of $1,250 per ounce.

19.2 Gold Market

The gold chart presented in this section shows the price topping out at $1,900 per ounce in the 2H of 2011 coinciding with the US dollar bottoming during that same period.

Since then a much stronger dollar along with more confidence in the US economy has negatively affected the price of gold.


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Figure 19.2-1 Gold Price Trend on the last 10 years

But it was really 2013 the year in which gold sold off relentlessly closing at $1,238.60 per ounce. Year 2014 has actually been stable in comparison.

The gold price in 2014 has traded to a high of $1,392.60 in March and to a low of $1,130.40 in November.

Currently gold is trading either side of $1,200 per ounce so it is barely down for the year notwithstanding the important fall in the price of crude oil and while the US dollar index is up nearly 10% for the year.

So while gold is down marginally in US dollars for 2014 it is actually up 8% in Euro terms.

Gold’s resilience in the face of a rallying US dollar may could be signaling a bottom for the so far 3 year-long down cycle for gold that started in the 2H of 2011.

Physical Demand

Physical demand is a key indicator of the price direction gold. There are signs that demand for physical gold is recovering.

The Indian government recently lifted restrictions on gold imports, which will surely cause a pickup in physical buying. There has been good and steady buying in gold by central banks in 2014. According to the World Gold Council, "signs are that 2014 will be another solid year of strengthening reserves with gold." The Russian Central Bank has been a notable official sector buyer.

Russian nationals are possibly buyers of gold as they flee the deterioration of the ruble in light of the imposed sanctions and the fall in crude oil prices. Fearing further devaluations and the potential for hyperinflation, Russian nationals will be exchanging their rubles for as much gold as possible.


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Another factor to consider is that lower crude oil prices are a shot in the arm for China, which has been suffering from sluggish growth in 2014. Any pickup in the Chinese economy will increase demand for physical gold.

Gold Price Forecasts and AMR’s recommended long term prices for the Financial Evaluation of the La Arena sulfides project

The consensus from analysts is for the gold price to average $1,200 per ounce in 2015. There are no strong opinions on gold going forward but most believe that the downside for the gold price is surely limited now.

Although now the dollar is strong and chances are it will remain strong in the short term, interest rates around the world remain low as fears of deflation have gripped Central Bankers. AMR believes that a continuation of low yields is positive for gold.

In the medium to long term AMR believes that the gold price cannot stay down in the event of improving prices in base metals and other commodities.

As AMR is proposing to use an average copper price of $7,500 per MT it is difficult to imagine that the price of gold would remain at $1,200 per ounce with such improvement in the copper price. AMR recommends to use a long term price of $1,350 per ounce of gold for valuing the forward payable gold from the La Arena oxide and sulfide project as this price would be more in line with a long term copper price of $7,500 per metric tonne.

19.3 Copper Supply and Demand

Copper’s long lived bull cycle which actually started in 2003 peaked in the first half of 2011. However as the market began to top out, there was ample evidence that the copper price would not come crashing down as per the previous boom-bust cycles which were very characteristic of copper in the late 70’s, 80’s and 90’s.

This time the price has behaved in a similar manner to how the price traded in the late 50’s to the early 70’s. During that period the copper price remained well supported above the operating cost of even marginal copper producers. This was a period when global economic condition where buoyant and rapid industrialization was experienced in Japan and in the US.

This time it is China and a number of emerging economies which have entered into a period of rapid industrialization and urbanization. Expectations are that this trend will continue for the foreseeable future with China continuing to lead the way. It is quite clear that emerging economies will demand proportionately more copper over a longer period of time than that which was demanded by Japan and the US during the copper boom of the late 50’s and 60´s.


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On the supply side, the long overdue increases in supply as a result of the response by the mining industry to high prices did finally materialized with the start-ups of new mines and brownfield expansions. Expectations are for the market to stay in a – slight- surplus of between 300k and 600k tons of metal per year through 2016.

Beginning in 2017 the market balance will switch to a deficit mode because of the lack of new mines in the pipeline. Producers are facing a number of obstacles including higher capital costs to build new mines, higher labor costs, ore grades declining at mature mines, more stringent environmental permitting, higher taxes and pressures from local governments and communities that host mining operations.

The difficulties that producer’s face in building new mines will mean that the market will need to witness an exceptional price of copper to provide the necessary incentive for new investment in the development of new copper mines.

Therefore in the long run we do not believe the current $2.85 per lb., or $6,300 per ton price of copper is sustainable. However, in times of over supply conditions the price of copper falls to the average cost at which the marginal (higher cost) producer operates. During such times the price may even overshoot this level on the downside.

This results in the closure of high cost production, thus helping to eliminate the over-supply condition and preventing the prices from falling further. It is estimated that the marginal copper production costs, including sustaining capital costs, at the 90th percentile today averages around $6,800/MT.

In this 2014 and for the next two years 2015 and 2016 we expect an unusual large number of green-field copper projects to come on-stream. Since demand is not very strong due to slow worldwide economic growth, we have seen the copper price dip below this level. The price will remain depressed until either demand catches up or until there is a significant cut in the supply or a combination of both.

Another factor that will support copper prices is the consolidation among the copper mining industry where more and more production is being controlled by a fewer number of mining houses. The current school of thought in this regard says that since the world production is in the hands of few mining houses it is likely they will be disciplined and delay development of large green-field projects until the timing is right.

We have seen a number of producers announcing delays in the development of green-field projects. Currently only the projects are extremely high quality and sufficiently low cost are getting the thumbs up

19.4 Contracts


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In 2011, Rio Alto entered into a Gold Prepayment Agreement and simultaneously entered into an off take agreement - the Gold Purchase Agreement, both of which being subsequently amended and restated. Rio Alto received advance sales proceeds of $50 million under the Gold Prepayment Agreement. At July 31, 2014, Rio Alto had satisfied the delivery and all other requirements under the Gold Prepayment Agreement. Pursuant to the Gold Purchase Agreement, Rio Alto agreed to sell gold produced from the Calaorco breccia and Ethel breccia bodies of gold oxide mineralization located within the property encompassing the La Arena Project, as more particularly described in the July 31, 2010 technical report prepared by Coffey Mining Pty Ltd. (the "2010 Report"). Rio Alto had previously reported that it estimated this amount to be up to 634,000 ounces of gold and that it had delivered the estimated remaining ounces of gold to fulfill its obligations under the Gold Purchase Agreement. The Gold Purchase Agreement makes reference to the 2010 Technical Report, which estimated that the ore bodies contained approximately 634,000 ounces of gold. However, as more particularly described in the 2015 Technical Report, the gold mineral resources and mineral reserves of the body of gold oxide mineralization within the Calaorco pit described in the 2010 Report have increased, while the Ethel pit has been exhausted through mining.

