EX-99.1 2 ex99_1.htm EXHIBIT 99.1

Exhibit 99.1

Report to:

Form 43 - 101F1 Technical Report:

Tailings Retreatment - Process Options

Document No. 0551920100-REP-R0002-01

WARDROP

	Report to:
	FORM 43 - 101F1 TECHNICAL REPORT: TAILINGS RETREATMENT - PROCESS OPTIONS
	March 2006
	Prepared by		Date
		Andre de Ruijter
	Reviewed by		Date
		Rick Alexander
	Authorized by		Date
		Ron Hall
	WARDROP
	905-1130 West Pender Street, Vancouver, British Columbia V6E 4A4 Phone: 604-408-3788 Fax: 604-408-3722 E-mail: vancouver@wardrop.com

WARDROP

REVISION HISTORY

REV. NO	ISSUE DATE	PREPARED BY AND DATE	REVIEWED BY AND DATE	APPROVED BY AND DATE	DESCRIPTION OF REVISION
00	March 9, 2006	Andre de Ruijter	Rick Alexander	Ron Hall	Draft issue for client review
01	March 31, 2006	Andre de Ruijter	Rick Alexander	Ron Hall	Final issue to client

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TABLE OF CONTENTS

1.0	SUMMARY	1
2.0	INTRODUCTION	2
3.0	RELIANCE ON OTHER EXPERTS	3
4.0	PROPERTY DESCRIPTION AND LOCATION	4
5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY	5
6.0	HISTORY	6
7.0	GEOLOGICAL SETTING	7
8.0	DEPOSIT TYPES	8
9.0	MINERALIZATION	9
	9.1  MMI 2003 SAMPLES	9
	9.2  MMI 2004 SAMPLES	9
10.0	EXPLORATION	10
11.0	DRILLING	11
12.0	SAMPLING METHOD AND APPROACH	12
13.0	SAMPLE PREPARATION, ANALYSES AND SECURITY	13
14.0	DATA VERIFICATION	14
15.0	ADJACENT PROPERTIES	15
16.0	MINERAL PROCESSING AND METALLURGICAL TESTING	16
	16.1  A METALLURGICAL REVIEW	16
	16.1.1  A HISTORICAL EVALUATION OF THE OXIDE TAILINGS	16
	16.1.2  HISTORICAL METALLURGICAL TEST RESULTS	16
	16.2  THE MMI TECHNICAL REPORT - PREAMBLE	17
	16.2.1  INTRODUCTION TO THE MMI 2003 METALLURGICAL TEST PROGRAM	18
	16.2.2  INTRODUCTION TO THE MMI 2004 METALLURGICAL TEST PROGRAM	19
	16.2.3  EVALUATION AND REVIEW OF METALLURGICAL TESTS	20

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	16.3  PROCESS ENGINEERING EVALUATION	33
	16.3.1  GRAVITY CONCENTRATION	33
	16.3.2  FLOTATION	33
	16.3.3  CYANIDE LEACHING	34
	16.3.4  COLUMN LEACH TEST	35
	16.3.5  PRECIOUS METAL RECOVERY	35
	16.3.6  POTENTIAL REVENUE ESTIMATION	36
	16.3.7  HEAP LEACH OPTION - ASSUMPTIONS	37
	16.4  PROCESS FLOWSHEET	41
	16.4.1  INTRODUCTION	41
	16.4.2  PROCESS DESCRIPTION	41
	16.4.3  PROCESS DESIGN CRITERIA	43
	16.4.4  CAPITAL AND OPERATING COST ESTIMATES	43
17.0	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES	49
18.0	OTHER RELEVANT DATA AND INFORMATION	50
	18.1  FINANCIAL MODEL	50
	18.2  COMMENTS REGARDING THE COST ESTIMATES	54
19.0	INTERPRETATION AND CONCLUSIONS	55
20.0	RECOMMENDATIONS	56
21.0	REFERENCES	57
22.0	DATE AND SIGNATURE PAGE	58
23.0	ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES	59
24.0	ILLUSTRATIONS	60

List of Appendices

Appendix A  General Arrangements, Drawings and Sections

Appendix B  Process Flowsheet

Appendix C  Electrical Single Line Diagram

Appendix D  Process Design Criteria

Appendix E  Capital Costs

Appendix F  Slusher System Description & Design Calculations

Appendix G  Preliminary Economic Evaluation

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Appendix H  “A Tailings Resource” by Bryan Slim

Appendix I  PRA Tailings Report

Appendix J  Cia Minera 1990 Sampling Program (from Avino Mines)

LIST OF TABLES

Table 16.1	Oxide Tailings Dam Data	16
Table 16.2	Cyanidation Test Results	17
Table 16.3	Flotation Test Results	17
Table 16.4	Test Procedures - MMI 2003 Test Program	18
Table 16.5	Test Procedures - MMI 2004 Test Program	20
Table 16.6	Moisture Content of Samples	22
Table 16.7	Head Assays	24
Table 16.8	Bulk Density and Specific Gravity	25
Table 16.9	Summary of Results of Gravity Concentration Tests	27
Table 16.10	Summary of Results of Flotation Tests	29
Table 16.11	Summary of Results of PRA Cyanidation Tests	30
Table 16.12	Summary of Cyanidation Test Results used by the MMI Reports	31
Table 16.13	Summary of Results of Column Leach Tests	32
Table 16.14	Cyanide Leaching Parameters	34
Table 16.15	Inherent Value of Oxide Tailings	36
Table 16.16	Summary of Cost Estimates - Four-year Treatment	37
Table 16.17	Summary of Cost Estimates - Two-year Treatment	38
Table 16.18	Estimated Inherent Value of Sulphide Tailings	40
Table 16.19	Capital Cost Estimate - Summary	44
Table 16.20	Estimated Cost of Reagents	46
Table 16.21	Plant Staff Complement And Labour Cost - 3-Shift Basis	47
Table 16.22	Estimated Operating Costs	48
Table 17.1	Oxide Tailings Dam Data	49
Table 18.1	Potential Net Present Value (NPV)	52
Table 18.2	Potential Internal Rate of Return (IRR)	53
Table 19.1	Summary	55

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1.0   SUMMARY

Avino Silver and Gold Mines Ltd requested Wardrop Engineering Inc [Wardrop] to produce a scoping study based primarily on the MineStart Inc. [MMI] Technical Report dated October 2005 addressing the Companie Minera Mexicana de Avino, SA de CV property in the Durango mining district of Mexico. Avino Mine wished to determine if there was sufficient financial justification to develop the property by way of processing some, or all, of the tailings material deposited over the years that the mine was in operation. The principal asset of the company is the Avino Mine where about 497 tonnes of silver, 3 tonnes of gold and 11,000 tonnes of copper as well as an apparently undocumented amount of lead were produced between 1976 and 2001.

The project is a re-treatment of oxide tailings from previous milling operations. The tailings grade averaged 95.5 g/t silver and 0.53 g/t gold, with the particle size of the tailings material ranging from minus 37 microns to plus 210 microns. The valuable minerals found in the oxide zone during the mining operations have been reported to be argentite, bromargyrite, native silver, and native gold. Other potentially economic minerals included chalcopyrite, chalcocite, galena, sphalerite, bornite, and native copper. The gangue minerals included hematite, chlorite, quartz, barite, pyrite, arsenopyrite and pyrrhotite. Malachite, anglesite and limonite were common accessory minerals in the quartz zones of the weathered parts of the oxide material.

Wardrop reviewed all available data and developed different mining and processing flowsheets during the evaluation of the property. Much of the previous data could not be validated and, therefore, could not be used in the review. The MMI Technical Report indicated that an economic evaluation of the tailings deposit could not be completed at this stage since the deposit could only be classified as an inferred resource. MMI had arrived at this conclusion based on their acknowledgement that their study was limiting with respect to the sampling of the tailings dam and incomplete metallurgical characterisation.

However, a conceptual financial model was developed by WARDROP using the estimated grade values and testwork results as reported by MMI and Process Research Associates, who conducted the metallurgical tests.

Wardrop investigated gravity separation, flotation, cyanide leach, carbon-in-pulp [CIP], and heap leach processing options. Using the recoveries and process conditions resulting from these tests, the capital costs to construct a processing plant using selected process options were developed while the operating costs associated with each option were determined and a financial model compiled. A heap leach operation with a four-year minelife indicated the best financial alternative.

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2.0   INTRODUCTION

As part of the completion of the Avino Mines tailings recovery study, a review and validation of the metallurgical test program previously undertaken by others, was conducted. The test results obtained were also assessed and the conclusions drawn reviewed and validated. This information was subsequently used to establish process plant design criteria for the flowsheet and subsequent conceptual mass and water balances of the selected recovery process, namely heap leach operation with a project duration of 4-years. For the heap leach recovery process selected to treat the tailings material, a mechanical equipment list was compiled. The infrastructure design was outlined and included the sizing of the main items of equipment required for the treatment of the tailings dam material. Some financial information was collected to compile the capital cost estimates and equipment quotations and selection, to the level of plus/minus 35%. A financial model was drawn up incorporating the main variables such as metal price and recoveries and the estimated operating costs.

The following constitutes the review of the test data obtained and the validation of the processes selected for treating the tailings dam material.

The source of reference for this review is almost exclusively the following report since it is intended to form the basis for use by Avino Mines for any further financial dealings. This report, titled “Avino Silver and Gold Mines Ltd, A Tailings Resource”, dated July 2005, and submitted as a NI 43-101 document by MineStart Management Inc [MMI] was used as the source of information. This document was subsequently revised and re-issued during October 2005. The MMI report based their conclusions on the results obtained from the metallurgical testwork program conducted by Process Research Associates [PRA] but directed by MMI. The PRA report reviewed was titled “Metallurgical Test Work on Avino Tailings, Durango, Mexico”, Project No. 0406407, and dated 28 March 2005. The metallurgical results obtained by PRA were reviewed and assessed and these are discussed in the following sections of this report. The implications of the reviewed results and the effect on the conclusions reached in the MMI report are discussed in this report.

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3.0   RELIANCE ON OTHER EXPERTS

See the report “A Tailings Resource” dated October 2005 by Bryan Slim of MineStart Management Inc. and similar reports from the same author in Appendices H and I.

Wardrop Engineering Inc. has relied upon information contained in the above-mentioned report, “A Tailings Resource”. Wardrop has used the information in this report under the assumption that it has been prepared by Qualified Persons.

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4.0   PROPERTY DESCRIPTION AND LOCATION

See the report “A Tailings Resource” by Bryan Slim in Appendix H.

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 5.0     ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

The minesite, which lies between the towns of Panuco de Coronado and San Jose de Avino, is at an elevation of about 2,200 metres. Relief is estimated at 100 metres. The vegetation is typically sparse.

The climate is temperate and arid/semi-arid. Mean monthly rainfall range is from over 120 mm per month in July and August to less than 20 mm per month in March and April.

As a result of the extensive history of previous mining activity at Avino Mine since 1976 and prior to 1976, there is a complete infrastructure for the area. The national grid supplies power from a line capacity quoted at 4MW.

(See the report “A Tailings Resource” by Bryan Slim in Appendix H.)

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6.0   HISTORY

Silver and gold was discovered in 1555 by Juan de Tolosa, a member of the Spanish Army, at what eventually developed into the Avino Mine. Mining operations commenced seven years later. Miners were encouraged to settle in this area and what would become the Durango district, thereby providing protection for the settled families from the indigenous Indian inhabitants. The Avino Mine is considered to be first operating mine in the Nueva Vizcaya area, later re-named Durango. In 1880, the deposits were merged into a larger scale mining operation with new technology and equipment, and named Avino Mines Ltd. Operations were abandoned in 1912 as a result of the Mexican Revolution.

In 1968, the Ysita family and Avino Mines and Resources Ltd of Vancouver, British Columbia, Canada, jointly formed the Companie Minera Mexicana de Avino, SA de CV, which acquired the property rights and the mineral rights, including the mine. The new company implemented an exploration company resulting in the re-opening of the mine with limited open-pit production commencing in 1970. A lead flotation concentrate was produced with silver and gold content credit received from the lead smelter. From 1974, open-pit production was continuous with a total of approximately 2 million tonnes of oxidised ore treated until 1993 when the underground mining of sulphide ore commenced. A copper concentrate was then recovered by flotation with silver and gold credit received from the copper smelter. The mine was closed in November 2001 as a result of the closure of the copper smelter treating the Avino Mine concentrates. About 3 million tonnes of sulphide ore had been treated during the period from 1993 to the time of the mine closure in 2001.

Avino Silver and Gold Mines Ltd recently increased its ownership of Cia Minera Mexicana de Avino, SA de CV, from its previous holding of 49%, held since 1968, to 100%. The company now intends to resume exploration and possibly re-start mining at the Avino Mine, while recovering metals from the existing tailings deposited over the years that the mine was operating. The principal asset of the company is the Avino Mine where about 497 tonnes of silver, 3 tonnes of gold and 11,000 tonnes of copper, as well as an apparently un-documented amount of lead, have been produced from 1976 to 2001. No mention has been made of zinc being recovered or receiving credit from the smelters. No metal production records appear to exist for the period 1970 to 1975.

A metallurgical test program was designed and implemented by MMI in 2003 and 2004, and the results were used in the Technical Report submitted to the Canadian Securities Regulatory Authorities [CSRA] for Avino Silver and Gold Mines Ltd. This report was originally titled “Preliminary Feasibility” and was dated May 2005, and referred to the oxide tailings as an “indicated resource”. A subsequent version of the MMI report appeared in July 2005 titled “A Tailings Resource”. Several deficiencies were noted by the CSRA and the report was returned to MMI for clarification. These deficiencies were subsequently addressed by MMI and the report dated October 2005, and in its final version as reviewed by the CSRA, refers to the oxide tailings as an “inferred resource”.

