The Caylloma Mine is within the historical mining district of Caylloma, northwest of the Caylloma caldera complex and southwest of the Chonta caldera complex. Host rocks at the Caylloma Mine are volcanic in nature, belonging to the Tacaza Group. Mineralization is in the form of low to intermediate sulfidation epithermal vein systems.

Epithermal veins at the Caylloma Mine are characterized by minerals such as pyrite, sphalerite, galena, chalcopyrite, marcasite, native gold, stibnite, argentopyrite and silver-bearing sulfosalts (tetrahedrite, polybasite, pyrargyrite, stephanite, stromeyerite, jalpite, miargyrite and bournonite). These are accompanied by gangue minerals, such as quartz, rhodonite, rhodochrosite, johannsenite (manganese pyroxene) and calcite.

Mineralization vein systems occurs in steeply dipping ore shoots ranging up to several hundred meters (“m”) long with vertical extents of over 400 m. Veins range in thickness from a few centimeters (“cm”) to 20 m, averaging approximately 1.5 m for silver veins and 2.5 m for polymetallic veins.

The mining method applied in the exploitation of the two main veins (Animas and Bateas) is overhand cut-and-fill using either mechanized, semi-mechanized or conventional extraction methods. All mining is undertaken in a southwest to northeast direction following the strike of the veins. Production capacity at the mine is 1,430 tonnes per day (tpd) since the upgrade to the plant in March 2016 but has the potential of achieveing 1,500 tpd.

Mechanized mining utilizes a jumbo drill rig and scoop tram for loading. The ore haulage is performed by a combination of locomotives and trucks. Rock support is applied through rock bolts and shotcrete. The average mining width ranges between 3.5 m and 17 m. Mechanized mining is regarded as only being suitable for the Animas vein based on the geological structure and geotechnical studies. Approximately 90 percent of production comes from the Animas vein.

Semi-mechanized mining is performed using handheld drilling equipment (jacklegs) and scoops for loading. Ore haulage is performed by a combination of locomotives and dump trucks. Rock support is supplied using rock bolts installed using manual drilling and installation techniques. Semi-mechanized mining is applied to narrow veins with average widths between 0.8 m and 2.0 m.

Conventional mining is performed using handheld drilling equipment (jacklegs) and scrapers for loading. The ore haulage is done with trains and rail system and the support is applied with rock bolts in manual form. This system is applied in narrow veins with average widths between 0.5 m and 0.8 m.

The underground mine is operated by a mining contractor selected by Minera Bateas based on a competitive bidding process. The scope of work for the mining contractor generally includes mine decline and raise development, stope preparation development, stoping, backfilling, and all related services required for the operation of a 1,500 tpd narrow vein silver-gold-lead-zinc mine.

The Minera Bateas processing plant is a typical flotation operation and consists of five stages: crushing; milling; flotation; thickening and filtering and tailings disposal.

The Caylloma Mine concentrator plant resumed operations in September 2006, treating 600 tpd of polymetallic mineral. Capacity increased progressively, with the installation of a 1.8 m by 2.4 m ball mill in 2009 the plant reached a treatment capacity of 1,300 tpd, and with the installation of two Derrick Stack Sizer vibrating wet screens the plant achieved a treatment capacity of 1,500 tpd at the end of March 2016, although this has since been reduced to 1,430 tpd for the rest of 2016. The treatment process is differential flotation. Initially, two concentrates were obtained: lead-silver and zinc. From late 2009 to January of 2011, a copper-silver concentrate was also produced, but due to unfavorable commercial terms the production of copper concentrate was suspended and the copper circuit put on standby.

The crushing process is fed from the 10,000 t capacity stock pile used for ore storage and blending. The process commences with feed to the coarse hopper, which has a 450 t active capacity with 30 cm separation grates. The mineral is extracted from the coarse hopper through the apron feeder that feeds the vibrating grizzly with variable separation that, in turn, feeds the Kurimoto jaw crusher, resulting in a product size varying between 76 mm and 90 mm. The mineral is transported on two conveyor belts (1-A and 2-A) to the twodeck vibrating 6 by 14 foot screen. The screen’s undersize is fed to the stockpile through belts 3-A, 17, 3, 4, and 15 with the oversize going to a Sandvik CH-420 secondary crusher through belt 4-A, the product of which goes to the two-deck vibrating 5 by 14 foot screen through belts 1 and 18, the undersize of this screen feeds the stockpile through belts 3, 4 and 5. The oversize is fed through conveyor belt 2 to the Sandvik CH-430 tertiary crusher, the discharge of which returns to belt 1, closing the circuit.

Additionally, there is a standby primary crushing circuit that starts at a 100 t capacity coarse hopper. From the hopper, the mineral is fed to a Kueken 24 by 36 inch jaw crusher through a Ross chain feeder. The discharge from this crusher is transported via conveyors 19 and 20 to conveyor 2-A. There are three permanent magnets and one electromagnet on the conveyors to prevent the entry of tramp iron.