As a result of the increased gold mineral resources and mineral reserves, the contracting parties are working together to reach an agreed upon interpretation of the Gold Purchase Agreement. In the interim, Rio Alto has voluntarily re-commenced delivery of gold produced from the Calaorco pit to the counterparty in accordance with the terms and conditions contained in the Gold Purchase Agreement.


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20	ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

The existing Environmental Impact Assessment (EIA) was approved on July 20, 2010. La Arena is actively working to ensure the commitments and recommendations of the EIA are followed, including environmental monitoring and social management plan programs.

Currently the La Arena gold oxide dump leach mine is working at full capacity of the production permit of 35,990 t/d.

The environmental assessment report for the copper-gold sulfide project was initiated in the last quarter of 2012 and was presented to the Ministry of Energy and Mines in June 2013. This study was approved on December 27th, 2013. The surface footprint of this permit is not outside the footprint of the initial EIA for the oxide mine and according to the current regulations was treated as a modification of the original EIA. New environmental commitments include water quality monitoring stations, air, noise quality and social management programs. All the new commitments result from a compilation and update of all plans and programs currently ongoing at the operation.

The environmental impacts of the expansion from the existing oxide gold operation and required facilities will be delivered shortly as it is currently being prepared by a consulting company as a new modified EIA (MEIA). This new MEIA is expected to be approved by the third quarter of 2015.

20.1 Environmental Risk

The main environmental issues that may be considered intermediate risks are:

The long term management of fresh water supply is being assessed and permits are on- going.
New areas that will be significantly impacted by the disposal of both tailings and waste generated by the oxide and sulfide operations that could require environmental and social compensation measures.
For the future sulfide project the time it takes to obtain licenses and permits from regulators.
The long term management plan for acid rock drainage (ARD) for the sulfides in waste dumps and tailings and water quality. Studies have started and a conceptual model is being assessed.
Possible compensation for wetland areas that would be impacted by tailings and waste dump disposal areas. Studies of wetland locations, uses and services are on-going.
The costs associated with the closure of the mine.

These risks are mitigated by setting sound social and environmental policies together along with professional management programs.


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20.2 Social

La Arena signed a general agreement in December 2009 with the La Arena community (APEULA) that includes the relocation of the school, construction of a new town and other social support programs. The obligations committed under this agreement are ongoing

The main social aspects that can be considered as project risks are:

The need for ongoing relocation of occupants and acquisition of surface land from individual owners for the expansion of the sulfide project.
The existence of other mining operations located in the vicinity of the project whose community management methods may affect the surface land acquisition as well as how communities will perceive the project in relation to social and environmental demands.
The expectations that the project development will generate within the population living in or near the project. This includes local suppliers and potential contractors for the construction and operation.
New regulations to be put in place by governments that can affect the project.
The actual adverse social-political situation against mining activities elsewhere in Peru.

20.3 Mine Closure

The current Mine Closure Plan was approved by The Ministry of Energy and Mines (MINEM) through R.D, 394-2013-MEM-AAM (October 22, 2013).

La Arena is presently hiring consultants to prepare the modification of the Mine Closure Plan to include the changes of the Modified Environmental Assessment for the copper-gold sulfide project. The initial design has been optimized and state-of the art technology has been incorporated into the plan. The new modification of Closure Plan will comply with the current mining regulations. According to the regulations, the Closure Plan needs to be updated every 3 years.


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21	CAPITAL AND OPERATING COSTS

21.1 Oxide Gold Project - Operating Expenditures

Operating costs are tracked and well understood. The mine is operated through a local contractor under an alliance contract. The agreement provides the contractor the rights to operate and manage most of the mining activities from surveying to earthmoving work. At the end of each month, Rio Alto reimburses the contractor for their costs plus a pre-established management fee (contractor’s fee) which is treated like interest on principal. The equipment is effectively financed through the contractor. Contractor reimbursable expenditures are mostly limited to equipment rental and spare parts.

The mine contractor operates a fleet which comprises approximately thirty 92t trucks, four 170 t shovels, 3 blast hole drilling rigs and various support equipment for road maintenance, personnel movement and miscellaneous services.

Diesel is used to operate the mining fleet, support vehicles and electrical generators. Processing costs predominantly include materials (reagents), diesel generated electrical power and labour. The Table 21.1-1 shows the total operating cost by tonnes mined used for the life of mine plan for the oxide project.

Table 21.1-1 Operating Cost as at January 2015


	Operating Cost as at January 2015	($/tmined)
	Direct Cost
	Ore	$ 1.20
	Waste	$ 1.66
	Other Direct Cost	$ 0.02
	Total Direct Costs ($/t mined)	$ 1.46

	Indirect Costs
	Alliance Indirects	$ 0.47
	Non Alliance Indirects	$ 0.81
	Mine Operations Labour	$ 0.15
	Mine Maintenance Labour	$ 0.07
	Total Indirect Costs	$ 1.51
	Total Mining Costs	$ 2.97

	Mine Geology Cost	$ 0.04
	Total Processing Cost	$ 0.36
	Electrical Maintenance	$ 0.12
	Total Geo, Processing & Maint Costs	$ 0.52
	Total Cost ($/t mined)	$ 3.49


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All gold oxide Dump Leach mine site costs including general and administration functions such as accounting, insurance, logistics and other support services are included in the G&A costs detailed in Table 21.1-2.

Table 21.1-2 General and Administration Cost G&A (‘000 $)


General and Administration Cost (‘000 $)
	2015	2016	2017	2018	2019	2020	Total
Indirect Costs -Alliance	16,045	16,045	16,045	16,045	16,045	16,045	96,272
Indirect Costs -Non Alliance	27,380	23,561	23,561	23,561	23,561	23,561	145,186
Total G&A Cost	43,425	39,607	39,607	39,607	39,607	39,607	241,458

Taking into consideration the above costs and production inputs described in Section 16, the total operating cost and operating profits for the gold oxide project are presented in Table 21.1-3.

Table 21.1-3 Total Operating Costs and Operating Profit (‘000 $)


Total Operating Costs and Operating Profit (‘000 $)
	2014	2015	2016	2017	2018	2019	Total
Net Revenue	257,284	222,737	212,365	187,506	181,111	171,881	1,232,884
Total Operating Expenses	117,967	137,879	141,303	127,538	126,869	98,758	750,314
Closure Expenditures	1,500	1,500	1,500	1,500	1,500	1,500	9,000
Operating Profit (EBITDA)	137,817	83,358	69,562	58,467	52,743	71,624	473,570
Operating Profit Margin	53.57%	37.42%	32.76%	31.18%	29.12%	41.67%	38.41%

21.2 Oxide Gold Project - Capital Expenditures

The gold oxides Dump Leach operation has been in commercial production since January 2012. The capital costs (Capex) for the oxide gold project are developed and revised on an annual basis as part of the budget cycle. The capital costs include ongoing sustaining capital for the mine and dump leach operations as well as capital for the expansion of some of these facilities. The capital cost estimates exclude financial charges, working capital, taxes, sunk costs or future expansions.