(See the report “A Tailings Resource” by Bryan Slim in Appendix H.)

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7.0   GEOLOGICAL SETTING

The orebody is epithermal and made up of veins and stockwork structures as reported in the MMI May 2005 report. The structure was normally weathered and leached in the upper section as a result of contact with the atmospheric waters. The oxide tailings material under consideration in this study is primarily from this source.

(See the report “A Tailings Resource” by Bryan Slim in Appendix H.)

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8.0   DEPOSIT TYPES

A study by MineStart Management Inc [MMI] investigated the potential for recovering the contained silver and gold from the Avino Silver and Gold Mines tailings deposit. This MMI study identified an inferred resource of 2 million tonnes of oxide tailings with a grade of 95.5 g/t silver and 0.53 g/t gold. The sulphide tailings were excluded from this MMI study, which delineated only the oxide tailings deposit as the [inferred] resource.

(See the report “A Tailings Resource” by Bryan Slim in Appendix H.)

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9.0   MINERALIZATION

Avino Silver and Gold Mines Ltd supplied the information in this section.

The valuable minerals found during the period of mining of the oxide zone were reported to be argentite, bromargyrite, chalcopyrite, chalcocite, galena, sphalerite, bornite, native silver, gold and native copper. The gangue minerals include hematite, chlorite, quartz, barite, pyrite, arsenopyrite and pyrrhotite. Malachite, anglesite and limonite are common in the quartz zones of the weathered parts of the oxide material. Although the above description was extracted from the MMI Technical Report, the reference literature citing the origin of this information was not recorded. It is anticipated that the minerals listed above would be present in the oxide section of the tailings deposit.

Two specific mineralogical assessments were conducted by MMI on samples arising from the MMI metallurgical testwork programs, one during 2003 and the other in 2004. The results and implications of these findings are discussed below.

9.1       MMI 2003 SAMPLES.

A flotation test tailings sample was submitted for mineralogical evaluation. No silver minerals were identified, and no gold particles were observed. Therefore no silver and/or gold associations with other minerals were observed. The report concluded that the silver present in the sample was probably occluded in, or adsorbed onto, secondary iron or silver minerals, such as argentojarosite. The minerals identified in the study included those already listed above.

9.2       MMI 2004 SAMPLES.

MMI requested that a mineralogical examination was to be conducted on sample material taken from specific sample bags shortly after these samples had been received by PRA. The rationale for selecting these samples was not disclosed to the PRA staff. The results of the investigation were not made available to PRA either, even though this may have assisted in the understanding of the nature of the material being tested, and assisted with the design of the metallurgical test program. The results of this examination were not disclosed in the MMI Technical Report.

(See “A Tailings Resource” by Bryan Slim in Appendix H.)

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10.0 EXPLORATION

This section is not applicable.

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11.0 DRILLING

This section is not applicable.

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12.0 SAMPLING METHOD AND APPROACH

The sampling method and approach adopted by Bryan Slim/MMI is described in the MMI report “A Tailings Resource” [Reference #7]. Essentially this incorporated the following steps.

1. A backhoe was used to excavate sample pits to a depth of 4 meters. Hand samples were taken at 1 meter vertical increments from the sidewalls of each pit.

2. The sample mass collected from each sampling point generally amounted to between 2 and 5 kg.

3. The MMI sampling program was ostensibly based on the 1990 Cia Minera sampling program. The Cia Minera sampling program was based on the results of 34 holes drilled to bedrock/soil level and generated 461 samples, which were submitted for silver and gold assay. The MMI sampling program actually excavated 14 sample pits to a depth of 4 meters and generated 86 samples.

Wardrop did not independently verify nor compare the results of the sampling programs.

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  13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

The samples collected by MMI from the Avino Mines tailings dam in 2004 were airfreighted to Process Research Associates [PRA] in Vancouver, B.C, from Durango, Mexico. The samples had been bagged and sealed with identification tags attached, as soon as each had been taken. This was done under MMI supervision. The samples were unpacked at the PRA facility in the presence of MMI personnel. The samples were then allotted new identification numbers by Bryan Slim of MMI, and were subsequently un-bagged and dried. The dry samples were individually mixed and blended, and then split into four one-quarter fractions as directed by MMI. One fraction was used to determine the head grade assay, while another quarter was used to create composite samples used for the subsequent metallurgical testwork program. MMI instructions were followed with the compositing of the samples, and the testwork program. Excess sample was archived for future testwork or analyses.

The analytical techniques employed during the testwork program were the following. Initially the standard fire assay [with atomic absorption spectrophotometric finish] was used for the silver analyses. However, this method is not very accurate for silver values < 100 g/t Ag. Subsequently, the Inductively Coupled Plasma Spectroscopy, which uses multi-acid digestion, was used for silver [ICP-M]. This method also resulted in analyses being obtained for other elements of interest, eg copper, zinc, lead, etc.

The standard fire assay method was used for gold analyses. Cyanide and lime concentrations were done using standard titrametric methods. Total sulphur was measured using a standard Leco furnace, and sulphide sulphur assays were done using the standard wet chemical gravimetric analysis.

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14.0 DATA VERIFICATION

MMI substantiated the quality of the results of the various assays obtained by utilising the laboratory International Plasma Labs Ltd [IPL] in Vancouver, B.C. This laboratory utilises a standard system of duplicate samples, standards and blanks for all their analytical determinations. Also, IPL are ISO9001:2000 registered.

As part of the initial set up of the project, the results of different analytical methods for silver were compared. This indicated that the fire assay method was not suitable for analysing solids samples for silver with grades below 100 g/t Ag. The AA method was also checked for reliability, but deemed less accurate when compared with the ICP-M method. The ICP-M method was subsequently used for silver determinations from solids samples throughout the metallurgical testwork program. Another unidentified analytical laboratory was also reportedly utilised for comparison of silver assays from specified samples submitted by MMI. However, the outcome of this exercise was not made available by MMI to PRA staff, and the results were not published in the MMI report “A Tailings Resource”.

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15.0 ADJACENT PROPERTIES

This section is not applicable.

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  16.0 MINERAL PROCESSING AND METALLURGICAL TESTING

This present section will review and validate the metallurgical characterisation and study which was generated by MMI, and review all the previous test reports and comments to establish the treatment process and the design criteria required for the recovering of silver and gold from the tailings resource.

16.1     A METALLURGICAL REVIEW

As mentioned above, the MMI report titled “Avino Silver and Gold Mines Ltd, A Tailings Resource”, dated July 2005 [Reference #7], used the metallurgical results obtained and conclusions drawn by PRA [Reference #1 and #3] in their NI 43-101 document. The revised and final MMI document was dated October 2005.

This study will initially review the available historical information and this will be followed by a review of the more comprehensive test program conducted by MMI at PRA, in 2003 and particularly in 2004.

16.1.1   A Historical Evaluation of the Oxide Tailings

In 1990, Cia Minera sampled the oxide tailings dam and calculated the tonnage and overall grade to be as given in the table below. No metallurgical characterisation tests were conducted on these samples. In their report, MMI have concluded that the Cia Minera data constitutes a reasonable estimation of this tailings material as a resource, and have reported the following tonnages and metal grades in Table 16.1. These values are based on their calculations. [See also Section 17].

Table 16.1 Oxide Tailings Dam Data

Source	Tonnes, t	Assays, g/t		Bulk Density, g/cm3
		Ag	Au
Cia Minera, 1990	2,092,178	93.0	0.50	1.605
MMI, 2005	2,091,074	95.5	0.53	1.605

16.1.2   Historical Metallurgical Test Results

A number of metallurgical evaluations have been made on various samples from the oxide tailings dam, according to the MMI report [Reference #2]. Apparently the first cyanidation tests were conducted during 1982, and this was followed by further tests performed over the years. The summarised cyanidation test results are reproduced in the following table [Table 16.2] taken from the MMI report, while the reported flotation test results are given in Table 16.3. The MMI results obtained from the testwork program initiated by MMI in 2003 and 2004 and which have been reported in the MMI Technical Report, are included in the table for purposes of comparison. The results will be discussed in greater detail later in the report.

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Table 16.2 Cyanidation Test Results

Author	Date of Test	Extraction, %		Leaching Time, hr	Particle Size, Microns
		Ag	Au
Denver Equipment	1982	69.3	66.7	24	66.6% < 149
Penoles	1987	78.3	88.9	24	87% < 74
Maja	1990	85.9	80.9	24	100%<105
Chryssoulis	1990	85.9	80.9	24	no data
Rosales	1996	83.9	76.9	23	75%< 74
MMI	2003	77. 1	71.4	24	86%< 74
MMI	2003	88.8	88.4	48	86% < 74

No details have been provided regarding the location nor the manner in which the samples for any of the above tests were taken, why these particular samples were taken, the test parameters employed, the assay techniques used, etc. The first set of results for tests conducted on MMI samples from the 2003 sampling campaign indicate a silver extraction of 77.1% and gold extraction 71.4%. However, these results cannot be verified since the origin of this set of numbers as quoted in the MMI Technical Report is not known. The second set of results was reported in the PRA report titled “Flotation and Cyanidation Scoping Tests and Specific Gravity”, Project No. 0302303, dated 28 March 2003 [Reference #1]. Considered in general terms, it would appear as if the cyanidation test results over the indicated period of time are reasonably consistent. However, no specific conclusions should be drawn since nothing is known about the head grades of the samples, nor the samples used, nor the test and assay procedures used at the time that these tests were conducted.

Table 16.3 Flotation Test Results

Author	Date of Test	Recovery, %		Particle Size, microns
		Ag	Au
Penoles	1987	60.2	47.1	87% < 74
Rosales	1996	69.4	66.9	75% < 74

The flotation results vary widely for similar particle sizes with recoveries ranging from 60 to 69% for silver and 47 to 67% for gold. However, since the test details of these reported cyanidation and flotation tests are unknown, any further discussion will be meaningless.

16.2     THE MMI TECHNICAL REPORT - PREAMBLE

MineStart Management Inc [MMI] was commissioned by Avino Silver and Gold Mines Ltd to produce a document that was NI 43-101 compliant with respect to detailing the indicated oxide tailings resource [subsequently referred to as an inferred resource] and to define the metallurgical characterisation and assay results for this material. The proposed economic processing of this tailings material could then be used to form the financial basis for restarting the mine. The first report prepared by MMI was titled “Tailings Evaluation” and was dated November 2003 [Reference #2]. Two further reports were subsequently prepared by MMI. One was titled “Preliminary Feasibility” and was dated May 2005 [Reference #4], while the second report was titled “Tailings Evaluation” and was also dated May 2005 [Reference #5]. This report [Reference #5] was subsequently revised and re-titled “A Tailings Resource” and dated July 2005 [Reference #6]. This July 2005 MMI report [Reference #6] was reviewed by the CSRA and returned to MMI for revision. The revised MMI report was subsequently re-issued as “A Tailings Resource” and dated October 2005 [Reference #7], and this was re-submitted for reviewing by the CSRA. This revised document [Reference #7] was titled “A Tailings Resource”, dated October 2005, and was produced for Avino Silver and Gold Mines Ltd, Cia Minera Mexicana, Durango, Mexico, by Bryan Slim, of MineStart Management Inc, North Vancouver, B.C., Canada. The document was submitted as a Technical Report to the CSRA.

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A review of the assaying methods and the metallurgical testwork, as directed by MMI and conducted by PRA, will now be carried out. This review will detail the processing techniques applied and comment on the test procedures used and validate the results obtained. It will then use the test data obtained to develop the proposed process flowsheet and the plant design criteria to establish the treatment process for the economic recovery of silver and gold from the Avino Mines tailings resource.

Two sets of test programs were directed by MMI and conducted at PRA. One was conducted during 2003, for which no sample origin can be determined [Reference #1], and the other, more detailed test program, was conducted during 2004 [Reference #3]. The 2004 testwork and assaying program was designed and supervised by MMI. It was conducted on samples collected from the tailings dam by MMI during 2004, while also using the results from the preliminary metallurgical scoping tests done during 2003 as a guide. PRA staff at their facilities in Vancouver, B.C., conducted all the testwork from both MMI test programs.

16.2.1   INTRODUCTION TO THE MMI 2003 METALLURGICAL TEST PROGRAM

The 2003 test program consisted of the following tests as summarised in Table 16.4. The cyanidation extraction results obtained were used in a preliminary report by MMI titled “Tailings Evaluation” dated November 2003 [Reference #2]. MMI considered using a 2000 tonnes per day vat leaching process to recover the silver and gold from the oxide tailings. However, this treatment process option was revised when the results of the 2004 test program became available.

Table 16.4 Test Procedures - MMI 2003 Test Program

Process/ Procedure	Details of Test	Sample Identity
Sample Preparation	No details documented	Sample L and Sample U
Head Assays	Fire assays, AA, and ICP multi-acid	Composite of L and U
Specific Gravity	Standard pycnometer test	Composite of L and U and Cyanide Residue
Cyanidation Leach	P80 = 68 microns; 40% solids; pH 10.5; 1.0g/l NaCN; 48 hours; dO2 > 7.9 mg/l 0.4 kg sample	Composite of L and U
Flotation	Rougher and 2 scavenger stages; P80 = 85 microns; 35% solids; pH 5.5; PAX & A208 with MIBC; 1kg sample	Composite of L and U
Mineralogical	Examination of flotation tailings	Composite of L and U

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The exact origin of Sample L and Sample U is not known, and does not appear to have been documented. The manner that each of the samples was collected by MMI has apparently also not been documented. The size of both samples, namely 0.8 kg for Sample L and 0.9 kg for Sample U, is small and its representation is questioned. Also, there appears to be no documentation relating to the arrival and receiving of these samples at PRA. There is no Receiving Log in the PRA Report No. 0302303 [Reference #1]. Also, no assay certificates have been recovered to date. Even though the above tests were considered to be scoping tests only, the results cannot be validated. When considering all the above factors, it is apparent that these results cannot be used with any degree of validity for the reviewing of process options for the recovery of silver and gold.

