The oxide tailings deposit (see Figure 1.2 for location) comprises historic recovery plant residue material that was wasted from processing plants during the earlier period of open pit mining of the Avino Vein. The oxide tailings are partially covered by younger unconsolidated sulphide tailings on the northwest side.

The Avino tailings dam is located approximately 500 m west southwest of the main shaft to the old underground workings and 2.5 km southwest of the San Gonzalo Vein.

Within the tailings dam, there are three distinct benches:
* lower oxide bench
* middle oxide bench
* upper bench or sulphide bench.

Due to the historical processing sequence, the oxide tailings are primarily derived from weathered and oxidized rocks close to the surface on the Property, whereas the sulphide tailings are predominantly derived from material sourced at depth from the underground workings, below the weathered/leached zone.

In the oxide zone, mineralization is primarily hosted by the minerals argentite, bromargyrite, chalcopyrite, chalcocite, galena, sphalerite, bornite, native silver, gold, and native copper. Other minerals present in mineralized areas, but not hosting the metals of interest, 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.

The oxide tailings Mineral Resource will be mined/moved using a conventional truck/loader surface mining method. The production cycle consists of loading and trucking.

The production schedule has been developed for the oxide tailings based on a treatment rate of 500 kt/a. This would be equivalent to a throughput rate of 1,370 t/d. This will give an overall Project duration of approximately eight years. This eight-year period includes one year of pre-production and excludes the time required for remediation of the heap after the leaching process has been completed. Only oxide tailings will be considered for treatment, while sulphide materials will be considered waste.

The mining operations include loading and trucking. A 3.85 m3 rated (5.0 yd3) front-end loader will be used to load a 24 t articulated truck that will either deliver the sulphide tailings to the sulphide stockpile or the oxide tailings to the 160 t oxide tailings hopper. Once the hopper is filled, excess tailings will be stockpiled around the hopper to be loaded by the process plant group.

Certain areas of the tailings might contain high amounts of moisture that can lead to equipment getting stuck. To mitigate this challenge, wider, oversized tires with chains will be installed on the front-end loader. Also, the front-end loader bucket will be downsized to 3.06 m3 (4.0 yd3). This will lighten the load on the front tires preventing them from sinking into saturated material. The trucks will not enter the soft zones so there will be no modifications to the trucks.

The proposed treatment plant will consist of agglomeration and cyanide heap leaching, followed by a Merrill-Crowe process to recover silver and gold from pregnant solution.

For an oxide tailings treatment rate of 0.5 Mt/a, an equivalent throughput rate of 1,370 t/d or 63.4 t/h is required. This will give an overall Project duration of approximately seven years. This seven year period will exclude the time required for site establishment and remediation of the heap after the leaching process has been completed.

A dribble chute will feed the tailings from the hopper onto a conveyor belt. Cement and lime will be added to the tailings at controlled addition rates. 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 reagents to the tailings material conveyor belt feeding the agglomerator drum. The design treatment rate will be 63.4 t/h 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 approximately 12.5 to 15% to the leach pad feed material. Two 1 t capacity cyanide mixing and storage tanks will be positioned at the Merrill-Crowe facility. Cyanide preparation system will produce a cyanide solution with a strength of 20% sodium cyanide. The cyanide solution will then injected into the solution distribution system going to the heap pad and precipitation filter press.

The agglomerator will be a drum type unit with a diameter of 1.8 m and a length of 6.7 m rotating at 10.5 rpm and with a variable angle of 2.5, 5.0 or 7.5°. Agglomerated material will be discharged onto a conveyor belt, then on to a series of jump conveyors, and then deposited on the heap leach pad by a radial telestacker. A curing time of 5 d will be allowed before spraying of the agglomerates with cyanide-bearing leach solution commences.

There will be only one leach pad. The leach heap pad dimensions are estimated to be 288 m wide and 428 m long and includes a surrounding berm of 6.5 m in width. There will be four lifts over the seven-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:
* a barren solution pond
* a pregnant solution pond
* an event or overflow pond.

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 approximately 0.5 g/L sodium cyanide 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 approximately 7.3 L/h/m2 (0.002 L/s/m2). A total leaching duration of 130 d will be allowed, followed by a wash/rinse cycle of seven days resulting in a total loading, leaching and rinsing cycle of 142 d.

The total calculated amount of area of pad under irrigation per day will be approximately 22,000 m2, with 1,210 m2 being rinsed every day. The calculated volume of solution pumped to the heap will be 173 m3/h of which a nominal 9 m3/h will be rinse solution. A total solution evaporation loss of 10% is assumed.

The 173 m3/h 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 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 tower 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 approximately 2 to 5 minutes for completion. The reaction should be complete by the time the new-barren solution exits the precipitate filter to barren solution tank and from there it 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 pad for leaching after cyanide concentration is adjusted to approximately 0.5 g/L sodium cyanide.

The addition of zinc dust has been calculated on the basis of 10 g of zinc dust per 1 g 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 test work has been conducted on the 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 added to improve the precipitation efficiency and its dosage is based on approximately 2 mg lead nitrate per litre of solution. Approximately 50% of the total required amount of the lead nitrate will be added to the pregnant solution prior to the clarifier filter where impurities present in the solution will be removed by 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 which contains the gold and excess zinc will flow to the acid vat tank where the excess zinc can be dissolved by mixing with adequate amount of sulphuric acid and from there it will be pumped to the digest precipitate filter press. This precipitate from the filter press will be dried in an oven prior to being melted in a smelting furnace for doré production. It is anticipated that the total metal precipitate production per day will be approximately 420 kg (dry basis) with approximately 20% of silver and gold, or approximately 85 kg of silver and gold.