Capex has been estimated by Rio Alto based on current operations. The total Capex estimates are detailed in Table 21.2-1.


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Table 21.2-1 Capex Additions for Oxide Gold Project (‘000 dollars)


	Capex Additions for Oxide Gold Project (‘000 dollars)
	2015	2016	2017	2018	2019	2020	Total
Construction	17,259	24,712	2,300	12,600	-	-	56,871
Plant	2,166	8,012	-	-	-	-	10,178
Community Relations	2,560	-	-	-	-	-	2,560
Permits & Engineering	1,396	-	-	-	-	-	1,396
Other Capex	2,103	2,000	-	-	-	-	4,103
Road Diversion	-	-	7,000	8,000	-	-	15,000
Land Purchases	14,548	-	-	-	-	-	14,548
Total Capex	40,032	34,723	9,300	20,600	-	-	104,656

21.3 Sulfide Copper Project - Operating Expenditures

The estimate is considered Feasibility Study level with an accuracy of ±15%. Project operating costs were determined by estimating for the three major cost centres: Mining, Processing and G&A.

Mining costs were estimated by Stracon. Process operating costs were estimated for processing each of the major ore types (Argillic and Phyllic Alterations), and vary with time according to the throughput and ore type. General and Administration (G&A) operating costs were estimated by Ausenco in conjunction with La Arena.

The operating costs were determined from first principles using input from a variety of sources:

Soft quotes from the current mining contractor
Reagent consumptions from metallurgical test work
Reagent & consumable costs from supplier quotations
Logistics & transport costs as supplied by La Arena
Manning levels from La Arena for mining and Ausenco for G&A and processing
Personnel salaries & overheads as supplied by La Arena
Previous study assessments

21.3.1 Mining Cost

The costs was prepared by the existing local contractor for the sulfide copper project, which would involve an expansion of the current oxide gold mine where the contractor is providing mining and civil contracting services in an alliance relationship with La Arena. The mining cost estimates are outlined in Table 21.3-2.


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The direct mining cost includes labour (operators and mechanics), fuel, consumables and ownership cost. The indirect mining cost includes staff, mine supervision, ore control, surveying and mine management cost.

21.3.2 Processing Cost

The battery limits for the determination of the process operating costs commence from the crushing facilities and continue through to tailings discharge into the TSF, and include the concentrate load-out and plant services.

The costs per tonne provided in this report are the average costs for each ore type over the life of mine. Processing costs include reagents, consumables, plant power, maintenance spares, mobile equipment as well as process plant labour. Process operating costs are summarized in Table 21.3-1

Table 21.3-1 Operating Process Cost Summary


	Description	$/t Ore
	Power	1.73
	Reagents	0.66
	Consumables	0.63
	Labor	0.66
	Mobile Equipment	0.07
	Maintenance	0.86
	Total Cost/t processed	4.61

21.3.3 General and Administration Costs

The G&A costs include camp operations, G&A personnel, offsite offices as well as miscellaneous project costs. An average G&A operating cost of $22.6 M per annum was factored into the cash flow model.

21.3.4 Total Operating Cost

Operating costs were divided into mine, plant and G&A departments. Mine cost is presented in direct and indirect Cost. Plant costs include labour, power, reagents and consumables, maintenance spares and consumables, vehicles and mobile equipment, metallurgical testing, and cover plant areas from ore reclaim through to the concentrate load out, including plant services. G&A costs include labour, light vehicle running costs, camp, off-site office, and miscellaneous costs.

Power consumptions for plant areas and G&A buildings were assumed to remain constant over the life of operation.

The vehicle fleet was assumed to remain constant for all years of operation. Mobile equipment used in mining operations is covered in the Mining Cost Estimate. Utilization of vehicles was based on the labour rosters of relevant personnel.


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Annual operating costs for maintenance spares and consumables were estimated at 5% of the capital cost estimate for each plant area.

The camp unit cost was based on the current operation at La Arena. These camp costs include camp lodging and catering personnel costs. Labour rosters were used to determine occupancy at the camp.

An overall summary of operating costs for the sulfide project is presented in Table 21.3-2 Total Operating Cost Summary

Table 21.3-2 Total Operating Cost Summary


		Units Cost
	MINING
	Mineral ($ / t mined)
	Dewatering & Grade Control	0.06
	Drilling	0.13
	Blasting	0.24
	Loading	0.29
	Hauling	0.47
	Auxiliary Services & Road M	0.20
	Stockpile Rehandling ($ / t rehandle)	1.22
	Waste ($ / t mined)
	Dewatering & Grade Control	0.06
	Drilling	0.09
	Blasting	0.17
	Loading	0.29
	Hauling	0.46
	Auxiliary Services & Road M	0.19
	Mining Direct Cost ($ / t mined)	1.32

	Mining Maintenance ($/mined)	0.02
	Mining Indirects ($/mined)	0.60

	PROCESSING
	Power	1.73
	Reagents & Consumables	1.30
	Labour	0.66
	Maintenance & Mob. Equipment	0.93
	Total Processing Cost ($ / t milled)	4.61

	GENERAL AND ADMINISTRATION (G&A)	$22.6 M/y


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21.4 Sulfide Copper Project - Capital Expenditures

The estimate for the process plant and associated infrastructure was developed with the classification of a Feasibility Study – Class 3 (AACEI Class 3) with an accuracy range of +15 to -8%.

The initial capital cost estimate is summarised in Table 21.4-3, with all costs stated in millions of dollars with no provision for forward escalation on any costs. The base date of the estimate is 4th quarter 2014.

Table 21.4-3 Initial Capital Estimate Summary


WBS	Level 1 Description	(Millions of Dollars)
2000	MINING	$0.9
3000	PROCESS PLANTS	$125
4000	SITE SERVICES AND UTILITIES	$11.3
5000	INTERNAL INFRASTRUCTURE	$44.5
7000	PROJECT PRELIMINARIES	$32.3
8000	INDIRECT COSTS	$54.5
9000	OWNERS COSTS	$17.0
8800	PROVISIONS	$28.6
Grand Total		$314

Sustaining capital expenditure of 2% of the process plant direct costs, 1% of the site services and 1% of the internal infrastructure was allocated annually to the project cash flow. Additionally the following items were included:

27.1 million dollars were split between in years 4 and 6 for waste dump expansion.
4.0 million dollars for liners for the Calaorco tailings facilities.
3.1 million dollars for contingency on sustaining capital
40.0 million dollars for mine closure in year 11.

A total sustaining capital cost of 61.2 million and mine closure cost of 40.0 million dollar are included in the financial model.


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22	ECONOMIC ANALYSIS

Under NI 43-101 rules, producing issuers are required to disclose economic results if a material change to the property has occurred. Oxide Intrusive material represents a material change in the Mineral Reserve statement. For that reason, an economic analysis of the mineral reserves of the La Arena oxide gold project was conducted and the results are presented below. A positive cash flow was achieved for all the periods which support the Mineral Reserves statement. Updated mining and processing input parameters and prices were used to estimate the Mineral Reserves, as discussed in Section 15.

