The Trinidad silver-gold deposit at the San Jose Mine is a typical low-sulfidation epithermal deposit according to the classification of Corbett (2002), having formed in a relatively low temperature, shallow crustal environment. The deposit is characterized by structurally controlled hydrothermal breccias, crackle breccias and quartz-carbonate veins hosting silver-gold mineralization plus trace to minor base metal mineralization.

Precious metal mineralization at the San Jose Mine is hosted by hydrothermal breccias, crackle breccias, quartz-carbonate veins and zones of sheeted and stockwork-like quartz-carbonate veins emplaced along steeply dipping north and north-northwest trending fault structures.

The Trinidad vein system (Tv) is emplaced in the footwall fault zone of the extensional system hosting the mineralized vein systems at San Jose. The Trinidad vein system strikes 355° and dips 70° to 80° to the east-northeast. The vein system ranges from less than 1 meter to locally over 15 meters in true width, with higher grade mineralization generally being present in zones with greater widths. Significant portions of the Trinidad structure are characterized by late black matrix silicified fault breccias with only trace to weak mineralization. Higher grade precious metal zones in the Trinidad vein system range up to approximately 1,300 g/t Ag Eq across the width of the vein. Combined copper, lead and zinc values are generally less than one percent but locally higher concentrations are present. At approximately the 1,100 meter elevation in the central portion of the Trinidad Deposit, four drill holes intercepted higher grade base metal mineralization with combined copper, lead, and zinc values ranging up to 21.6 percent across the width of the vein system. Fault gouge seams are commonplace at the footwall and hanging wall of the Trinidad vein system. The Trinidad hanging wall splays and the Trinidad footwall veins are considered to be part of the Trinidad mineralized structure.

The Bonanza vein system (Bv) is emplaced in the hanging wall zone of the structural corridor hosting the mineralized vein systems in the Trinidad deposit. The Bonanza vein system generally strikes 350° and dips steeply to the east to sub-vertical. The Paloma vein (Pv) is considered to be part of the Bonanza vein system. Mineralization within the Bonanza vein system is present in the form of shoots plunging shallowly to moderately to the north-northwest, reflecting the dominant dip-slip movement of the controlling fault structures. Combined copper, lead and zinc values for the Bonanza Vein range from negligible in the upper portions of the vein system to approximately 0.1 to 0.5 percent at depth.

The Trinidad North discovery refers to an area located between 1846200N and 1847800N and below the 1200 meter elevation. Brownfields exploration in the Trinidad North area has been successful in identifying high-grade silver and gold mineralization associated with the northern extensions of the Bonanza and Trinidad vein systems and associated stockwork zones. Similar to the main Trinidad deposit, the mineralization is directly associated with the presence of hydrothermal breccias, crackle breccias and sheeted and stockwork-like quartz veinlets.

The Fortuna vein (Fv) strikes north-south and in contrast to the other major veins in the deposit, dips steeply to the west. The Fortuna vein has been extensively mined on levels 2, 3 and 4 of the historic mine workings with vein widths ranging from 2 meters to approximately 5 meters.

The main Stockwork Zone is located between 1846550N to 1847200N and 1,000 masl to 1,300 masl, being located in an extensional environment between the principal Bonanza and Trinidad structures. The main Stockwork Zone is present over 650 horizontal meters and 300 vertical meters being elliptical in shape, with a variable thickness ranging to greater than 50 meters. Stockwork-style mineralization is also present in the Trinidad North area between the Trinidad and Bonanza structures.

The primary silver bearing mineral in the Stockwork Zone is acanthite, usually in association with traces of pyrite. Secondary minerals accompanying the acanthite are silver-rich electrum, fine grained galena, sphalerite, chalcopyrite and gangue minerals including hyaline quartz, white quartz, and calcite along with minor concentrations of adularia and fluorite.

The mining method applied at the San Jose Mine is overhand cut-and-fill using a mechanized extraction methodology. Production capacity has been increased to 3,000 tpd as a result of the mill expansion commissioned in June 2016.

The method chosen for underground mining is overhand cut-and-fill which removes ore in horizontal slices, starting from the bottom undercut and advancing upwards. When ore widths are greater than 8 m, a combination of overhand cut-and-fill and room-and-pillars has been selected as the best method for the conditions encountered.

Mechanized mining utilizes a Jumbo drill rig to drill blast holes, scoop trams for loading and trucks for ore haulage. Rock support is provided through rock bolts and shotcrete. The ore deposit width ranges from 4.5 m to 17 m for the Bonanza and Trinidad vein systems and can be more than 30 m in the Stockwork Zone.