16.2.2   INTRODUCTION TO THE MMI 2004 METALLURGICAL TEST PROGRAM

The 2004 test program was a better structured program, which included the pre-concentration processes such as gravity concentration and flotation, both with and without regrinding, in an attempt to upgrade the material into a smaller mass for the subsequent treatment for the recovery of silver and gold. Also, cyanidation leach tests were conducted on as-received samples as well as samples that were reground in order to attempt to improve the liberation of silver and gold from the associated minerals. A single column leach test was also performed.

Additional work done included the establishing of the Specific Gravity and Bulk Density of the material, determining the Bond Mill Work Index on an oxide sample from the open pit, settling and filtration tests following cyanidation tests, and electrowinning tests using EMEW technology. All the different test procedures are summarised in the following Table 16.5.

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Table 16.5 Test Procedures - MMI 2004 Test Program

Process/Procedure	Details of Test	Sample Identity
Sample Preparation	Individually numbered ; dried; weighed; subsequently composited	Composites A, B and C
Head Assays	Fire assays, AA and ICP multi-acid	Individual samples, and Composites A, B and C
Specific Gravity	Standard pycnometer test	Composites A, B and C
Bulk Density	Standard volume displacement test	Composites A, B and C
Mineralogical	Examination of as-received samples	Selected Samples
Test Product Assays	Fire assays, AA and ICP multi-acid	All test products
Bond Mill Work Index	6 cycles; closing screen size 150 microns	Oxide sample
Size-Assay Distribution	Screened and assayed the size fractions	Selected samples
Gravity Concentration	Various test conditions	Composites A, B and C
Cyanidation Leach	Various test conditions	Composites A, B and C
Flotation	Various test conditions	Composites A, B and C
Column Leach Test	Agglomerated feed; 81 day duration; 0.5 to 1.0 g/l NaCN; pH 10.5; 0.05 ml/s	Composite of A and B
EMEW Electrowinning	Various test conditions	PLS from leach test
Acid Base Accounting	Acid generation tests	Composites A, B and C

The results obtained from this test program led MMI to include the heap leach process as the recommended treatment option in their report “Preliminary Feasibility” dated May 2005 [Reference #4].

16.2.3  EVALUATION AND REVIEW OF METALLURGICAL TESTS

The metallurgical tests conducted according to the MMI 2004 test program will now be reviewed. The process implications of the procedures and processes investigated, and the results obtained, will be discussed below. The most promising process option will be selected as the recommended process treatment route based on the evaluation of the results obtained from the test program. This process option will then be evaluated with respect to capital and operating cost estimates.

Sample Preparation and Characteristics

Bagged samples carrying the MMI identification tags were prepared at Avino Mine under the direct supervision of MMI personnel. These samples were then transported from the mine-site to Durango, Mexico, and shipped via airfreight to Vancouver, BC. The samples were delivered to the PRA facility, and unpacked in the presence of MMI personnel to ensure that no tampering had occurred to the samples en-route. The samples were subsequently re-numbered by MMI prior to PRA staff un-bagging and drying the samples. These details are shown on the PRA Sample Receiving Log [Reference #3]. The individual samples were initially air-dried, followed by low-temperature, < 50 degrees Centigrade, oven drying.

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The individual samples were subsequently homogenised, and riffled, and split into four one-quarter fractions. One of these fractions was used for head assay determinations. A second fraction was used for compositing selected individual samples to create the sample Composite A, representing the oxide material of the lower bench of the tailings dam. Similarly, Composite B, representing the oxide material of the middle bench of the tailings dam, was prepared by compositing selected individual samples, as was Composite C, representing the sulphide tailings of the upper bench.

Although the samples had arrived at PRA from Avino Mine without any indication of tampering, it is the sampling regime itself, which is considered to be deficient. Firstly, the sampling of the oxide section of the tailings dam was incomplete. The sampling did not replicate the 1990 Cia Minera program, and certain parts of the tailings dam were not sampled. Secondly, the samples that were taken by MMI only represent the first 4 metres of depth of the tailings dam. Indications are, however, that the overall depth of the oxide section of the tailings dam varies between 7 and 27 metres. These two major deficiencies were also recognised by the CSRA as deficiencies during their review. Both these items were addressed in the final MMI report “A Tailings Resource” dated October 2005 [Reference #7]. The October 2005 report recommended a more detailed program of sampling of the whole tailings dam up to bedrock or ground soil level, as well as conducting metallurgical characterisation tests using representative material from this more detailed sampling process whenever this is to be performed. However, since the MMI Technical Report, as reviewed by the CSRA, subsequently referred to the oxide tailings as an “inferred resource” [Reference #7], this and other sampling discrepancies noted in the MMI test program, will not be discussed any further.

Moisture Content

The moisture contents of the samples as received from the Avino Mine tailings dam were found to vary widely, namely from a low value of 5.12% to a high value of 28.25% moisture. A frequency distribution for moisture content of all the oxide tailings samples as received by PRA is given in Table 16.6 below. The bi-nodal distribution is apparent.

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Table 16.6 Moisture Content of Samples

Frequency Distribution
Moisture Content - Range, %	Number
5.00 - 7.50	9
7.51 - 10.00	14
10.01 - 12.50	19
12.51 - 15.00	16
15.01 - 17.50	5
17.51 - 20.00	5
20.01 - 22.50	12
22.51 - 25.00	5
25.01 - 27.50	0
27.51 - 30.00	1

The particular presence of these high moisture content values in the tailings dam apparently confirms the high moisture content values found during the 1990 sampling program conducted by Cia Minera. Although the precise sampling procedure and drying conditions are unrecorded, a data sheet provided by Avino Mines as ostensibly related to this sampling program, provides assay values and moisture contents obtained during the program. The moisture values obtained varied from a low moisture value of 13.89% to a high value of 29.4% and a calculated average of 22.87% moisture. This data sheet, titled ‘Bloque de Reservas, Presa de Jales’, together with the individual assay values recorded during the 1990 drilling program, is included in Appendix J.

A possible reason for the high moisture content of the tailings material is that the mine was operational during this period when the sampling program was undertaken, i.e. 1990, and that routine tailings deposition was still in progress.

The specific reason for the relatively high moisture contents found during the 2004 MMI sampling program, is not apparent. The MMI Technical Report has referred to the possibility of the original manner of deposition of the tailings which has resulted in the localised areas of high moisture content. Also, the presence of artesian springs under the tailings dam has also been mentioned as a possible reason. It was also observed that any rain water run-off from the higher levels above the tailings dam would collect at the head of the tailings dam and subsequently seep through the dam exiting at the foot of the dam. Whatever the reason(s) may be, areas of high moisture content do exist and will influence the method of recovery of the tailings and the subsequent agglomeration process.

Head Assays and Test Products Assays

Gold assaying was done using the standard fire assay procedure. Initially the silver was also analysed by the fire assay procedure followed by an atomic absorption spectrophotometric [AA] finish. However, this fire assay based method for silver is not very accurate in the low concentration range of < 100 g/t for silver. Assaying for silver was then done using the Inductively Coupled Plasma Spectroscopy method preceded by the total digestion of the sample in a suite of mineral acids [ICP-M]. A further method was also investigated, namely that of total acid digestion followed by an AA finish. The results obtained with this acid digestion and AA method were similar to the ICP-M method. The assay method selected for all the silver assays was therefore the ICP-M method preceded by the total digestion of the sample in a suite of mineral acids [ICP-M]. All the other analyses for the various products arising from the metallurgical tests were done by the standard and universal methods using titration, ICP-M or AA methods.

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All the various head sample analyses conducted during the test program have been listed in Table 16.7. The reference to the Test Number relates to the stage of the testwork that the sample was submitted for analysis. The average values for the four different composite samples tested, namely Composite A, Composite B, Composite C and the Composite A + B blended sample, have all been calculated and are given in the table together with the respective standard deviation values. The standard deviation of the head samples representing Composite A and Composite B are shown to be within 10% of the deviation from the average value. This is considered to be reasonable.

However, the average silver value of all the head assay analyses assayed as head samples representing both Composite A and Composite B together is only 86.8 g/t Ag. This average value of 86.8 g/t Ag is less than the 95.5 g/t Ag as given in the MMI Technical Report as being the overall silver grade of the material of the whole oxide tailings dam [Reference #7]. Similarly, the average gold value of all the head assay analyses assayed as head samples representing both Composite A and Composite B [i.e. representing the oxide tailings dam] taken during the testwork program, is 0.44 g/t Au which also is less than the 0.53 g/t Au as quoted in the MMI Technical Report. For silver, this amounts to a difference of about 9% based on the MMI quoted head grade of 95.5 g/t Ag, while for gold the difference is larger at 17% based on the MMI quoted gold value of 0.53 g/t Au. It is of interest that the average head assay for the Composite A + B sample is closer to the calculated average for Composite A and for Composite B, namely 89.6 g/t compared with 86.8 g/t for silver, and 0.41 g/t compared with 0.44 g/t for gold. The above discussion assumes that the tonnages of the tailings dam labelled Composite A (lower bench) will be mixed in equal proportion to the area of the tailings dam designated as Composite B (middle bench). In the absence of specific tailings dam volumes, or tonnages, this assumption may be an oversimplification and may therefore not be entirely valid. However, the assay values representing the tailings area (Composite B) will lower the overall average head grade of the tailings being treated.

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Table 16.7 Head Assays

Test No.	Composite A Assays, g/t		Test No.	Composite B Assays, g/t
	Ag	Au		Ag	Au
SA9	99.8	0.37	SA10	88.3	0.55
Ave. 1	103.4	0.34	Ave. 1	82.6	0.68
Ave. 2	105.3	0.36	Ave. 2	88.4	0.51
C1	95.2	0.35	C4	76.3	0.52
C2	94.3	0.35	C5	70.6	0.49
C3	94.1	0.36	C6	71.4	0.50
C7	88.7	0.36	C9	70.3	0.52
C8	88.7	0.36	C10	70.3	0.52
C13	95.9	0.28	C15	77.2	0.49
C14	98.9	0.37	C16	78.3	0.52
C17	95.2	0.35	C18	77.2	0.49
Average Value:	96.32	0.350		77.35	0.526
Standard Deviation:	5.27	0.025		6.72	0.054
	Composite C			Column Composite A + B
C11	39.8	0.34		87.4		0.42
C12	39.8	0.34		90.1		0.40
Ave. 1	31.7	0.29		91.4		0.42
Ave. 2	39.8	0.39		-		-
Average Value:	37.78	0.340		89.63		0.413
Standard Deviation:	4.05	0.041		2.04		0.012

A further comment regarding the assay results above relates to the methods employed for the assaying techniques for silver from these samples. The MMI Technical Report states that for the Cia Minera 1990 tailings drilling program, the silver assaying was done using the mine standard practice of fire assay followed by acid digestion and AA finish. The PRA metallurgical testwork program used multi-acid digestion followed by ICP assay method for silver analyses. It is anticipated that there will not be a significant difference between the silver assays as reported in 1990 and those from the MMI test program as conducted by PRA, but the extent of this difference cannot be quantified in this review. Similarly, no comment can be given as to the accuracy of the assays conducted by Cia Minera since the standards of precision of sampling, sample preparation and detailed methodology of the assaying methods are unknown. However, a summary sheet containing assay values has been provided by Avino Mines as being the silver and gold grades obtained from the 1990 Cia Minera sampling program. The summary sheet is attached in Appendix J. No calculations have been performed using these assay values and it is only included in this report since it is part of the Cia Minera sampling program. The MMI report ‘A Tailings Resource’ [Reference #7] provides a grid map identifying the various sample holes.

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Mineralogical Evaluation

At the start of the 2004 metallurgical test program, MMI requested that a sample from some of the individual samples be submitted for mineralogical analysis. The mineralogical findings have not been reported in the PRA Report No. 0406407 [Reference #3], and also were not alluded to in the MMI Technical Report [Reference #7], nor in any of the preceding reports. The reason(s) why these results have apparently not been communicated to Avino Mine or to the investigators of the test program at PRA, is not known.

Bond Ball Mill Work Index

Although this information was not required for the treatment of the oxide tailings dam material, a Bond Ball Mill Work Index determination test was done on an oxide ore sample. The work index was determined to be 12.3 kWhr/t using a closing screen size of 74 microns [200 mesh] with convergence of the specific energy input (grams of product per revolution) found after 5 cycles of testing. This makes the sample tested a moderately hard rock type. The details regarding the origin of this sample have not been documented and its relevance as data is therefore questioned.

Bulk Density and Specific Gravity

Bulk density and specific gravity determinations were conducted on samples specifically identified by MMI. The specific gravity measurements were done using the standard pycnometric method, while the bulk density values were obtained by measuring the volume of dry solids in a measuring cylinder. The values obtained are reproduced in the table below.

Table 16.8 Bulk Density and Specific Gravity

Location/ Bench	Sample Identity	P80 Size, micron	Bulk Density,g/cm3	Specific Gravity
Upper Bench	S2	226	1.66	2.74
Lower Bench	S10	326	1.73	2.62
Lower Bench	S22	367	1.73	2.76
Middle Bench	S45	254	1.60	2.76
Middle Bench	S50	201	1.63	2.74
Upper Bench	S74	301	1.57	2.72
Average	-	-	1.653	2.723

The bulk density values determined for the oxide tailings material were found to vary between 1.57 and 1.73 g/cm³ with an average of 1.653 g/cm³. This average value is in reasonable accord with the bulk density of 1.605 g/cm³ as quoted in the MMI Technical Report. The specific gravity values obtained were generally consistent with an average value of 2.723.