The data for gold recovery is based on the metallurgy reported by Ausenco in Section 13. Capex and Opex costs are discussed in Section 21.

22.1 Peruvian Mining Taxes and Royalty

Revenue taxes and income taxes were considered in the financial assessment. Revenue based tax, operating profit based tax and income tax must be incorporated within the cash flow. Revenue tax includes Osinergmin Tax (Energy and Mining Regulatory Agency) and OEFA Tax (Environmental Supervision Agency). Operating based tax includes a Special Mining Tax and a Modified Royalty. The corporate income tax after the worker participation is 28% for 2015 and 2016, 27% for 2017 and 2018 and 26% from 2019 onwards. Pension plan is 0.5% of income before taxes. The worker profit participation is 8% of income before income tax and pension tax.

The Peruvian mining tax system was revised during 2011. Rio Alto is subject to the revised system which created two new forms of taxation on mining enterprises. One bill modified the existing royalty on sales of mineral resources. Of the two new forms of tax, only one applies to La Arena along with the modified royalty (MR). The amended bills applicable to La Arena may be summarized as:

Special Mining Tax (SMT)

It is applied on operating mining income based on a sliding scale with progressive marginal rates ranging up to 8.40%. As a tax on operating profit, normal operating costs excluding interest are deducted from revenue, an operating profit margin is determined and a progressive tax rate would be applied on the Operating Profit Margin Ratio. The calculated SMT used for economic analysis varies from 2.15% to a maximum of 3.18%.

Royalty Based on Operating Income (MR)

The modified royalty revises the mining royalty enacted in 2004 that required a payment ranging from 1% to 3% of the commercial sales value of mineral resources. The MR is applied on a company’s operating income rather than sales and is payable quarterly (the previous royalty was payable monthly). The amount payable is determined on a sliding scale with marginal rates ranging up to 12% applied to the operating margin. As a company’s operating margin increases, so does the marginal rate of the royalty. The MR use for this economic analysis varies from 1.28% to a maximum of 3.21%.


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Under Peruvian law, workers in the mining industry are entitled to participate in a company’s income before income tax (“Worker participation” on Table 22.2-1). This participation amounts to 8% of income before income tax. Each of the MR, SMT and worker profit participation taxes are deductible for the purposes of corporate income tax and pension tax.

22.2 Oxide Gold Project

The Table 22.2-1 summarizes the annual total taxes and royalties for oxide gold project.

Table 22.2-1 Annual Taxation for Oxide Gold Project


Annual Taxation for Oxide Gold Project (dollars)
	2015	2016	2017	2018	2019	2020	Total
Royalty Tax	3,568,451	1,044,413	578,134	369,269	433,050	1,102,026	7,095,343
Special Mining Tax	3,535,101	1,364,921	882,827	620,256	683,600	1,322,431	8,409,136
Subtotal Royalties	7,103,552	2,409,334	1,460,961	989,525	1,116,650	2,424,457	15,504,479

Worker Participation	8,100,941	3,984,029	2,737,076	1,965,495	2,191,366	3,622,017	22,600,926
Osinergmin Tax on Revenue (0.19%)*	488,839	356,379	-	-	-	-	845,218
OEFA Tax on Revenue (0.15%)**	385,925	289,558	-	-	-	-	675,483
Pension Tax on Incomes (0.5%)	465,804	229,082	157,382	113,016	126,004	208,266	1,299,553
Corporate Tax on Incomes (28%)***	26,085,031	12,828,573	8,498,622	6,102,863	6,552,185	10,829,831	70,897,106
Total Taxes	35,526,541	17,687,620	11,393,080	8,181,374	8,869,555	14,660,115	96,318,286

Total Taxes & Royalty	42,630,093	20,096,954	12,854,042	9,170,899	9,986,205	17,084,571	111,822,765

*	Osinergmin tax is 19% for 2015, and 16% for 2016.
**	OEFA tax is 15% for 2015, and 13% for 2017
**	Corporate taxes are 28% for 2015 and 2016, 27% for 2017 and 2018, and 26% from 2019 onwards

The Annual Cash Flow and Net Present Value (at a 5% discount rate) are presented in Table 22.2-2. The economics results show robust financial results for the oxide gold project with positive outcomes for all periods and a total cash flow of $ 218.6M or a discounted value of $ 180.5M.

The effects of changes in the major project assumptions and estimates were evaluated using the traditional approach of assessing variations in the metal prices, grades, operating cost, capex and metallurgical recovery. The analysis was carried out by changing the input parameters within the cost model and assessing the NPV (at a 6% discount rate). A +/-10% range was used as shown in Table 22.2-3


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Table 22.2-2 Net Cash Flow (After-Tax)


Net Cash Flow (After-Tax) (’000 dollars)
	2015	2016	2017	2018	2019	2020	Total
Pre Tax Cash Flow	59,296	48,635	60,262	37,867	52,743	71,624	330,426
Taxes	42,630	20,097	12,854	9,171	9,986	17,085	111,823
After Tax Cash Flow	16,666	28,538	47,408	28,696	42,757	54,539	218,603
Cumulative Cash Flow	16,666	45,203	92,611	121,307	164,064	218,603
Net Present Value (5%)	15,872	25,885	40,952	23,609	33,501	40,698	180,517
Cumulative NPV (5%)	15,872	41,757	82,709	106,318	139,819	180,517

Table 22.2-3 Sensitivity Analysis on the NPV


Sensitivity Analysis on the 5% NPV (’000 dollars)
	Low Range (-10%)	Base Case	High Range (10%)
	Gold Price of 1,150 $
Inputs Var.	$ 1,035	1150	$ 1,265
NPV (5%)	114,317	180,517	245,781
	Grade Avg of 0.39g/t
Inputs Var.	0.35 g/t	0.39 g/t	0.42 g/t
NPV (5%)	114,552	180,517	245,565
	Opex Avg of $ 3.00 /tonne mined
Inputs Var.	$2.70 /tmined	$3.00 /tmined	$3.30 /tmined
NPV (5%)	244,266	180,517	116,767
	LOM Capex Avg of $107,582M
Inputs Var.	$96,824	$107,582	$118,340
NPV (5%)	190,256	180,517	170,778
	LOM Recovery Avg of 83.99%
Inputs Var.	75.59%	83.99%	92.39%
NPV (5%)	114,552	180,517	245,565

22.3 Sulfide Copper Project

Ausenco has estimated an initial CAPEX of $ 314 million for the processing plant and all associated infrastructure such as camp relocation, power, drainage system, tailings facilities and contingency. An additional $ 61.2 million has been allocated for sustaining capital during the 10 year mine life, which totals an estimated CAPEX of $ 415.5 million (including $40 million Mine Closure), for the 63Mt starter project.