The grinding circuit has two stages. The primary stage operates in an open circuit, consisting of two ball mills (Comesa 2.4 m by 3.0 m and a Denver 2.1 m by 2.1 m). The secondary stage operates in closed circuit and consists of three ball mills, a Magensa 1.8 m by 1.8 m, a Hardinge 2.4 m by 0.9 m and a Liberty 1.8 m by 2.4 m. The final product of the grinding circuit is 60 percent passing 75 microns.

The Comesa and Denver primary grinding mills are fed independently by conveyor belts. The Comesa primary mill operates with the Magensa and Libertad secondary mills. The Comesa mill discharge feeds a flash cell (SK 240) with concentrate from the flash cell being sent to the lead thickener. Tailings are fed to a 6 by 6 inch horizontal pump which in turn feeds the Derrick Stack Sizer. The undersize goes to the flotation circuit and the oversize feeds the three secondary ball mills.

The Denver primary ball mill operates with the Hardinge secondary ball mill. Discharge from this mill feeds a 6 by 4 inch horizontal pump, which in turn feeds the D-15 cyclone. The cyclone’s overflow goes to the flotation circuit and the underflow returns to the three secondary ball mills.

Metallurgical treatment is through a process of differential flotation; the first step is the flotation of lead-silver followed by zinc flotation.

Lead-silver flotation circuit.
The D-15 cyclone overflow and the Derrick Stack Sizers undersize are fed to a conditioner before going to the TC-20 unit cell of 20 m3 capacity. The unit cell tailings are fed to three TC-20 rougher cells. The unit cell concentrate together with the rougher concentrate are fed to the primary cleaner cells, consisting of four 3 m3 OK-3 cells. The primary cleaner concentrate is fed to the secondary cleaner cells, consisting of three 3 m3 OK-3 cells. The secondary cleaner concentrate is fed to the tertiary cleaner cells, consisting of two 3 m3 OK-3 cells. Concentrate from the tertiary cleaner forms the final lead-silver concentrate. Tailings from the secondary and tertiary cleaner cells return to the head of the primary and secondary cleaner cells, respectively.

The rougher tailings feed the scavenger flotation bank, consisting of two TC-20 cells. The scavenger concentrate, as well as the primary cleaner tailings are pumped to join the Derrick Stack Sizers with oversize returning to the secondary grinding circuit. The scavenger tailings feed the zinc flotation circuit.

Zinc flotation circuit.
The lead-silver flotation tailings are sent to three conditioners (two 2.4 m by 2.4 m, one 3.0 m by 3.0 m). The conditioned pulp is fed to the zinc rougher flotation stage, consisting of six 8 m3 OK8U cells working in series. The rougher concentrate is fed to the cleaner flotation circuit, comprised of three stages consisting of five, three and two 2.8 m3 Sub-A30 cells for the primary, secondary and tertiary cleaner stages respectively. These stages work in series, the concentrate from the primary cleaner feeds the secondary cleaner and the concentrate of this feeds the tertiary cleaner. The concentrate from the latter is the final product from the zinc flotation circuit. The zinc concentrate goes through an automatic sampler and is then sent to the zinc thickener.

The rougher tailings feed the scavenger flotation circuit that is comprised of two 8 m3 OK8U cells. The scavenger concentrate is sent to a conditioner before returning it to the rougher circuit. The scavenger tailings are the final tailings of the whole process.

The flotation process achieves a lead concentrate containing an average of 55 percent lead and a zinc concentrate containing an average of 51 % zinc. Historical data show achievable metallurgical recoveries of 91 percent for lead, 81 percent for silver (in the lead concentrate) and 87 percent for zinc.

Concentrates thickening and filtration.
The lead-silver concentrate is thickened in an Outotec 9.0 m diameter thickener; the underflow is pumped to a 1.8 m diameter disc filter (six discs). The filtered lead concentrate contains on average 7.5 percent moisture.

The zinc concentrate is thickened in an Outotec 12.0 m diameter thickener; the underflow is pumped to a 1.8 m diameter disc filter (eight discs). The filtered zinc concentrate contains on average 9.0 percent moisture.

Each filtered concentrate is discharged into a covered temporary storage area from where it is loaded by a front-end loader into trucks for transport to the concentrate purchaser’s storage facilities in Matarani, Arequipa for the zinc concentrate and Callao, Lima for the lead-silver concentrate.

Tailings disposal.
Tailings from the concentration process are pumped and classified through cyclones. The underflow is accumulated in a temporary storage area for later transportation to the mine as hydraulic backfill. Approximately 35 percent of the tailings are used as backfill material in the mine.

The overflow is pumped to the tailings facility for final disposal. The water collected from the tailings impoundment is pumped back to the processing plant and reused in the process. Usage of the new tailings storage facility (N° 3) commenced in January 2013.