Particle Size - Assay Analysis

A particle size - fraction analysis was done on the same samples as were used for the bulk density and specific gravity determinations. These tests were conducted to determine whether the silver and gold were predominantly occurring in a particular particle size range. The size-assay analyses indicated that the metal distributions were varied according to the location, but that all displayed the bi-nodal distribution for silver, gold and mass to varying degrees.

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Sample S10 from Composite A from the Lower Bench of the tailings dam indicated one maximum metal distribution occurring in the size range 149 to 210 microns, and another in the minus 37 microns size range. The maximum mass distributions are generally similar although it occurs over a wider size range in the coarse size, namely 105 to 210 microns. The second sample from this bench, Sample S22, was similar but with a shifted maximum metal and mass distribution in the 210 to 297 microns size range, and a secondary maximum metal and mass distribution in the minus 37 microns size range.

Sample S45 from the Middle Bench of the tailings dam, and part of Composite B, indicated maximum metal distribution in the 149 to 210 microns size range with maximum mass distribution in the 105 to 149 microns size range. The secondary maximum metal and mass distribution was found in the minus 37 microns size range. The second sample from the Middle Bench, namely Sample S50, had the maximum metal and mass distributions in the 105 to 149 microns size range as well as the minus 37 microns size range.

The two samples from the Upper Bench of the tailings dam of Composite C displayed totally different particle size distributions. Sample S2 was bi-nodal with one maximum for metal and mass distribution in the size range 105 to 149 microns and the second maximum occurring for the size range of minus 37 microns. Sample S74 displayed only one maximum metal and mass distribution over the relatively wide coarse particle size range of 105 to 297 microns. This sample was almost entirely devoid of slimes, or minus 37 microns material.

These samples reflect the operating discharge conditions and history at the time of plant operations and tailings deposition. The results typify the use of a tailings cyclone situated on the tailings dam wall discharging the coarse undersize material onto the wall area with the finer cyclone overflow material flowing downstream and settling within the tailings dam. Changes in the size distribution would be anticipated with downstream distance from the point of discharge by the cyclones at the tailings dam wall. This is typified by the size distribution of Sample S74 which purports to be a cyclone underflow sample taken at the point of discharge and which was found to be almost totally devoid of fines, or minus 37 microns material.

Gravity Concentration Tests

Pre-concentration tests using the centrifugal gravity concentration method were conducted to evaluate the potential upgrading of silver and gold. The laboratory size concentrator used was the Falcon Model SB40 centrifugal concentrator. The tests were conducted on samples from Composites A, B and C. MMI dictated the test parameters used for these tests, including a set of tests where the samples were reground prior to conducting the gravity concentration test. The results from the gravity concentration tests are summarised in Table 16.9 below.

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Table 16.9 Summary of Results of Gravity Concentration Tests

Sample Identity	Head Grade		Concentrate Grade		% Recovery			P80 micron	Remarks [Note: All tests are 3-pass tests]
	Ag, g/t	Au, g/t	Ag, g/t	Au, g/t	Mass	Ag	Au
Comp. A	93.8	0.35	124.7	0.52	24.1	32.1	36.5	269	Pressure 1.5 psig; no regrind
Comp. B	70.3	0.50	96.9	0.71	23.6	32.5	33.3	180	“
Comp. C	39.7	0.33	58.0	0.65	24.1	35.2	47.0	254	“
Comp. A	92.1	0.33	126.1	0.71	19.7	27.2	42.1	76	Pressure 1.0 psig; reground
Comp. B	70.5	0.56	96.5	1.29	22.4	30.7	51.5	77	“
Comp. C	40.7	0.38	65.5	0.98	24.8	39.9	64.3	79	“

The mass recoveries varied between 20 and 25% indicating that the tests were performed in a uniform and consistent manner. The highest silver recovery obtained was 40% (after regrind) for Composite C and decreasing to 31% for Composite B (after regrind) and about 27% for Composite A, also after regrind. The gold recoveries were higher than the equivalent silver recoveries, particularly after regrind, indicating that the liberation of the precious metals could be incomplete.

However, the upgrading factor for both silver and gold is very low, namely about 1.4 for silver and up to 2.3 for gold. No further upgrading or silver and gold recovery tests were conducted on the gravity concentrates produced possibly as a result of the relatively low grades and recoveries obtained. Also of interest is the fact that no historical testwork was documented by MMI where gravity concentration was used to produce a saleable high-grade concentrate.

Flotation

Different scoping flotation tests were conducted on samples from Composite A and Composite B using various reagent schemes and conditions as dictated by MMI. The results of the flotation tests are summarised in Table 16.10. The test results reported have lead to the following conclusions.

For Composite A, a regrind from a P80 size of 238 microns [as received particle size] to a P80 of 72 microns, improved the flotation recovery of silver from 18 to 23%, and that of gold from 18 to 39%. The standard suite of reagents was used for these tests [Tests F1, F2 and F3]. For Composite B, a regrind from a P80 size of 173 microns [as received particle size] to a P80 of 74 microns, improved the flotation recovery of silver from 22 to 33%, and that of gold from 12 to 32% [Tests F2, F5 and F6]. A particle size fraction analysis distribution conducted on the tailings of Test F4 [Composite A] indicated that the major proportion of the mass and the silver and gold is present in the slimes, or minus 37 microns, size fraction. However, significant losses of silver, and particularly gold, occurred in the coarser sizes, namely the size range 53 to 105 microns. This indicates that the degree of liberation could be improved and that some metal appears to be occluded in the coarser particle sizes. Some silver may also be adsorbed onto secondary oxide minerals and be unrecoverable by flotation. A similar mass and metal distribution was obtained in the case of Test F9 [also Composite A] which was a flotation test performed using a sulphidisation reagent.

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In testing the various flotation reagent suites, variable mass and metal recoveries and concentrate grades were obtained. However, the maximum silver grade obtained for a rougher concentrate was 909 g/t Ag, while the overall recoveries for silver could not be improved beyond about 40%. This indicated that mineral surface alteration or oxidation, or occlusion of precious metals in gangue, was inhibiting the flotation process. Since the silver recoveries obtained were deemed low and unsatisfactory, no further flotation tests were conducted and no extraction tests were performed on flotation concentrates.

The head assays obtained during the flotation testing stage gave inconsistent results. Table 10 shows the actual head assays obtained for each flotation test compared with the head assay obtained for silver for the composite samples. For Composite A, the individual silver head values for each flotation test conducted are all higher than the assay for the composite sample, except in the case of Test F11. The gold [and silver] values obtained for Tests F7, F8 and F9, are known to have been the result of poor sampling technique adopted for these three tests. The composite head assay gold value of 0.36 g/t Au is probably a reasonably representative assay value for Composite A. For Composite B, the silver head value for the composite sample is slightly lower than the assays for the individual flotation tests. For gold, the composite sample value is higher at 0.52 g/t Au than the assays for the individual tests.

The historical results of the flotation tests reported in Table 16.3 are significantly higher at 60 to 69% recovery for silver and 47 to 67% for gold. However, in the absence of information regarding the origins of these samples, the lack of head grade data and the absence of sampling and flotation procedures involved, these results will not be taken into consideration in selecting of the processing options for the oxide tailings dam material.

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Table 16.10 Summary of Results of Flotation Tests

Sample Identity & Test No.	Head Grade		Conc. Grade		% Recovery			P80 Micron	Remarks
	Ag, g/t	Au, g/t	Ag, g/t	Au, g/t	Mass	Ag	Au
Comp. A /F1	112.2	0.35	908.7	3.17	2.1	17.8	18.4	238	3-stage ro., pH 8;
Comp. A /F3	119.2	0.39	734.6	3.88	2.6	21.0	30.4	103	Conditioning NaCN +
Comp. A /F4	104.6	0.40	630.9	3.36	3.8	22.6	38.6	72	Na2CO3; A404, PAX
Comp. A /F7	111.9	1.39	654.6	5.56	2.3	16.3	34.9	~ 75	2-stage ro., nil NaCN
Comp. A /F8	108.5	2.38	887.2	11.91	0.9	7.8	30.7	~ 75	2-stage ro., nil NaCN
Comp. A /F9	114.5	1.67	723.9	5.86	2.7	20.8	45.0	~ 75	2-stage ro., NaS2, PAX
Comp. A /F10	103.5	0.36	401.3	1.62	8.9	34.6	39.8	~ 75	with NaCO3, CuSO4
Comp. A /F11	99.6	0.34	484.8	1.83	8.8	42.2	48.3	~ 75	with CuSO4, A208
Comp. B /F2	88.4	0.42	695.4	2.65	2.6	22.0	12.2	173	3-stage ro., pH 8
Comp. B /F5	89.7	0.47	806.1	4.18	2.9	27.0	24.6	92	conditioning NaCN +
Comp. B /F6	89.9	0.51	867.1	5.45	2.9	32.5	32.1	74	Na2CO3; A404, PAX
Comp.A: Head	99.8	0.36	-	-	-	-	-	-	-
Comp.B: Head	88.3	0.52	-	-	-	-	-	-	-

Cyanidation Tests

Cyanide leaching tests were conducted on samples from Composite A, Composite B and Composite C using different leaching conditions. The first set of tests were to determine the effect of regrinding the tailings samples prior to leaching while subsequent tests determined the effect of cyanide concentration in the leach solution.

For Composite A, the silver extractions varied from 66% for the unmilled [as received] sample to 80% for the samples that were reground, while the gold extractions varied from 82 to 89% respectively. For Composite B, the silver extractions ranged between 69% for as-received material, to 77% for samples that were reground. The corresponding gold extractions varied between 82 and 87%. Although the cyanide consumption increased with the regrinding of samples tested for both Composite A and Composite B, the increase in extraction may compensate for the additional cost of cyanide reagent and regrinding provided that the filtration characteristics are not detrimentally effected. Higher cyanide concentrations in the leach solution tended to improve the extractions of silver and gold, but increased the cyanide consumption significantly as well.

The results from the sulphide tailings, namely Composite C, indicate that between 80 and 83% of the silver can be extracted, with between 77 and 85% of the gold. However, the cyanide consumption values were higher than the results from the oxide tailings. Two leach tests only were conducted on reground samples from Composite C, each having a P80 of about 69 microns.

A summary of the cyanide leach test results is given in the following table.

Avino Silver and Gold Mines Ltd.	29	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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Table 16.11 Summary of Results of PRA Cyanidation Tests

Sample Identity & Test No.	% Extraction		Reagent Usage, kg/t		NaCN Conc., g/l	P80 micron
	Ag	Au	NaCN	Lime
Comp A+ / C1	66.4	81.5	1.8	1.4	1.0	269
Comp A+ / C2	79.3	85.7	1.6	1.8	1.0	103
Comp A+ / C3	80.4	89.1	2.6	1.6	1.0	78
Comp A+ / C7	78.6	82.7	2.2	1.8	0.5	74
Comp A+ / C8	89.7	85.5	5.1	0.8	2.0	74
Comp A* / C 13	79.7	86.8	1.5	1.3	0.5	74
Comp A* / C 14	83.1	82.1	3.7	0.8	2.0	74
Comp A* / C 17	79.4	90.9	1.0	1.2	1.0	74
Comp B+ / C4	69.1	82.0	2.6	1.8	1.0	180
Comp B+ / C5	77.1	88.3	1.7	1.8	1.0	100
Comp B+ / C6	77.3	86.9	1.7	1.9	1.0	84
Comp B+ / C9	73.2	86.0	2.6	1.2	0.5	84
Comp B+ / C10	79.5	86.4	4.5	1.0	2.0	84
Comp B* / C15	72.9	82.6	1.6	2.0	0.5	84
Comp B* / C16	75.4	83.4	3.8	1.0	2.0	84
Comp B* / C18	67.7	78.6	0.9	1.3	1.0	84
Comp C+ / C11	73.8	77.3	4.0	2.8	1.0	69
Comp C+ / C12	86.6	85.0	7.3	2.6	2.0	67
+ indicates Original Composite Sample
* indicates New Composite Sample
Tests C17 & C18 = 24hr leach duration; other tests = 72 hr leach duration.

During the cyanide leach test program, a new Composite A and Composite B sample had to be prepared since the original composite samples had been exhausted. Comparison of results from the two composite samples indicated similar behaviour patterns, although there are some noticeable differences in the extractions. Also, the cyanide and lime consumption values as recorded are inconsistent. This indicates that absolute numbers cannot be assigned to a single test although any observed trends would be valid. The averages of similar tests would more likely predict the overall responses more accurately. It is also apparent that non-systematic variations in the assay results could have arisen from subtle variations in mineralogy, sample preparation, the sample regrinding process and possibly daily variations in temperature.

The cyanide leach extraction results quoted by MMI in Table 16.2, and the averaged results from the present test program, are summarised below in Table 16.12, and will be discussed in the following section.

Avino Silver and Gold Mines Ltd.	30	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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Table 16.12 Summary of Cyanidation Test Results used by the MMI Reports

Sample Identity & Test No.	% Extraction		Remarks
	Ag	Au
Comp. A / C1	66	82	as-received; 1.0 g/I NaCN
Comp. A / C7 & C13	80	88	average; reground; 0.5 g/I NaCN
Comp. B / C4	69	82	as-received; 1.0 g/L NaCN
Comp. B / C9 & C15	73	84	average; reground; 0.5 g/l NaCN
MMI 2003	77	71	results from 2003 test program
MMI 2003	88	88	origin of results unrecorded
MMI 2004 / C8 & C10	85	86	average; reground; 2.0 g/l NaCN

The average extraction results obtained from samples from Composite A and Composite B in the present study are generally lower than the results from the historical testwork as detailed in Table 2. However, in the absence of details, these historical results cannot be used in the overall evaluation of this process. The MMI claim of a 77% silver extraction, based on the MMI 2003 test program, cannot be considered an acceptable result since only one test was done. The sample origin is purported to be four holes dug at approximately 25 metre intervals with samples scraped into a bag, one for the lower bench and one for the upper bench of the oxide tailings dam. Clearly, a sample collected in this manner cannot be considered to be representative. Also, the other MMI 2003 claim for an extraction result of 89% silver and 88% gold cannot be validated. All these test results can therefore not be considered as valid and will not be used in any further discussions or evaluations.