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The sensitivity analysis on the NPV and IRR for the sulfide project is shown in the Table 22.3-1 :

Table 22.3-1 Sensitivity Analysis on the NPV and IRR (in dollars)


					After Tax NPV
Interest Rate		2.5/lb Cu		2.75/lb Cu		3.0/lb Cu		3.5/lb Cu		4/lb Cu
0%	$	115,931,418	$	202,256,504	$	285,492,823	$	448,325,631	$	606,298,959
4%	$	57,148,889	$	124,695,433	$	189,213,692	$	314,925,219	$	436,670,624
6%	$	34,108,807	$	94,249,838	$	151,479,447	$	262,807,486	$	370,539,024
8%	$	14,441,544	$	68,205,343	$	119,203,512	$	218,270,395	$	314,068,574
10%	$	(2,366,623)	$	45,879,743	$	91,521,670	$	180,078,151	$	265,657,574
					After Tax IRR
		2.5/lb Cu		2.75/lb Cu		3.0/lb Cu		3.5/lb Cu		4/lb Cu
IRR		9.70%		15.41%		20.15%		28.22%		35.16%


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23	ADJACENT PROPERTIES

The region displays a particularly rich endowment of metals (Cu-Au-Ag) occurring in porphyry and epithermal settings, including the Lagunas Norte mine, the Santa Rosa mine, La Virgen mine, the Quiruvilca Mine, Shahuindo project, Igor exploration project and Tres Cruces development project (see Figure 7.1-1).

Barrick Gold operates and owns the Lagunas Norte mine which is located 13 km from La Arena deposit. According with Barrick’s website, the production of Lagunas Norte, is anticipated to be 570,000-610,000 ounces of gold for 2014 at an all-in sustaining costs of $560-$620 per ounce. Proven and probable gold reserves as at December 31, 2013 were estimated in 3.75 million ounces of gold.

During the 2^ndQuarter of 2014, Rio Alto announced the acquisition of Shahuindo project to Sulliden Mining Capital. Shahuindo Gold project is located 30 km from La Arena and contains similar geology conditions with total gold reserves of 1.02 M oz as reported in the NI-43-101 report of November, 2012. Optimization of the project’s design is ongoing by Rio Alto, however, is not the intention of this report to disclaim or update any information related to Shahuindo project.


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24	OTHER RELEVANT DATA AND INFORMATION

24.1 Oxide Project Development

The La Arena oxide project ongoing development schedule includes two key elements: the expansion of the leach pad and the waste dump. All other significant components are complete and operational.

Pad phase 3 will be completed by the first quarter of 2015. Phase 4 is also under development. The next expansion on phase 5, 6, 7 and 8 will provide an additional leaching capacity of 87.9 Mt.
The waste dump will be expanded to the East side of current leach pad where sufficient area for expansion exists.
Power supply from the 220 kV substations was completed in October 2014.

24.2 Other

There is no specific other relevant data and information not already included in other Sections.


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25	INTERPRETATION AND CONCLUSIONS

25.1 Mineral Resources

The increase in the gold oxide Resource is primarily due to the definition of extra resource to the west and at depth in Calaorco, as a direct result of the 2014 drilling program.
The Calaorco gold oxide Resource remains open along strike and should be further tested in 2015. The oxide mineralization is plunging to the North, along the Tilsa system.
Mineral Resources for sulfides have reduced due to the use of updated metal prices and other costs as determined by the sulfide project study.

25.2 Mining and Mineral Reserves

Oxide Mineral Reserves have increased due to the physical extension of the mineralization of the oxide deposit reflected in the new Mineral Resource estimates. The gold price of $ 1,200 per ounce was not changed from the previous estimates and only costs were updated based on the performance of the year 2014.
The mine is currently being efficiently operated by contractors using a conventional mining method of drill, blast, load, haul, and dump. There is ongoing work to improve efficiencies in all aspects of this cycle. The greatest operational cost in the cycle is haulage and the greatest focus is on achieving and maintaining efficiency in this area.
The updated oxide Mineral Reserve Statement comprises a total of four pits where one is main pit Calaorco currently in production and other three are located close to the main pit.
The oxide intrusive ore has been drilled and metallurgical and pilot tests performed. However, this will represent new material to be leached on the production pads. In addition, oxide material from the North and South Pits have higher copper grade. Rio Alto will need to introduce these two variables in the ore control procedures in order to ensure proper materials blending.
The Reserve Pit at 63 Mt is the starter pit which provides 10 years of steady mill feed at 18,000 t/d to the processing plant. Economics trade off study presented in Section 15 shows that this pit size represents the best discounted value for the project with lowest CAPEX. However, this pit is only a portion of a potentially larger pit from the 274 Mt resources.

25.3 Oxide Treatment

Oxide gold leaching process is currently operating as-planned. The infrastructure is in-place and operates efficiently. The performance of the gold recovery (Table 6.4-2) on the first three year of operations (from 2011 to 2014) falls within industry standards.


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25.4 Sulfide Treatment

Comminution testing indicates a low to moderate competency (SMC testing) and a moderate hardness (BWi) for the ARC and PHC ores, indicating that the ore is amenable to grinding in a conventional SAG/ball mill circuit.
KC ore, which is excluded from the current mine plan, indicates high competency and would require modifications to the comminution circuit for treatment
Initial investigations into the site hydrology indicate adequate water is available for the operation, although the Calaorco Oxide pit will need to be used to capture sufficient water for start-up.
Geotechnical investigations at the proposed plant site indicate very poor geotechnical conditions requiring considerable over-excavation and replacement fill to improve foundation conditions
The topography and footprint of the selected site results in an unusually large amount of earthworks excavation
The selected site for the process plant requires the relocation of existing site infrastructure including administration and accommodation facilities, truck work shop and warehouses

25.5 Project Infrastructure

25.5.1 General

Road access to the La Arena Project is very good and is being further improved. The sealed road between site and Trujillo is complete.
The La Arena mine site is now connected to the Peru grid power supply from September 2014.
All site offices and workshops are constructed and operating.
There is one fresh water borehole operational with a total of 3 L/s of water being produced.

25.5.2 Waste Dump Facilities

The geotechnical site investigation and laboratory testing for sulfide waste rock facility was adequate for feasibility level of foundation characterization; data from these was used to for stability analysis of the dumps under a variety of lift height and slope angle scenarios.
The ongoing kinetic geochemical testing for the ARD potential of the waste rock may lead to the requirement for an alternative encapsulation strategy. This data should be reviewed when available and incorporated at the detailed design stage.
The sulfide waste rock facility may require additional raises depending on the final volumes of the waste rock. This raise is likely to be insignificant and can potentially be accommodated and the final dump height could be raised to safely accommodate this material.