Expansion of the concentrate plant was successfully completed in June of 2016 taking the ore throughput capacity from 2,000 dry tpd to 3,000 dry tpd. The principal stages are follows:
1. Crushing
2. Milling
3. Flotation
4. Thickening, filtering and shipping.

Crushing at the San Jose mine is a dry process, where ore extracted from the mine is reduced in size from 406 mm to 12.7 mm to be fed to the mill.

The crushing process begins at the reception hopper, where ore from the mine is deposited. The ore is fed from the bottom of the hopper via a plate feeder into a jaw crusher that crushes the ore to a 102 mm product size prior to it being transported via conveyors to one 2.44 m by 6.1 m primary screen deck. The screen deck operates with one mesh of 35 mm opening. Material that does not pass through the 35 mm mesh is sent to a secondary crusher via a chute, where it is reduced to 25 mm and the product returned to the primary screen deck. The material that passes the 35 mm mesh is sent via a conveyor to one 3 m by 7.3 m secondary screen deck. The screen deck operates with one mesh of 12.7 mm opening. Material that does not pass through the 12.7 mm mesh is sent to a tertiary crusher where it is reduced to 12 mm size before being sent back to the jig to close the circuit. The fine ore that passes through 12.7 mm mesh is sent to fine ore storage, achieving a final product of 12.7 mm that is stockpiled before being fed into the milling circuit.

The fine ore stock is sent via conveyor belts to either a 3.96 m by 5.94 m or 4.57 m by 6.6 m ball mill with 25 to 30 percent of their volume filled with three inch wrought steel balls used to further reduce (grind) the ore size. The product of the mills is pumped to the classification process comprised of hydro-cyclones, where two products are generated; 1) a fine ore, which is expulsed thorough the top of the cyclones, and 2) a coarse ore that exits through the bottom and is recycled back into the mills for further grinding. The fine ore must comply with the metallurgical conditions for metal recovery, which indicates 80 percent of the product must be under the 200 mesh size (equivalent to 74 microns), before being sent to the flotation process.

The pulp (water + mineral) received from the fine ore of the hydro-cyclone, is first sent to a flotation stage performed in ten mechanic cells, six 14.2 cubic meter and four 17 cubic meters in size, which generate agitation through a propeller and diffuser that distributes the pulp and injects air. The agitated pulp allows reagents to act on the elements of value and adhere to the bubbles formed by the injected air, freely spilling over the edges of the cells into a collection trough. The resulting product is known as the primary concentrate. Upon conclusion of this first stage, the pulp is sent by gravity to a second stage.

The second flotation stage is similar to the first, utilizing an additional six 17 cubic meter mechanic cells to generate the same conditions, from this process a secondary scavenger concentrate is obtained. This secondary scavenger concentrate is returned to the beginning of the process. Mineral that did not float in the second flotation stage is regarded as tailings and passes to the thickening process.

The primary concentrate is sent to a first cleaning stage that is carried out in twelve 2.8 cubic meter mechanical cells, whose function is to eliminate impurities and increase the grade of the concentrate. The product obtained is a first clean concentrate and the residue is returned to the beginning of the process. The first clean concentrate is sent to a second cleaning stage performed in three 2.8 cubic meter mechanic cells having a similar function to the first where impurities continue to be removed to obtain a second clean concentrate and the residue is returned to the first cleaning stage. The second clean concentrate is sent to a third cleaning stage performed in two 2.8 cubic meter mechanic cell to obtain a final concentrate that passes to the thickening stage and a residue that returns to the second cleaning stage.

The third cleaning concentrate is sent to a thickening tank where, using a flocculating reagent the particles are agglomerated and sediment generated. Solids and liquids are separated so as to recover water to put back into the process (recovered water) while the thickened solid is pumped to a two press-type pressure filter of twelve tarpaulin covered plates, where part of the water is eliminated and then re-circulated to the process. The concentrate cake is discharged from the filters to the concentrate storage for transportation.

The underflow of the final bank of the second flotation (exhaustion) is sent to a thickening tank where a solid-liquid separation is performed through the application of a flocculating reagent that agglomerates fine particles into sediment. Recovered water is returned to the process while the rest of the pulp is pumped to a three press-type pressure filter of one hundred and forty-five tarpaulin covered plates, where most of the water is eliminated and then re-circulated back into the process. The tailings cake is discharged from the filters to the tailings stock for transportation to the dry stack disposal area.

Part of the pulp pumped to the pressure filter is deviated to the paste fill plant for backfilling purposes with 30 percent of the mines backfilling requirements being supplied by the paste fill plant.