The MMI 2004 results, as claimed in the Technical Report and listed in Table 16.12 above, are also considered unusable. The reasons for this statement are that these results were obtained with a reground sample and leached at a high cyanide concentration of 2.0 g/l NaCN, whereas the other tests were done using 1.0 g/l NaCN. Both these conditions, that is, the regrinding of the tailings material and a high cyanide concentration leach condition, will not be implemented in a recovery process and these results are considered to be unrealistic.

The extraction results from the cyanidation tests obtained using as-received samples from Composite A and Composite B, namely 66 to 69% for silver and 82% for gold, were encouraging.

Column Leach Test

One column leach test was conducted on a 30.9 kg sample being an equal mix of material from Composite A and Composite B. The sample was mixed with water, Portland Cement and lime and then agglomerated to a P80 size of 2614 microns. After curing, the sample was put into a column with a diameter of 102 mm and a height of 3 m. The column test was run for a total of 81 days after the solution flowrate and pH had been stabilised. The silver extraction obtained was 73.0% while the gold extraction was 78.9%. These results compare very well to the average extraction values calculated from the cyanidation tests of the individual composite samples leached in the as-received condition, namely 67.8% for silver and 81.8% for gold. The cyanide consumption values are also comparable. The results obtained from the column test, as well as the calculated average extraction values obtained from the tests conducted on the as-received samples of Composite A and Composite B, have been summarised in Table 16.13.

Avino Silver and Gold Mines Ltd.	31	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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Table 16.13 Summary of Results of Column Leach Tests

Sample & Test No.	% Extraction		Reagent Consumption, kg/t			NaCN Conc., g/l	P80 micron	Remarks
	Ag	Au	NaCN	Lime	Cement
Column Test, Comp. A & B	73.0	78.9	2.32	13.73	21.8	0.5 & 2.0	2614	pH 11; flowrate 0.05 ml/s
Comp. A & B Average, Tests C1&C4	67.8	81.8	2.18	1.59	-	1.0	225	pH 10.5/11; bottle roll

The kinetics of leaching had slowed down significantly by Day 81 when the test was terminated, although there was evidence that some leaching was still in progress.

A particle size - assay analysis of the leach residue of the column test found that the highest unleached [undissolved] silver grade was in the coarsest size range of plus 210 microns, while the highest gold value was found in the minus 37 microns size range. This suggests both inadequate liberation of the silver grains and/or minerals, and occlusion of gold possibly by clay minerals, or the presence of tarnished/coated mineral surfaces, or the presence of refractory minerals. The subsequent leaching of de-agglomerated column leach test residue resulted in a negligible extraction of silver and gold. This indicates that the column leach test had virtually reached its maximum potential extraction, which confirms the observation that the leaching rate had slowed down.

Only one column leach test was conducted. Also, the material tested was a mixture of samples from Composite A and Composite B, that is, a mixture of material from the lower and the middle benches of the oxide tailings dam. During the test, flow problems were encountered which resulted in the column having to be unloaded and the material having to be re-agglomerated with the test subsequently re-started after filling the column. In general terms, the results from one test only can not be regarded as representative of the whole oxide tailings dam. However, despite these limitations and problems encountered, the encouraging results obtained and the close comparison with the bottle-roll tests, implies that the results are relatively reliable. The extraction values obtained from the column test, namely 73.0% for silver and 81.8% for gold, will therefore be used in the evaluation of this treatment process. The reagent consumption values also appear to be very high, namely 13.73 kg/t for lime, 21.8 kg/t for cement and 2.32 kg/t for cyanide. However, lime and cement consumption values obtained in laboratory tests generally approximate commercial operations although, in this case, they seem to be unrealistically high. The cyanide consumption of a commercial operation would typically only be 30 to 50% of that measured in a laboratory test.

Acid Base Accounting [ABA]

The ABA results predict the overall acid generating potential of selected samples. A net acid general potential was found for the sulphide tailings but not the oxide tailings. The processing of the sulphide tailings for silver and gold recovery could modify the ABA and increase the stability of the ultimate residues. Alternatively, the sulphide tailings would require the addition of lime during the process of relocating this material. This would ensure that the sulphide tailings would not cause acid-generating environmental problems.

Avino Silver and Gold Mines Ltd.	32	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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Electrowinning

Electrowinning metal recovery tests were conducted using EMEW technology [from Electrometals Electrowinning company], specifically designed for the electrodeposition of metals from dilute solution tenors. The tests were carried out using filtered cyanide leach pregnant solutions. Although the test results were favourable, it appears unlikely at this stage that this technology could be applied in this situation given the high solution volumes generated and the very low silver concentrations anticipated in the pregnant solution from the heap. However, further testwork using the EMEW metal recovery system should be undertaken if the project advances to the feasibility level because the potential for savings in capital cost and operating cost needs to be investigated.

16.3      PROCESS ENGINEERING EVALUATION

This section will review the process engineering options based on the results of the metallurgical tests conducted and will set the metallurgical design criteria for purposes of estimating the operating costs and capital costs. The costs of the processing options selected for review will be estimated based on conventional plant layouts. The design criteria, layouts and cost estimates will be kept on a general basis and will be based on the knowledge gained from the site visit and the interpretation of conditions and the availability of existing plant equipment, manpower and infrastructure for this Technical Report.

16.3.1    GRAVITY CONCENTRATION

Review of Results

As indicated in Table 16.9, the upgrading for silver from the as-received oxide tailings was poor with a maximum concentrate grade of 125 g/t Ag with a mass recovery of 20%. The upgrading of gold is similarly poor. The re-grinding of the samples prior to gravity concentration leads to an almost negligible improvement in the upgrading of silver to 126 g/t Ag, while for gold a maximum concentrate grade of 1.29 g/t Au was obtained. The sulphide tailings response to gravity concentration is equally poor with even lower grade gravity concentrates being obtained despite slightly improved recoveries being observed for both silver and gold.

Conclusion

The poor results obtained in that no high-grade metal concentrate could be produced, coupled with the fact that no extraction tests for silver and gold were conducted on the gravity concentrates produced, has resulted in the gravity concentration treatment option not being selected for further consideration.

16.3.2   FLOTATION

Review of Results

The flotation results have been summarised in Table 16.10. The results indicate that the overall recoveries for both silver and gold are low, namely between 8 and 42% for silver and 12 to 48% for gold. The re-grinding of both tailings samples (Composite A and Composite B) are seen to improve the recoveries, while the testing of various reagent regimes also resulted in improvements to the overall recoveries of both silver and gold in some cases. However, the overall recoveries are generally considered to be low at < 40% for silver and < 48% for gold, and this is coupled with a very low grade concentrate being produced. This poor flotation response is probably the result of surface alterations and/or inadequate liberation of the silver and gold. No extraction tests were conducted on any of the flotation concentrates produced and so the total extent of extraction is not known. No tests were conducted on the sulphide tailings material [Composite C] and its response to flotation as a pre-concentration process is therefore not known.

Avino Silver and Gold Mines Ltd.	33	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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Conclusion

Flotation will not be considered as a treatment option for the recovery of silver and gold from the oxide tailings dam material. For the reasons specified above, namely a generally low recovery of silver and gold, the option of using flotation to recover silver and gold will not be considered as a processing method in the treatment of the oxide tailings dam material.

16.3.3   CYANIDE LEACHING

Review of Results

Cyanidation leach tests were done on samples from Composite A and Composite B under different conditions of particle size and solution cyanide concentration. The results have been summarised in Table 16.11. The results generally indicated that cyanidation was still occurring after 72 hours of the leaching time used for the laboratory tests, but at a much reduced rate. The base metals copper and zinc also dissolved during the cyanide leach and will contribute to the overall consumption of cyanide. Increasing the cyanide concentration in the leach solution generally improved the extraction of silver and gold, but also increased the overall cyanide consumption. The extraction of silver and gold from Composite A increased with fineness of grind, while Composite B did not improve the extraction for finer grinds than P80 of 100 microns. The cyanide consumption figures are inconsistent in some cases although trends are apparent. Although limited testwork was done on material from Composite C, namely the sulphide tailings, a set of results has been included in Table 16.14 below for purposes of comparison.

Table 16.14 Cyanide Leaching Parameters

Sample Identity	Head Grade		% Extraction		Reagent Cons, kg/t		NaCN Conc. g/l	P80 Micron	Remarks
	Ag, g/t	Au, g/t	Ag, g/t	Au, g/t	NaCN	Lime
Comp. A	94.7	0.35	66.4	81.5	1.8	1.4	1.0	269	As-received sample
Comp. B	95.9	0.28	69.1	82.0	2.6	1.8	1.0	180	“
Avg of A & B	95.3	0.32	67.8	81.8	2.2	1.6	1.0	225
Comp. A	94.7	0.35	79.3	85.7	1.6	1.8	1.0	103	Reground sample
Comp. B	70.3	0.52	77.1	88.3	1.7	1.8	1.0	100	“
Avg of A & B	82.5	0.44	78.2	87.0	1.7	1.8	1.0	102
Comp. C	39.8	0.34	73.8	77.3	4.0	2.8	1.0	69	“

Conclusions

As-received [unmilled] and reground tailings dam material will be expected to show the following extraction results under normal leaching conditions of about 68% for silver and 82% for gold. The reground material will give higher extractions at about 78% for silver and 87% for gold [see results in Table 16.13]. Although the regrinding of tailings material is considered to be an expensive treatment method, cyanidation with and without regrinding as a treatment option will be reviewed and discussed in Sections 16.3.6 and 16.3.7 below.

Avino Silver and Gold Mines Ltd.	34	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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16.3.4    COLUMN LEACH TEST

Review of Results

One column leach test was conducted using a blend of equal proportions of as-received [unmilled] Composite A and Composite B oxide tailings material. Despite interruptions in the leaching cycle as a result of the de-agglomeration of material in the column and the resultant percolation of fines, the overall extraction of silver was 73% and 79% for gold [see Table 13 for the results]. Although the test was terminated after a total leaching time of 81 days, indications were that the leaching process was nearing completion but had not finalised at that stage. The above extraction results compare very well with the average extraction results obtained from the bottle roll leach tests, namely 68% extraction for silver and 82% for gold. The cyanide consumption of 2.3 kg/t for the column test was also comparable with that obtained for the bottle roll leach tests, namely 2.2 kg/t. The lime consumption for the column test was significantly higher probably as a result of the two repeated agglomeration exercises.

Conclusions

Although only one column leach test was performed, the extraction results are in keeping with those obtained from the bottle roll tests. The results as given in Table 16.13 will be used for developing the process design criteria.

16.3.5   PRECIOUS METAL RECOVERY

Review of Results

Only one technology was tested for recovering precious metals from cyanide leach solutions. The pregnant solution arising from leach tests performed on oxide tailings 1aterial was used to conduct electrowinning tests. Three tests were conducted using the EMEW electrowinning technology. These tests indicated that silver could be electrowon from solutions with a starting concentration of about 58 mg/l Ag to a depleted electrolyte with about 3 mg/l Ag. The deposition was also shown to be very selective with respect to the co-deposition of base metals. However, the pregnant solution from a leaching heap is expected to be significantly less than 58 mg/l Ag, possibly as low as 16 mg/l Ag. It is unclear whether the EMEW technology could operate efficiently under such low silver tenors.

The alternative process options for the recovery of precious metals would likely be either activated carbon, or the zinc precipitation method. No tests were conducted on these two process options. The use of an activated carbon circuit to recover silver is not recommended because of the added operational complexity. Also, the relatively high grade of the silver in solution will result in the treating of relatively large amounts of carbon, which will add to the cost of the project.

Conclusions

No other historical testwork results were reported by MMI, nor are any alternative technology results known to have taken place, which tested the recovery of silver from the Avino Mine tailings material. The Merrill-Crowe process will therefore be the preferred technology to recover the silver and gold from pregnant leach solutions.

Avino Silver and Gold Mines Ltd.	35	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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16.3.6   POTENTIAL REVENUE ESTIMATION

Preamble

A very preliminary economic view has been taken in eliminating the process options in the above sections. The only purpose of conducting this revenue estimate exercise is to determine order-of-magnitude costs, which can be compared in order to eliminate the obviously unrealistic process options. The only potential process options that will be developed are those selected from the above discussion. The inherent value of the precious metals in the oxide tailings material of the Avino Mine tailings dam will now be calculated using the three process treatment options discussed previously. These treatment options are:

· cyanidation of the oxide tailings material without regrind

· cyanidation of the tailings with regrind

· heap leaching of the tailings without regrind.

The following table compares the estimated potential revenue of the three process treatment options for the oxide tailings only. The sulphide tailings are discussed in a later section in this report. Note that all references to costs are in US dollars.