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The phase 2 expansion of the waste dump Nº2 will be located to the north side of phase 1 (to level 3330). The expansion of the phase 2 is outside of the approved modified EIA area for the project as well as being outside the limit of LASAC’s property boundary.
The Calaorco Pit expansion has changed from that shown in the approved modified EIA. The west side of the final Calaorco pit configuration is outside the approved modified EIA project area.
The relocated camp site is located outside of the approved modified EIA project limits but inside of the LASAC property limits.
The sewage treatment plant for ADR admin, is located outside of the approved modified EIA project limits but inside of the LASAC property limits
A large proportion of the low grade ore stockpile is located outside of the approved modified EIA project limits and the property limits of LASAC
The waste rock facility configuration outlined in Figure 18.11-3 (Section 18) is reasonable and has been determined based upon stability analysis using (SlopeW) software.

25.6 Contracts

Pursuant to the Gold Purchase Agreement, Rio Alto agreed to sell gold produced from the Calaorco breccia and Ethel breccia bodies of gold oxide mineralization located within the property encompassing the La Arena Project, as more particularly described in the July 31, 2010 technical report prepared by Coffey Mining Pty Ltd. (the "2010 Report"). Rio Alto had previously reported that it estimated this amount to be up to 634,000 ounces of gold and that it had delivered the estimated remaining ounces of gold to fulfill its obligations under the Gold Purchase Agreement.
The Gold Purchase Agreement makes reference to the 2010 Technical Report, which estimated that the ore bodies contained approximately 634,000 ounces of gold. However, as more particularly described in the 2015 Technical Report, the gold mineral resources and mineral reserves of the body of gold oxide mineralization within the Calaorco pit described in the 2010 Report have increased, while the Ethel pit has been exhausted through mining.
As a result of the increased gold mineral resources and mineral reserves, the contracting parties are working together to reach an agreed upon interpretation of the Gold Purchase Agreement. In the interim, Rio Alto has voluntarily re-commenced delivery of gold produced from the Calaorco pit to the counterparty in accordance with the terms and conditions contained in the Gold Purchase Agreement.

25.7 Overall

The La Arena oxide mine continues to exceed budget expectations due to positive grade variances between resource models and mining, and the definition of additional resources at the mine.
The La Arena sulfide project requires more work to determine the optimum size and scope of the project.


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26	RECOMMENDATIONS

26.1 Geology and Resources

Define the northern strike extensions to the current gold oxide Resource through ongoing RC infill and extensional drilling.
Continue to refine the geology model, particularly the Tilsa structural system (feeders) which may prove to be economic for small scale underground mining.
Continue the reconciliation process as a standard procedure.
Continue to explore in the district.
Update the sulfide resource with new domains based on detailed re-logging completed in 2014.

26.2 Mining

Geotechnical review of the new pit design must be undertaken to ensure pit wall stability. Any changes to the design must be implemented immediately.
The mine planning department should conduct a life of mine plan of the leach pad and waste dump facilities, to identify any potential issues with the scheduling and to ensure space availability for all material throughout the remainder of the mine life.
Review the mine production plan for the sulfide project and smooth the total rock moved per period. An opportunity to reduce the peaks and lows on the mine production schedule was identified which allow better equipment utilization.
An opportunity to reduce the size of the low grade stockpile exists with a detailed mine schedule.

26.3 Oxide Mineral Metallurgy

Pilot dump leach test works were carried out on intrusive and sandstone samples, mixed in different proportion. However the test could not determine the different gold extraction rates attributable to the sandstone and intrusive samples. Carry out additional leaching test on both blended and unblended material including a pilot dump leach, column and bottle roll test with the blended sample (sandstone plus intrusive) and column leach and bottle roll tests with the unblended samples to assess the relative gold extraction attributable to the sandstone and intrusive samples.
Pilot dump leach tests were carried out on an intrusive sample from one a zone of the deposit. Assess the spatial variability of the oxide deposit at different copper head grades, by conducting bottle roll and column leach tests on intrusive variability samples, then scale up the gold extraction using the pilot dump leach results.


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Metallurgical test work to date has only assessed leaching. Carry out carbon adsorption tests with pregnant liquor in order to determine carbon loading capacity when high soluble copper samples are being leached.

26.4 Sulfide Mineral Metallurgy

The latest metallurgical test work samples were selected from 6 metallurgical drill holes, of which only 3 are inside the current 63 Mt Feasibility Study pit. Additional variability flotation tests at optimized conditions should be conducted with new samples from drill holes samples inside the current pit design.
Cleaner gold losses for ARC and PHC are high at about 30%-40% at a regrind size of 25 m. Evaluate coarsening the target regrind size (e.g. from 25 m to 53 m) to determine if gold recoveries in the cleaner circuit can be improved.
Collector addition to the cleaner stage was 10 g/t Hostaflot 3403 of primary collector and 3 g/t Aerophine 3418A as secondary collector. Increase gold collector dosages (Aerophine 3418A) to the cleaner 1 and cleaner scavenger to determine if gold recovery in the cleaner circuit could be improved.
High copper concentrate grades were achieved at a pH of 10.5 in the cleaner circuit. Trials should be undertaken at lower pH levels to increase copper and gold recoveries i.e. pH levels of 9.5 and 10.
Higher copper and gold recoveries were achieved at the target copper concentrate grade of 22% Cu. Conduct cleaner and locked cycle tests using only two cleaner stages.
Diagnostic leach results indicate that around 15% of the gold in the cleaner scavenger tailings is cyanide soluble. Conduct leaching tests of pyrite concentrate and cleaner scavenger tailings using a regrinding stage to determine if economic recovery of gold from these streams is feasible.
High gold losses occur in the cleaner scavenger tailings. Future plant investigations may include gravity gold recovery tests on the cleaner scavenger tailings to determine if gold losses in tailings can be reduced and to produce a gravity concentrate that can be combined with the final copper concentrate.

26.5 Sulfide Process Plant

Investigations should be undertaken to identify an alternate plant site location that has better geotechnical characteristics and does not require relocation of existing infrastructure, this has the potential to significantly reduce the capital cost of the project.
The route survey identified key limitations for equipment departing Salaverry port as a width of 5 m, height of 4.5 m and weight of 48 t. Subsequent information has indicated that alternate transport routes, with a larger envelope, may be possible. This should be investigated as a possibility to reduce freight and fabrication costs.


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26.6 Infrastructure

Continue with the plan to relocate the national highway.
The following recommendations should be considered for the next phase of work based on the recent studies for waste rock disposal:

	o	Perform additional testing on the sulfide to better refine their physical and mechanical properties to further develop the stacking and the PAG waste rock encapsulation and leachate collection strategy;
	o	The ongoing kinetic geochemical testing for the ARD potential of the waste rock may lead to the requirement for an alternative encapsulation strategy. This data should be reviewed when available and incorporated at the detailed design stage;
	o	Perform a seepage model and fate of contaminate transport model in the next design phase to refine the leachate collection system and closure capping system; and
	o	The sulfide waste rock facility will require additional raises and potential lateral expansion to incorporate the Non-PAG and Low-PAG materials, which can be performed in the next phase.