Table 16.15 Inherent Value of Oxide Tailings

	Process Treatment Options			Remarks
	Cyanidation As-received	Cyanidation Reground	Heap Leach As-received
Tailings Dam, t	2,091,074	2,091,074	2,091,074	MMI Technical Report
Available for Treatment, %	100.0	100.0	100.0
Head Grade, Ag g/t	95.5	95.5	95.5	MMI Technical Report
Head Grade, Au g/t	0.53	0.53	0.53	MMI Technical Report
kg metal available, Ag	199,697.57	199,697.57	199,697.57
kg metal available, Au	1108.27	1108.27	1108.27
% Extraction, Ag	67.8	78.2	73.0	Data ex Tables 13/14
% Extraction, Au	81.8	87.0	78.9	Data ex Tables 13/14
% Precipitation, Ag & Au	96.0	96.0	96.0	Assumed value
kg metal recovered, Ag	129,979.15	149,916.96	139,948.05
kg metal recovered, Au	870.30	925.63	839.45
oz metal recovered, Ag	4,178,959.73	4,819,980.10	4,499,469.91	1 kg = 32.151 oz
oz metal recovered, Au	27,981.07	29,759.82	26,989.07
Silver Price, $/oz	8.00	8.00	8.00
Gold Price, $/oz	500.00	500.00	500.00
Silver Income Potential,$	33,431,678	38,559,841	34,428,011
Gold Income Potential, $	13,990,534	14,879,908	12,906,800
Total Potential Income, $	47,422,212	53,439,749	47,334,811

Avino Silver and Gold Mines Ltd.	36	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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Capital and Operating Cost Estimates

The estimated capital and operating costs of a standard lay-out cyanide leach plant will now be compared with the estimated capital and operating cost of a heap leach plant. A throughput of 1370 t/d will be assumed for each option based on a four-year treatment period. These basic treatment costs specifically do not include the reclamation and relocation of the sulphide tailings dam material. The costs of the treatment options are summarised below in Table 16.16. These estimated capital and operating costs are based on data from Western Mining Engineering 2005 Handbook.

Table 16.16 Summary of Cost Estimates - four-year Treatment

Process Option	Capital Cost US $(million)	Operating Cost US $/ t
Cyanide Leach - no regrind	29.18	8.60
Cyanide Leach - with regrind	30.97	11.65
Heap Leach - Estimate I	13.13	3.87
Heap Leach - Estimate II	14.74	5.24

The cost estimates figures in Table 16.16 indicate that the estimated cost of a cyanide leach plant, without or with a regrind circuit, amounts to almost 70% of the total potential income for the metallurgical parameters. These numbers are as determined by the MMI Technical Report and for the metal prices as quoted above. Even on the assumption that some of the original plant equipment could be salvaged and restored for use, and also assuming that used equipment could be acquired from elsewhere, the savings are not anticipated to be sufficiently significant to make this option financially attractive. Based on an assessment from the recent site visit, it would appear as if there is very little that could be salvaged from the Avino Mines plant for re-use in a new treatment facility. Both cyanidation process treatment options will therefore not be pursued any further. It should also be noted that this estimate is based on a treatment rate of 1370 t/d. Any increase in the throughput rate will increase the cost of a cyanidation treatment facility [see Section 16.3.7 below].

Heap Leach Plant

The capital cost for a 1370 t/d [500,000 t/year] heap leach plant has been estimated to be between $13 million and $15 million with an estimated operating cost of between about $4 /t and $5 /t treated based on information available on similar operations. On the assumption that these numbers are reasonable estimates, it is seen that the potential for profit is about $22 million over the four-year period of operation plus the time required for the project to be established, and project closure.

16.3.7    HEAP LEACHOPTION - ASSUMPTIONS

A number of factors, some unknown at this stage, could influence the above rudimentary calculations.

Treatment Period

The period of treatment of the oxide tailings is assumed to be four years. This treatment period gives the cost estimates as detailed in Table 16.16.

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Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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A higher throughput rate of 3000 t/d resulting in a shorter treatment period of two years has also been considered. The resulting cost estimates of the process options assuming a two- year treatment period are given in Table 16.17 below.

Table 16.17 Summary of Cost Estimates - Two-Year Treatment

Process Option	Capital Cost US $(million)	Operating Cost US $/ t
Cyanide Leach - no regrind	48.69	6.99
Cyanide Leach - with regrind	51.78	9.36
Heap Leach - Estimate I	21.91	3.15
Heap Leach - Estimate II	24.69	4.08

The estimated capital cost of a 3000 t/d cyanide leach plant, without or with a regrind circuit, exceeds the potential income as calculated in Table 16.15. This treatment option will therefore not be considered. The capital cost for a 3000 t/d heap leach plant has been estimated to be between $22 million and $25 million with an estimated operating cost of between about $3 /t and $4 /t of material treated. In this case, the potential for profit is only about $14 million over the two-year period of operation, but would result in a negative cash flow situation. The four-year treatment period as discussed in Section 16.3.6 above will therefore be the preferred treatment option.

Heap Leach Lay-out

The heap layout, heap lift height, and number of lifts, have been assumed for purposes of this report, and are detailed in Section 16.4.4 of this report. The maximum height has been restricted to 26 m as a result of the proximity of the proposed heap leach facility to the community of San Jose de Avino. This proposed height for the heap would require geotechnical verification. Despite the 26 m height proposed for the heap, this has resulted in a relatively large surface area being required for the leach pad. The site layout and available space, site drainage, and pad size have been done according to the most suitable surface area topography and the best available information. However, the close proximity of the proposed heap leach facilities to the community of San Jose de Avino, and its agricultural workings, may yet result in site and/or layout revisions.

Refurbished Equipment

The suitability and/or the possibility of refurbishing the existing equipment at the Avino Mines plant site also needs to be evaluated more specifically. Some water pumps, and possibly some conveyors, could be returned to service, and some of the existing tanks could be used for water storage/supply. However, it does not appear as if any other equipment could be salvaged and/or refurbished for use in a heap leach plant. The general availability of used equipment from other locations in the Durango district, or in Mexico, would need to be investigated.

Assay Laboratory Equipment

The assay laboratory would require re-equipping with suitable equipment. The availability of suitable used laboratory equipment and facilities is also not known.

Avino Silver and Gold Mines Ltd.	38	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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Sulphide Tailings

The option of whether the sulphide tailings should be re-treated on the heap, or re-located and dumped elsewhere while possibly being treated for environmental remediation, is not known at this stage. The absence of reliable sulphide tailings dump metallurgical information makes quantification of this option difficult. A detailed trade-off study should be undertaken to determine whether the re-treatment of this material would contribute to the profitability of the project. However, at this stage only limited metallurgical test data is available since no detailed metallurgical testwork was undertaken on this material during the MMI 2004 test program. Based on the assumption that 65% of the silver and gold would be recovered by heap leaching under the same conditions as given for the oxide tailings, the following table gives the potential income from treating this sulphide tailings material. The operating costs will probably be higher for this material than for the oxide material because of the increased requirements for lime and cyanide, but this cannot be quantified at this stage because of the lack of detailed test information. However, there will be a cost in relocating the sulphide material if it is not treated for silver and gold recovery on the heap. Indications are that the sulphide tailings will also require treatment for environmental remediation purposes in the future. These costs could be partially or completely be off-set by treating this material separately or together with the oxide material by the heap leach process.

The assumed sulphide tailings grade at 37.8 g/t Ag and 0.34 g/t Au is lower than the assumed oxide tailings grade of 95.5 g/t Ag and 0.53 g/t Au. Also, in the absence of specific metallurgical extraction information, the estimated recovery for the sulphide material has been set at 65% which is lower than that for the oxide material. Despite this, the apparent potential income of about $27 million is significant.

Avino Silver and Gold Mines Ltd.	39	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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Table 16.18 Estimated Inherent Value of Sulphide Tailings

	Process Treatment Options			Remarks
	Cyanidation Reground	CyanidationReground	Column Leach As-received
Tailings Dam, t	2,900,000	2,900,000	2,900,000	MMI Technical Report
Available for Treatment, %	100.0	100.0	100.0
Head Grade, Ag g/t	37.78	37.78	37.78	MMI Technical Report
Head Grade, Au g/t	0.34	0.34	0.34	MMI Technical Report
kg metal available, Ag	109,562.00	109,562.00	109,562.00
kg metal available, Au	986.00	986.00	986.00
Concentration, g/l NaCN	1.0	2.0	1.0	Data ex Table 11
% Extraction, Ag	73.8	86.6	65.0	Column extractions;
% Extraction, Au	77.3	85.0	65.0	assumed at 65.0%
% Precipitation, Ag & Au	96.0	96.0	96.0	Assumed value
kg metal recovered, Ag	77,622.49	91,085.46	68,366.69
kg metal recovered, Au	731.69	804.58	615.26
oz metal recovered, Ag	2,495,640.5	2,928,488.7	2,198,057.4	1 kg = 32.151 oz
oz metal recovered, Au	23,524.59	25,867.92	19,781.35
Silver Price, $/oz	8.00	8.00	8.00
Gold Price, $/oz	500.00	500.00	500.00
Silver Income Potential, $	19,965,124	23,427,910	17,584,459
Gold Income Potential, $	11,762,297	12,933,961	9,890,676
Total Potential Income, $	31,727,421	36,361,872	27,475,136

For purposes of the present Wardrop Technical Report, it will be assumed that the sulphide tailings will be moved to another location north-east of the proposed site for the leach pad. Again, it should also be mentioned that this proposed site is very close to the town of San Jose de Avino and this may result in objections from the local community.

Environmental Concerns

ABA tests have indicated that mild acid generation may already have started on the tailings dam. Processing of at least the oxide tailings will mitigate these concerns. However, the sulphide tailings have been shown to be potential net acid producers. Re-locating the sulphide tailings may afford a timeous opportunity to address this potential environmental problem. Alternatively, the treatment of the sulphide tailings for gold recovery will afford an opportunity to recover silver and gold from the material as well as treating this material with the lime to ensure that this material will not be a net acid producer [see discussion in the previous section].

Remediation work will also be required at the end of the life of the heap leach project. No cost estimates have been undertaken to ensure the heap meets the environmental requirements once the processing of the heap material has been terminated.

Avino Silver and Gold Mines Ltd.	40	0551920100-REP-R0002-01
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Other Socio-Economic Risks

It is assumed that no other costs or claims are associated with this project. Thus the cost of permitting has not been considered, and neither has the cost of expropriating agricultural land for the leach pad, nor the cost of water which would have to be re-directed to the heap leach project but which is currently used for agricultural purposes.

16.4      PROCESS FLOWSHEET

16.4.1     INTRODUCTION

It is assumed that only the approximately 2 million tonnes of oxide tailings will be reclaimed from the tailings dam for re-treatment for the recovery of silver and gold. The approximately 3 million tonnes of sulphide material will be moved from its present location and dumped at a new site near the proposed site of the heap leach plant. The sulphide tailings will therefore be loaded separately into trucks and dumped in a suitable newly created sulphide tailings deposit facility. Although it has been estimated that there is approximately 3 million tonnes of sulphide tailings, it is not certain how much of this material will be moved to the new sulphide tailings dam site. Some of the sulphide tailings could be used in the building of the heap leach pad and its facilities, but no quantities have been estimated at this stage.

16.4.2   PROCESS DESCRIPTION

The following is a conceptual description of the heap leach operation of the oxide tailings material.

The plant will be operated on a 24 hours per day, 365 days per year basis, with an overall utilisation of 90%. For a treatment rate of 0.5 million tonnes per year, this would be equivalent to a throughput rate of 1370 tonnes per day [t/d], and equivalent to 63.4 tonnes per hour [t/hr] at 90% running time. This will give an overall project duration of 4 years. This four year period will exclude the time required for site establishment and remediation of the heap after the heap leaching process has been completed.

A slusher-scraper system will be used to recover both the sulphide and oxide tailings material. A slusher-scraper method was selected for recovering the tailings because of the high moisture content of the tailings in certain areas of the dam which would make vehicular access dangerous. The sulphide tailings will be scraped into a heap and a front-end loader will be used to load the material onto a 40 t capacity truck. The truck will deposit the material at the site chosen for the sulphide tailings stockpile, namely the sulphide stockpile which will be situated immediately to the north-east of the heap leach pad. Another slusher-scraper will be used for the recovery of the oxide tailings material. The oxide tailings will similarly be scraped into a heap for loading onto trucks using a front-end loader. The 40 t capacity front-end loader will load the oxide tailings material into a dump truck which will then dump the material into a 160 t capacity tailings hopper.

A dribble chute will feed the tailings from the hopper onto a conveyor belt. Cement and lime will be added to the tailings under controlled addition rate conditions. Although some operations add solid dry, flake cyanide to the agglomerator feed material, this option will not be exercised in this case. The cement and lime will be added from their respective bulk storage silos. A 50 t capacity cement storage silo equipped with a dust collection filter and a cement blower will be required, as well as a 30 t capacity lime storage silo similarly equipped with a dust collection filter and a lime blower. Each reagent delivery system will be controlled by a weightometer prior to feeding the tailings material conveyor belt feeding the agglomerator drum. The design treatment rate will be 63.4 t/hr of tailings material with an average moisture content of 10%. Water, or barren solution, will be added to the agglomerator to provide for an overall moisture content of about 12.5 to 15% to the leach pad feed material. Two one- tonne capacity cyanide mixing and storage tanks will be positioned at the Merrill-Crowe facility. Cyanide will be diluted to 20% strength, and then injected into the solution distribution system going to the heap.

Avino Silver and Gold Mines Ltd.	41	0551920100-REP-R0002-01
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The agglomerator will be a drum type unit with a diameter of 2.0 m and a length of 6.0 m rotating at 10.5 rpm and with a slope of 4 degrees. Agglomerated material will be discharged onto a conveyor belt, then on to a series of conveyor belts, and then deposited on the heap leach pad. A curing time of 5 days will be allowed before spraying of the agglomerates with cyanide-bearing leach solution commences.

There will be one leach pad only. The leach heap dimensions will be an estimated 243 m wide and 282 m long and includes a surrounding berm of 6.5 m in width. There will be four lifts over the four year treatment period. Each lift will be 6.5 m high giving the heap an overall height of 26 m.

The heap leach process will operate with three solution ponds. Solution from the Barren Solution Pond will be pumped to the leach heap. Concentrated cyanide solution will be added to the Barren Solution Pond where it will be mixed to give a controlled cyanide concentration of about 0.5 g/l NaCN strength. The pH will be maintained at 10.5. This solution will be distributed over the leach pad using irrigation pipes and drips for an overall solution feeding rate of about 7.3 l/hr/m² [0.002 l/sec/m²]. A total leaching duration of 130 days will be allowed, followed by a wash/rinse cycle of 7 days resulting in a total loading, leaching and rinsing cycle of 142 days.

The overall life of a pad on a per lift basis is 365 days. The total calculated amount of area of pad under irrigation per day will be 22454 m², with 1209 m² being rinsed every day. The calculated volume of solution pumped to the heap will be 173 m³/hr of which a nominal 9 m³/hr will be rinse solution. A total solution evaporation loss of 10% is assumed.

The 173 m³/hr pregnant solution collected from the leach pad will be directed to the Pregnant Solution Pond. The solution from the Pregnant Solution Pond will be pumped to the Merrill-Crowe plant for silver and gold recovery by precipitation with zinc dust and filtration of the precipitate. The barren solution will then be returned to the Barren Solution Pond. Solution from the Pregnant Solution Pond can overflow into the Barren Solution Pond should this be required. Solution from the Barren Solution Pond can also overflow into the Overflow Solution Pond. This Overflow Solution Pond will also collect excess water and drainage solution from the heaps and the plant environs. The Overflow Solution Pond will also supply make-up water to the process by pumping the water back to the Barren Solution Pond. Alternatively, excess solution from this pond will be treated in the Effluent Treatment section prior to discharge in the river. Water drawn from this pond will be treated with calcium hypochlorite in an agitated treatment tank to reduce the cyanide levels to acceptable limits prior to discharging this water to the environment, or re-using this water as process water.

The Merrill-Crowe section will receive the pregnant solution, which will be pumped to the clarifier filter together with filter aid pre-coat and body feed. The slurry from the backwash cycles will be pumped to an inactive part of leach heap. The clarified solution will be pumped to the de-aeration tower where the solution will be de-oxygenated and a slurry of zinc dust, lead nitrate, cyanide and filter aid will be pumped into the de-aerated solution after the towers but ahead of the precipitate filters. The zinc dust, lead nitrate and filter aid will be made up into a slurry at the required dosage rate in the precipitate mixing tank and cyanide will be added as needed. The cementation reaction occurs at the point of introduction of the slurry to the de-aerated solution. This reaction normally requires about 2 to 5 minutes for completion. The reaction should be complete by the time the now-barren solution exits the precipitate filter. The barren solution will flow into the Barren Solution Pond where the pH will be adjusted to 10.5 with lime if necessary and then be pumped back to the heap for leaching.

Avino Silver and Gold Mines Ltd.	42	0551920100-REP-R0002-01
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The addition of zinc dust has been calculated on the basis of 8 grams of zinc dust per 1 gram of silver plus gold in order to ensure that the cementation reaction will be driven to completion. Although precipitation efficiencies are normally considered to be higher than 99.5%, in this case 96% has been selected since no testwork has been conducted on pregnant solution from this material. The cyanide concentration of the pregnant solution should be a minimum of about 100 mg/l as free cyanide, and will be monitored on a regular basis. The lead nitrate addition will be calculated on the basis of about 2 mg/l of solution and will be added to improve the precipitation efficiency. About 50% of the total amount of the lead nitrate to be added will be added to the pregnant solution prior to the clarifier filter in order to precipitate detrimental impurities. These impurities will then be removed in the clarifier filter. Although no anti-scalant reagents have been included in the study, any reagents of this nature should be tested to determine its effect on the precipitation efficiency.

The silver-rich precipitate will contain the gold and excess zinc. This precipitate from the filter press will be dried in an oven prior to bagging and its subsequent despatch to a smelter. It is anticipated that the total metal precipitate production per day will be about 461 kg/d (dry basis). The silver and gold content will be about 20%, or 92.2 kg (silver and gold).

16.4.3   PROCESS DESIGN CRITERIA

The process design criteria are attached as a separate section in Appendix D.

 16.4.4  CAPITAL AND OPERATING COST ESTIMATES

The estimated capital cost for the construction of the heap leach pad and the treatment facilities has been calculated to be US $16.2 million, and the estimated operating costs are US $8.64/ t. These cost items are described in the following sections and the details are given in the appendix.

Capital Cost Estimate

The capital cost for the Avino Mines heap leach project has been developed on the basis of the treatment of 1370 t/d, or 500,000 tonnes per year, of oxide tailings. The capital cost estimate has been assessed at US $16.2 million. Recognising that the duration of the project is only 4 years, the costs of the equipment and infrastructure have been kept as low as realistically possible. However, the capital costs are based on new equipment. The capital costs are summarised in Table 16.19.

Avino Silver and Gold Mines Ltd.	43	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

WARDROP

Table 16.19 Capital Cost Estimate - Summary

Item/ Description	Total Cost, US $
Mining, Agglomeration, and Pad Loading	2,099,533
Process Facilities	1,421,211
Reagents / Auxiliary Services	177,155
Buildings	653,938
Leach Pad and Infrastructure	4,159,464
Power Supply and Distribution	1,069,873
Total Direct Costs	9,581,173
EPCM, QA and Vendor Representatives	1,369,344
Freight and Construction Indirects	1,980,812
Contingency	3,232,832
Total Project Capital Cost Estimate	16,164,162

Mining, Agglomeration and Pad Loading

These costs include the facilities required for the recovery of the tailings using separate slushers for both the oxide and sulphide tailings material. Three refurbished slushers have been included in the estimate to meet the production requirements. It also includes the facilities required for the loading of the tailings into a bin to feed the conveyor to the agglomerator, and includes the agglomerator and its structural supports as well as the ancillary equipment. This section also includes the lime and cement silos.

Process Facilities

The costs in this section includes the various items of equipment, the tanks and their attendant pumps and agitators [if equipped], the Merrill-Crowe circuit [supplied as a modular package unit] and other miscellaneous process-related equipment. The process equipment is estimated as new cost items.

Reagents and Auxiliary Services

The costs derived for this section includes reagent tanks and related equipment as well as civil construction costs. Water will be supplied from the existing sources, namely from the dams and/or the wells. The costs shown for the fresh water supply includes the refurbishing of the equipment and pumps. Safety items related to reagent handling have also been included.

Buildings

The existing buildings and offices of the Avino Mine will be utilised for the project. An allowance has been included for the refurbishment of these facilities. No costs have been allocated for the truck shop since it is intended to have a transport contractor to provide all the transportation needs for the project. An allowance has been included for the procurement/refurbishing of laboratory equipment.

Avino Silver and Gold Mines Ltd.	44	0551920100-REP-R0002-01
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Leach Pad and Infrastructure

The civil construction costs of upgrading the roads and constructing the leach pad and ponds are given in this section. The leach pad will be constructed according to standard practise with liners and a leak detection system. The ponds will all be lined. Also included is the cost of fencing off the plant area, the telephone system, sewage disposal, water supply and treatment, and fuel storage facilities. The existing fuel storage facilities will be used but this will require refurbishing and this cost has been provided in this section.

Power Supply and Distribution

The refurbishing and expanding of the existing electrical power supply system, and provision for lighting, has been included in this section. It also includes the provision of power to the slushers, the agglomerator area and the Merrill-Crowe section.

Indirect Costs

Indirect costs have been included as costs associated with construction services, consulting services, spare parts and freight. A contingency of 25% has been included in the indirect costs.

No sunk costs, owner’s costs, taxes, or insurance costs have been added to the capital cost estimates as detailed above.

Avino Silver and Gold Mines Ltd.	45	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

WARDROP

Operating Cost Estimate

The operating costs have been determined on the basis of a treatment rate of 500,000 tonnes per year. The anticipated costs of reagents have been given in Table 16.20 and are based on calculations given in the process design criteria. The estimated plant complement and labour cost has been developed in Table 16.21. The costs of reagents have been obtained from suppliers in Mexico. The labour costs are based on available labour rates for Mexico. The labour requirements have been obtained from Avino Mines. It is not clear if the shift complement is based on three 8-hour shifts per day or two 12-hour shifts. If the staff is based on three 8-hour shifts, there is no provision for a relief shift. The daytime only labour is also indicated in Table 16.21.

Table 16.20 Estimated Cost of Reagents

Reagent	Reagent Consumption, kg/t	Reagent Cost, US $/ kg	Monthly Cost, US $
Cement	7.194	0.15	44,963
Lime	4.531	0.08	15,103
Cyanide	0.928	1.95	75,401
Zinc Dust	0.768	2.05	65,601
Lead Nitrate	0.154	0.21	1,348
Filter Aid	0.100	0.79	3,292
Pre-Coat	0.100	0.79	3,292
Monthly Cost of Reagents in US $:			209,000
Cost, US $/ t:			5.02

Avino Silver and Gold Mines Ltd.	46	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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Table 16.21 Plant Staff Complement And Labour Cost - 3-Shift Basis

Job Designation	No. of Shifts	Persons per shift	Total Persons	Monthly Salary (Pesos)	Monthly Cost
Plant Superintendent	1	1	1	30,000	30,000
Engineering and Planning Manager	1	1	1	25,000	25,000
Administration Manager	1	1	1	20,000	20,000
Shift Foreman	3	1	3	15,000	45,000
Slusher Operator	3	3	9	15,000	135,000
Plant Operator: Agglomerator	3	2	6	6,000	36,000
Plant Operator: Conveyors	3	1	3	6,000	18,000
Plant Operator: Merrill- Crowe	3	2	6	6,000	36,000
Day Crew Reagents	1	2	2	6,000	12,000
Day Crew Heap Piping	1	10	10	6,000	60,000
Maintenance Crew	1	4	4	15,000	60,000
Assay Supervision	1	1	1	15,000	15,000
Assayers	3	2	6	6,000	36,000
First Aid Attendants	1	1	1	15,000	15,000
Office Clerk	1	1	1	7,000	7,000
Warehouse Staff	1	2	2	10,000	20,000
Computer Technician	1	1	1	8,000	8,000
Environmental Supervisor	1	1	1	20,000	20,000
Purchasing Agent	1	1	1	20,000	20,000
Security	3	1	3	6,000	18,000
Total:			63		636,000
Exchange Rate: US $/Mexican Pesos					$0.09517
Monthly Costs in US $:					$60,525
Cost, US $/ t:					$1.45

The estimated cost of reagents has been calculated to be US $5.02/t treated while the labour cost has been estimated to be US $1.45/t treated. The total operating cost has been estimated to be US $8.64/t treated. The details are given in Table 16.22.

Avino Silver and Gold Mines Ltd.	47	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

WARDROP

Table 16.22 Estimated Operating Costs

Item / Description	Cost, US $/ t	Remarks
Transport: Tailings to Pad & Sulphide Dump	1.03	$1.66 per m3
Power	0.69	$0.069 per kWhr
Water	0.04	$0.14 per m3
Reagents	5.02	Table 20
Labour	1.45	Table 21
Miscellaneous	0.41	5% of total
Total Operating Cost, US $/ t	8.64	Estimated Total

Avino Silver and Gold Mines Ltd.	48	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

WARDROP

  17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

During the period from 1976 to 2001, the Avino Mine was in continuous operation and a nominal 5 million tonnes of ore was treated by the mill during that period of time. Ore production prior to 1993 [1992 has also been mentioned] was described as “oxide” and was mined from the open pit. As indicated in the MMI report dated May 2005, and the subsequently revised reports [References #4, #5, #6 and #7], just over 2 million tonnes of oxide ore was apparently treated.

From 1993 to 2001, ore described as “sulphide” was mined and processed. MMI records indicate that a nominal 3 million tonnes of ore were treated. Since the sulphide tailings material was not a part of this technical study, no values have been assigned to the sulphide tailings during this study except in the section where a conceptual economic value was required for purposes of demonstration. In this case, the head grade values used were those originating from the samples used for the limited metallurgical testwork program undertaken at PRA under the direction of MMI [Reference #3]. These assay values should in no way be construed as representing the assay values of the entire sulphide tailings deposited in the tailings dam. An independent and detailed sampling campaign would be required to evaluate the metal assays and content of the sulphide tailings.

As reported in Section 16.1.1, MMI estimated the volumes/tonnages, as well as the silver and gold grades of the oxide tailings [the inferred resource] to be as given in Table 16.1. This table is reproduced here as Table 17.1 for purposes of discussion.

Table 17.1 Oxide Tailings Dam Data

Source	Tonnes, t	Assays, g/t		Bulk Density, g/cm3
		Ag	Au
Cia Minera, 1990	2,092,178	93.0	0.50	1.605
MMI, 2005	2,091,074	95.5	0.53	1.605

Of interest is that the tonnage of oxide tailings material available for treatment and given by the MMI reports as arising from the Cia Minera sampling program of 1990 is 2,092,178 tonnes. This is the exact tonnage number given from a data sheet provided by Avino Mines and purporting to summarise the data from the 1990 sampling program. This summary sheet has been included in Appendix J and is titled ‘Bloque de Reservas, Presa de Jales’.

If the data presented on this sheet, ‘Bloque de Reservas, Presa de Jales’, was generated from the 1990 sampling program, then it would have excluded the oxide tailings deposited during the rest of that year [1990], as well as 1991, 1992 and the part of 1993 until the sulphide material became available. That would render the Cia Minera 1990 data as incomplete with respect to both tonnage calculations and tailings grade. The methodology adopted by Cia Minera to establish the values given in the data sheet, namely using a simplified block model, is also considered to be not technically valid but probably provides a reasonable estimation given the drilling pattern used for this exercise.