The following recommendations should be considered for the next phase of work based on the recent studies for tailings deposition:

	o	Perform additional testing on the tailings to better refine their physical and mechanical properties, including consolidation with depth to further develop the deposition and water reclaim strategies;
	o	The geological mapping of the new pit configuration to determine if there are any additional design considerations need to be taken;
	o	The technical information currently available regarding pit hydrogeology and geotechnical characterization of tailings indicates a minimal affect the natural aquifer.
		However, the design provides conservatism, placing geotextile along the west wall of the pit, above the natural water level, to create an impermeable barrier against and possible migration of tailings through the wall composed of fractured sandstone. This need to be review and revised, if necessary, based on the new Calaorco pit configuration;
	o	Additional hydrogeological and geochemical studies need to be conducted in the Calaorco pit area based on the new configuration, i.e. deeper and bigger and with the potential interaction between the Calaorco pit and Sulfide pits to better understand any potential interaction;
	o	Update the water balance using only the 63Mt Calaorco pit option;
	o	Determine if an emergency spillway is required in the late stages on the Calaorco in pit filling; and
	o	Finally formalize the process requirement to remove the last years tailings production of sulfide for the closure cap

26.7 Social

Complete purchasing the land required for the gold oxide project, for the public road deviation, and continued land purchases for the sulfide project.
Continue to build on training programs for the local communities.


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26.8 Environmental

Complete water supply investigations and water balance calculations within final footprint and specifications for the La Arena project.
The site closure plan needs to be updated with the new details of the proposed Sulfide operation.
The environmental area of influence, with the exception of the associated areas of hydrology and aquatic biota (direct and indirect), will be need to be expanded according to the new project footprint and the results of updated air and sound predictive modeling for all the environmental components and areas of human interest (ground, physiography, air quality, sound, vibrations, land biology, landscape, archeology, etc.)


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27	REFERENCES

Andes Mining Research S.A.C.,“AMR Report on RIO Gold Dore Marketing and Gold Market”, March 2014

Anddes Asociados S.A.C., Memorando “Análisis de Estabilidad y Optimización del Relleno en Sector 8 (Tajo Ethel) del Pad Fase 3”. Proyecto CQA La Arena Fases 2 y 3, Noviembre 2013

Anddes Asociados S.A.C., “Diseno Conceptual Pad Fase 4 - Oxidos de Au con Alto Cu, Mina La Arena.” (rev. C), Diciembre 2013

ALS Metallurgy Project KM3262, “Preliminary Metallurgical Assessment of the La Arena Project”, August 2012

ALS Metallurgy Project KM3526, “Further Metallurgical Assessment of the La Arena Deposit - DRAFT”, November 2012.

AMEC Perú S.A. (2013). DOC-171706-043 “Memorando Análisis de Deformaciones - Tajo Ethel”, Proyecto La Arena.

AMEC Perú S.A., “Ingeniería de Detalle del Pad de Lixiviación - Fase 1B, Informe de Diseño HLP Fase 1B”, Proyecto La Arena“, Agosto 2011

AMEC Perú S.A., “Ingeniería de Detalle del Pad de Lixiviación - Fase2, Informe de Diseño HLP Fase 2”, Proyecto La Arena”, Enero 2011

Ausenco Vector, “Estudio Geotécnico para la Fase 2 del Pad de Lixiviación y Poza de Mayores Eventos”, Proyecto La Arena, 2011

C.M Orr - George, Orr and Associates, "La Arena Project: Summary Report on January 2014 Site Visit", February 2014

Coffey Mining, “La Arena Project, Peru - Technical Report (NI 43-1010) July 31, 2010”, 28 Oct 2010).

Cube Consulting, “La Arena Project, Peru – Geostatistical Study February, 2012” February 2012

Greg Corbett, “Comments on the Exploration Potential of La Arena and Regional Prospects Northern Peru”, December 2011

Heap Leaching Consulting S.A.C., “Informe de Preparacion de Muestras Y Conformacion de Muestras Compositos – La Arena”, Mayo del 2010

Jeffrey W. Hedenquist, “Observations on the La Arena epithermal Au mine and porphyry Cu-Au deposit, and prospects in the district, La Libertad, Peru”, October 2012

Kirk Mining Consultants, “La Arena Project, Peru – Technical Report (NI 43-101) September 30, 2011” 17 February 2012

Kirk Mining Consultants, "La Arena Project: Technical Report (NI 43-101)" January 2013 Mining Plus Peru SAC, “La Arena Project: Technical Report (NI-43-101)” March 2014. Vector Peru S.A.C., “Feasibility Study Report.” May 2010


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28	CERTIFICATES

Certificate of Qualified Person

La Arena Project, Peru, Technical Report, December 31^st2014, Rio Alto Mining Limited

1. I, Enrique Garay, have been a Rio Alto Mining employee since November 2010. My residential address is Calle Carlos Enrique Ferreyros No. 377, San Isidro, Lima 27, Peru.

2. I am a member of the Australian Institute of Geoscientists (“MAIG”).I hold a Bachelor's Degree in Science with a major in Geology from the National Engineering University, Lima and a MSc in Mineral Exploration from Queens University, Canada.

3. I am a practising geologist for over 23 years in the precious and base metal resource industry with a focus on both exploration and mine geology. I have been previously employed by several mining companies including Barrick Gold Corporation, Hochschild Mining PLC, Trafigura and Consorcio Minero Horizonte S.A. From 1996 to 2004 I contributed to the resource definition work at Barrick Gold Corporation's, multi-million ounce Pierina Gold Mine located in Peru and I was the project's Chief Mine Geologist.

4. I have frequently visited the property that is the subject of this report, since November 2010.

5. I am responsible for Sections 7, 8, 9 and 10 of this report.

6. I am co responsible for Sections 1, 25 and 26 of this report.

7. I am not independent of Rio Alto Mining Limited as independence is described in Section 1.5 of NI 43-101.

8. I hereby consent to the use of this report and my name in the preparation of documents for a public filing including a prospectus, an annual information filing, brokered or non-brokered financing(s), or for the submission to any Provincial or Federal regulatory authority.

9. I have read National Instrument 43-101 and Form 43-101F1 and, by reason of education and past relevant work experience, I fulfil the requirements to be a “Qualified Person” for the purposes of NI 43-101. This technical report has been prepared in compliance with National Instrument 43-101 and Form 43-101F1.

10. At the effective date, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated 27^thday of February 2015 at Lima, Peru.