Estimated tonnage calculations by MMI, using string traverse and an assumed triangular plan, as well as an arithmetic check of the 1990 data using block volumes, gave reasonable confirmation that between 2.0 and 2.3 million tonnes of oxide tailings were present in the tailings dam [Reference #7]. However, no indication has been provided in the MMI reports as to the methodology used by MMI to establish the MMI 2005 value of 2,091,074 tonnes of oxide material present in the tailings dam.

Wardrop did not independently verify the tonnage or grade of the tailings deposit.

(See the report “A Tailings Resource” by Bryan Slim in Appendix H.)

Avino Silver and Gold Mines Ltd.	49	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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18.0 OTHER RELEVANT DATA AND INFORMATION

18.1      FINANCIAL MODEL

The MMI Technical Report “A Tailings Resource”, dated October 2005, was produced for Avino Silver and Gold Mines Ltd, Cia Minera Mexicana, Durango, Mexico, of MineStart Management Inc., North Vancouver, B.C., Canada, and submitted to the Canadian Securities Regulatory Authorities (CSRA) [Reference #7]. Deficiencies were noted by the CSRA and the report was returned to MMI for revision. The required amendments were subsequently made to the document and re-submitted to the CSRA. In its final version, the MMI report refers to the oxide tailings as an “inferred resource”. This term “inferred resource” is expanded upon in the following paragraph.

Section 7.2 from the MMI Technical Report of October 2005, is quoted verbatim below.

Start of quotation:

“7.2 THE OPINION

While the assays and their trends and for the estimates of volumes give reasonable indications of continuity of grades into those areas not fully sampled we note the limitations of non-confirmed metallurgical characterisation and the need for further testing in those areas not examined.

These limiting descriptions of the tailings means they are suitable nor ready for planning or economic evaluation neither [sic] and thus in our professional opinion only merit classification as an inferred resource for which we assign the dimensions of 2M t at the statistical mean grades of the 1990 sampling of 95 g/t silver and 0.5 g/t gold for the combined middle and lower benches. The estimate excludes oxide material lying coincident with and at the bottom of the upper bench. No assays from 2004 were included in the inferred resource classification since the former were an incomplete population and would skew the estimate”

End of quotation.

A copy of this document is attached in this report in the appendix.

In addition, it is the opinion of the author of the MMI Technical Report, as quoted above, that an economic evaluation can not be conducted at this stage.

The above quotation from the MMI Technical Report essentially states that:

· The extent of the sampling program was limited

· The metallurgical samples taken from the tailings deposit for testing did not represent

the complete deposit

· The metallurgical test results were not confirmed

· The volume of the tailings deposit was not accurately determined.

The lack of specifically accurate information therefore precludes the establishing of an accurate and detailed economic evaluation of the deposit and implies that the following steps need to be taken in order to complete a feasibility study. These steps include the following.

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· A detailed and comprehensive sampling and assaying program must be undertaken to define the inherent value of both the oxide and sulphide tailings material in the tailings  dam

·          A detailed evaluation of the volume of the tailings deposit must be undertaken to define the individual oxide and sulphide tailings volume [tonnage], and the combined volume, available for treatment

·           A comprehensive metallurgical testwork program must be conducted on representative samples taken from the tailings deposit in order to quantify the recovery of silver and gold from the tailings material

·           The metallurgical testwork should quantify the leaching time required, and give more accurate reagent consumption values and precipitation efficiency data

·           The metallurgical test results should be reproducible.

However, on a conceptual basis, two financial scenarios were produced, namely processing the oxide tailings over a 4-year span versus processing the tailings over a 2-year span. The silver price was varied between $7.00 and $9.00 per ounce with the gold price fixed at $500 per ounce, followed by the silver price fixed at $8.00 per ounce with the gold price changing between $450 t0 $600 per ounce (as USA dollar values).

For this comparison, the following parameters were used:

·           The Design Criteria silver and gold recovery rates were used to estimate metal recovery;

·           The Design Criteria 130-day leaching cycle was used to estimate the rate of recovery;

·           One month delay between the time ore was stacked and when leaching began;

·           30% of the total recovered metal was recovered in the first month;

·           20% of the total recovered metal was estimated for months 2 though 4;

·           10% of the total recovered metal was estimated for month 5; and,

·           Leaching continued for 5 months after the ore was exhausted, at half the operating cost.

The 4-year mine life capital cost was estimated at $16,200,000; the 2-year mine life plant was expanded at only 1.4 times over the smaller plant. No salvage value for either plant is included for this comparison.

The Net Present Value (NPV) for the 4-year minelife varied from $3.7 million to $10.0 million (US) based on metal prices, with the Internal Rate of Return (IRR) ranging from 20% to 35%. The NPV for a two-year minelife ranged from $0.2 million to $5.6 million for the same metal prices, with the IRR ranging from 11% to 31%. All profits are before taxes, royalties or similar.

The NPV and IRR tables are shown in Table 18.1 and Table 18.2.

Avino Silver and Gold Mines Ltd.	51	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

WARDROP

Table 18.1 Potential Net Present Value (NPV)

Avino Silver and Gold
Potential Net Present Value
Mine life		4 years		2 years
Estimated Capital:		$16,200,000		$22,680,000
Metal Price			Gold price fixed at $500 per ounce
Silver
$7.00		$3,722,727		$169,543
$7.50		$5,283,952		$1,879,119
$8.00		$6,845,176		$3,588,695
$8.50		$8,406,401		$5,298,271
$9.00		$9,967,625		$7,007,847
Gold			Silver price fixed at $8.00 per ounce
$450		$5,908,710		$2,563,244
$500		$6,845,176		$3,588,695
$550		$7,781,642		$4,614,146
$600		$8,718,108		$5,639,598

Avino Silver and Gold Mines Ltd.	52	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

WARDROP

Table 18.2 Potential Internal Rate of Return (IRR)

Avino Silver and Gold
Potential Internal Rate of Return
Mine Life		4 years		2 years
Estimated Capital:		$16,200,000		$22,680,000
Metal Price			Gold price fixed at $500 per ounce
Silver
$7.00		20%		11%
$7.50		24%		16%
$8.00		28%		21%
$8.50		32%		26%
$9.00		35%		31%
Gold			Silver price fixed at $8.00 per ounce
$450		26%		18%
$500		28%		21%
$550		30%		24%
$600		32%		27%

Avino Silver and Gold Mines Ltd.	53	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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18.2      COMMENTS REGARDING THE COST ESTIMATES

It can be seen that the calculated capital and operating cost estimates are higher than the estimates previously given in Section 16.3.6 and 16.3.7 which were referencing average industry costs. This possibly implies that there are costs which could be reduced in a future Feasibility Study for Avino Mines. The reagent consumption values, although considered to be reasonable estimates, are based on the outcome of one test only and all these numbers therefore require confirmation, while the staff complement for the proposed heap leach treatment facility should also be updated, and the work schedules require confirmation. It is anticipated that a relief [4^th shift] may have to be factored into the staff structure. A review of the processing section may also result in some power savings.

In summary, it should be noted that the above cost estimates are based on the best available information and assumptions made, and these have all been discussed in the report. The costs are very sensitive to the price of silver and gold, the extraction recoveries used in developing the cost analysis, reagent consumption values (especially cyanide and zinc dust), and the head grade of the material to be treated. The availability of more detailed and accurate metallurgical information will assist in providing more accurate information for the Feasibility Study phase of the tailings treatment project.

The capital costs are based on purchasing new equipment for the project. Selecting good and suitable used equipment [if available] could reduce the overall cost of construction.

Avino Silver and Gold Mines Ltd.	54	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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19.0 INTERPRETATION AND CONCLUSIONS

A possible process design was developed to recover the silver and gold present in the oxide tailings dam material at the Avino Mine per the MMI report dated October 2005.

The implied values contained in the oxide tailings dam are approximately US $47.3 million, based on US $8.00 per ounce of silver and US $500 per ounce of gold. The silver values are approximately three times higher than the gold values.

The capital cost to design and construct a 500,000 tonne per annum agglomeration heap leach operation treating tailings material is estimated to be US $16.2 million.

Based on information in the MMI report and using current reagent costs, plant operating and maintenance costs are estimated to average US $8.64 per tonne treated. It must be noted that no sulphide tailings relocation costs are included in this figure because of insufficient information to calculate the volumes of both the oxide and sulphide tailings.

These costs are summarized in Table 19.1 below.

Table 19.1 Summary

Avino Silver and Gold
Implied Values contained in Oxide Tailings	US $47.3 million
Capital Cost for 500,000 tonne per annum agglomeration/ heap leach operation	US $16.2 million
Estimated operating cost per tonne of tailings treated     (stripping costs not included)	US $8.64
Estimated Net Revenue using $8.00 per ounce Silver and                 $5.00 per ounce Gold	US $31,423,182

Avino Silver and Gold Mines Ltd.	55	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

WARDROP

20.0 RECOMMENDATIONS

All of the information used in this report should be verified, including but not limited to the following:

·    Survey the complete tailings dam, as well as specifically the oxide section and the sulphide section of the tailings deposit.

·    Drill the surface of both tailings dam areas to determine the volumes and bulk density of each of the oxide and sulphide tailings material.

·    Take sufficient amounts of samples from both oxide and sulphide tailings to obtain representative samples for assay and metallurgical testwork to confirm the grade of the deposit and the recovery of silver and gold from the heap leach process.

·    Use the metallurgical results from the testwork program to confirm/define the duration of leaching on the pad, the reagent consumption values and the silver and gold precipitation efficiencies.

·    Use the metallurgical results from the metallurgical testwork program to develop a heap leach flowsheet.

·    Based on accurate assay and reproducible metallurgical testwork data, develop a financial model for the treatment of the oxide tailings dam material, the sulphide tailings dam material, and for the treatment of both oxide and sulphide tailings material.

·    Design the plant treatment rate based on current metal prices and assess the sensitivity of the project in relation to the major variables.

Avino Silver and Gold Mines Ltd.	56	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

WARDROP

21.0 REFERENCES

REFERENCES.

1. “Flotation and Cyanidation Scoping Tests and Specific Gravity”, Process Research Associates Ltd., Project No. 0302303. Report by John Huang, 28 March 2003. Report to Bryan Slim, MineStart Management Inc.

2. “Tailings Valuation, Avino Silver and Gold Mines Ltd., Cia Minera Mexicana, Durango, Mexico”, MineStart Management Inc. Report by Bryan Slim, November 2003.

3. “Metallurgical Test Work on Avino Tailings, Durango, Mexico”, Process Research Associates Ltd., Project Number 0406407. Report by John Huang and Gie Tan, 28 March 2005. Report to Bryan Slim, MineStart Management Inc.

4. “Preliminary Feasibility, Avino Silver and Gold Mines Ltd., Cia Minera Mexicana, Durango, Mexico”, MineStart Management Inc. Report by Bryan Slim, May 2005.

5. “Tailings Valuation, Avino Silver and Gold Mines Ltd., Cia Minera Mexicana, Durango, Mexico”, MineStart Management Inc. Report by Bryan Slim, May 2005.

6. “A Tailings Resource, Avino Silver and Gold Mines Ltd., Cia Minera Mexicana, Durango, Mexico”, MineStart Management Inc. Report by Bryan Slim, July 2005.

7. “A Tailings Resource, Avino Silver and Gold Mines Ltd., Cia Minera Mexicana, Durango, Mexico”, MineStart Management Inc. Report by Bryan Slim, October 2005.

Avino Silver and Gold Mines Ltd.	57	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

WARDROP

22.0 DATE AND SIGNATURE PAGE

CERTIFICATE OF PROFESSIONAL QUALIFICATIONS

I, Richard C. Alexander, P. Eng., do hereby certify that: as the author of the “Avino Silver & Gold Mines Form 43-101F1 Technical Report: Tailings Retreatment - Process Options” dated March 2006, I hereby make the following statements:

· I am an Engineering Manager with Wardrop Engineering Inc. with a business address at #905 - 1130 West Pender St, Vancouver, BC, V6E 4A4.

· I am a graduate of the University of Alberta (1985) with a degree in B.Sc. (Mechanical Engineering).

· I am a registered Professional Engineer in the Province of British Columbia.

· I have practiced my profession of engineering, construction management and project management for 20 years.

· I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that, by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

· I am responsible for the preparation of this technical report titled “Avino Silver & Gold Mines Form 43-101F1 Technical Report: Tailings Retreatment - Process Options” dated March 2006.

· I have visited the Avino Silver & Gold Mines Project Property in December 2005.

· I have had no prior involvement with the property that is the subject of the Technical Report.

· I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

· I am independent of the Issuer applying the tests set out in Section 1.5 of National Instrument 43-101.

· I have read National Instrument 43-101 and the Technical Report has been prepared in compliance with National Instrument 43-101 and Form 43-101F1.

Signed and dated this 31 day of March 2006 at Vancouver, B.C.

________________________________

Rick Alexander, P. Eng.

Engineering Manager

Wardrop Engineering Inc.

Avino Silver and Gold Mines Ltd.	58	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

WARDROP

  23.0 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES

This section is not applicable.

Avino Silver and Gold Mines Ltd.	59	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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24.0 ILLUSTRATIONS

See Appendix A.

Avino Silver and Gold Mines Ltd.	60	0551920100-REP-R0002-01
Form 43- 101F1 Technical Report: Tailings Retreatment - Process Options

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Appendix A

General Arrangements, Drawings & Sections

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Appendix B

Process Flowsheet

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Appendix C

Electrical Single Line Diagram

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Appendix D

Process Design Criteria

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Appendix E

Capital Costs

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Appendix F

Slusher System Description & Design Calculations

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Appendix G

Preliminary Economic Evaluation

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Appendix H

“A Tailings Resource”

by Bryan Slim

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Appendix I

PRA Tailings Report

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Appendix J

Cia Minera 1990 Sampling Program Data (from Avino Mines)