[signed]



Enrique Garay, M Sc. (MAIG)
Vice President Geology
Rio Alto Mining Limited


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

Certificate of Qualified Person

La Arena Project, Peru, Technical Report, December 31^st2014, Rio Alto Mining Limited

1.	I, Tim Williams, am the Vice President of Operations at Rio Alto Mining.

2.	I am a Fellow of the Australasian Institute of Mining and Metallurgy (AusIMM). I graduated from Curtin University, Western Australian School of Mines, Western Australia with a B. Eng (Geology) degree in 1991. I also have a Masters Eng. Science and a Grad Dip Mining from Curtin University, WASM.

3.	I have practiced my profession continuously since 1991. I have been directly involved in the mining operations, on three continents in a variety of mineral commodities and mining methods.

4.	I have visited the property regularly and on numerous occasions during 2014.

5.	I am responsible for Sections 4, 5, 15, 16, 17-1, 18 (except 18-10 and 18-11), 19, 20, 21 and 22, of this report.

6.	I am co responsible for Sections 1, 25 and 26 of this report.

7.	I am not independent of Rio Alto Mining Limited as independence is described in Section 1.5 of NI 43-101.

8.	I hereby consent to the use of my name in the preparation of documents for a prospectus, annual information filing, initial public offering, brokered or non-brokered financing(s), for the submission to any Provincial or Federal regulatory authority.

9.	I have read National Instrument 43-101 and Form 43-101F1 and, by reason of education and past relevant work experience, I fulfil the requirements to be a “Qualified Person” for the purposes of NI 43-101. This technical report has been prepared in compliance with National Instrument 43- 101 and Form 43-101F1.

10.	At the effective date, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated 27^thday of February 2015 at Lima, Peru.

[signed]



Tim Williams,
Vice President Operation
Rio Alto Mining Limited


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Certificate of Qualified Person

La Arena Project, Peru, Technical Report, December 31^st2014, Rio Alto Mining Limited

1. I, Ian Dreyer, am the Corporate Development Geologist of Rio Alto Mining.

2. I am a Charted Professional of the Australasian Institute of Mining and Metallurgy (AusIMM). I graduated from Curtin University, Perth, Western Australia with a B. App.Sc (Geology) degree in 1982.

3. I have practiced my profession continuously since 1988. I have been directly involved in the mining, exploration and evaluation on three continents in a variety of mineral commodities.

4. I have visited the property regularly and on numerous occasions during 2014.

5. I am responsible for Sections 11, 12 and 14 of this report.

6. I am co responsible for Sections 1, 25 and 26 of this report.

7. I am not independent of Rio Alto Mining Limited as independence is described by Section 1.5 of NI 43-101.

8. I hereby consent to the use of my name in the preparation of documents for a prospectus, annual information filing, initial public offering, brokered or non-brokered financing(s), for the submission to any Provincial or Federal regulatory authority.

Dated 27^thday of February 2015 at Lima, Peru.

[signed]



Ian Dreyer BSc Geology, MAusIMM(CP)
Corporate Development Geologist
Rio Alto Mining Limited


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

Certificate of Qualified Person

La Arena Project, Peru, Technical Report, December 31^st2014, Rio Alto Mining Limited

1. I Fernando Angeles am currently employed as Senior Mining Consultant by Mining Plus Peru SAC., 1196 Javier Prado Este Avenue, San Isidro, Lima, Peru.

2. I hold bachelor’s and master’s degrees in mining engineering from the Pontifical Catholic University of Peru and The University of British Columbia in Canada, respectively.

3. I am registered with The Association of Professional Engineers and Geoscientists of British Columbia (APEGBC) as a Professional Engineer P.Eng (License No. 178264)

4. I have practiced my profession for 10 years in the areas of mining operation, mineral project, reserves estimates and mine economic for precious and base metals minerals.

5. I have visited La Arena property that is the subject of this report from September 8^thto September 11^thof 2014.

6. I am responsible for Section 2, 3, 6, 23 and 24 and co-responsible for portions of Sections 1, 25, and 26 of this report.

7. I have no prior involvement with the property that is subject of Technical report. I have no controlling or monetary interest involving Rio Alto Mining Limited or the property.

8. I am independent of Rio Alto Mining Limited., applying all of the tests in section 1.5 of NI 43-101.

9. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

10. As of the effective date of the Technical Report, to the best of my knowledge and information the Technical Report contains all scientific and technical information required to be disclosed to make the report not misleading.

Dated 27^thday of February 2015 at Lima, Peru.

[signed]



Mr. Fernando Angeles P.Eng.,
Senior Mining Consultant,
Mining Plus Peru SAC.


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Certificate of Qualified Person

La Arena Project, Peru, Technical Report, December 31^st2014, Rio Alto Mining Limited

1. I, Greg Lane am Chief Technical Officer, with Ausenco Ltd, 144 Montague Road, South Brisbane, Queensland, Australia.

2. I am a Fellow of Australasian Institute of Mining and Metallurgy (AusIMM). I graduated from the University of Tasmania, Australia, with a M. Sc. in1986.

3. I have worked as a process engineer in the minerals industry for over 28 years. I have been directly involved in the mining, exploration and evaluation of mineral properties internationally for precious and base metals.

4. I visited the property in December 2014.

5. I am responsible for Sections 13 and 17.2 of this report.

7. I am independent of Rio Alto Mining Limited as independence is described by Section 1.5 of NI 43-101. I have not received, nor do I expect to receive, any interest, directly or indirectly, in Rio Alto Mining Limited.

Dated 27^thday of February 2015 at Brisbane, Australia.

[signed]



Greg Lane
Chief Technical Officer
Ausenco


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	MP PERU SAC	Technical Report NI 43-101, La Arena Project, Peru

Certificate of Qualified Person

La Arena Project, Peru, Technical Report, December 31^st2014, Rio Alto Mining Limited

1. I, Scott Elfen P.E. am the Global Lead Geotechnical Services with Ausenco Engineering Canada Ltd, 855 Homer Street, Vancouver, British Columbia, V6B 2W2, Canada.

2. I graduated with a Bachelor of Science degree in Civil Engineering from the University of California, Davis in 1991.

3. I am a Registered Civil Engineer in the State of California by exam since 1996 (No. C56527). I am also a member of the American Society of Civil Engineers (ASCE) and Society for Mining, Metallurgy & Exploration (SME).

4. I have worked as an engineer in the minerals industry for over 20 years. I have been directly involved in the mining and evaluation of mineral properties internationally for precious and base metals.

5. I visited the La Arena property that is the subject of this report on August 20^thof 2014.

6. I am responsible for Sections 18.10, 18.11, and co-responsible for portions of Sections 1, 25, and 26 of this report.

7. I am independent of Rio Alto Mining Limited as independence is described by Section 1.5 of NI43-101. I have not received, nor do I expect to receive, any interest, directly or indirectly, in Rio Alto Mining Limited.

Dated 27^thday of February 2015 at Vancouver, Canada.

[signed]



Scott Elfen
Global Lead Geotechnical Services
Ausenco Engineering Canada Ltd


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Rio Alto Mining Inactive 6-KRio Alto Files Ni 43-101 Technical Report on the La Arena Project