The Yaxtché deposit shows alteration assemblages typical of high sulfidation epithermal deposits whereas the metal content and sulfide assemblages are characteristic of mineralizing fluids with an intermediate sulfidation state.

The transition from high- to intermediate-sulfidation state is thought to define an evolving epithermal system as high-sulfidation state metal-bearing fluids cooled and interacted with host rocks as they moved vertically and laterally though the Yaxtché structure. This is depicted with three stages of primary fluid evolution:

• Alteration and gangue mineral assemblages related to acidic magmatic–hydrothermal fluids created permeability through acid leaching (i.e. vuggy silica)
• High-sulfidation state mineral assemblages (namely enargite–luzonite–famatinite) and metal contents (copper–gold dominant) formed at lower elevations within the Yaxtché structure
• Transition of high- to intermediate-sulfidation state as metal-bearing fluids ascended and further interacted with host rocks. The final phase of fluid evolution was critical for precipitation of silver-bearing minerals as tennantite–tetrahedrite became stable.

Sillitoe and Hedenquist (2003) defined the following key features of intermediatesulfidation systems:

• Intermediate-sulfidation deposits occur in calc-alkaline andesitic–dacitic arcs, although more felsic rocks can locally act as mineralization hosts
• The arcs typically display neutral to mildly extensional stress states
• Deposits form under acidic, oxidizing conditions within 1 km of the surface and
between temperatures of 150º and 250ºC
• Deposits show a large range in metal content and characteristics and can vary along the spectrum from gold-dominant to silver-dominant mineralization
• Although there is a large range of sulfide and sulfosalt minerals, these are dominated
by sphalerite with low FeS content, and include galena, tetrahedrite–tennantite, and chalcopyrite. Sulfide abundance can vary from 5–20 vol%
• Mineral assemblages typically contain Ag ± Pb, Zn (Au)
• The typical Ag:Au ratio is > 20:1
• Minor mineral associations can include Mo, As, Sb; may have associated tellurides
• Silica alteration can include vein-filling crustiform- and comb-textured veins
• Typical alteration assemblages include advanced argillic, alunite and kaolinite with
pyrophyllite deeper in the system; the proximal alteration mineral is often sericite.

Mineralization at Yaxtché consists of fine-grained black sulfides and sulfosalts that are difficult to identify in hand specimens. The mineralization occurs variously as disseminations, open-space filling, and in massive veinlets or clots. The identified mineralogy is consistent with that expected within a high- to intermediate- sulfidation epithermal deposit.

Coote (2010) observed:

• Tennantite–tetrahedrite is both intergrown with and overgrowing/replacing enargite–luzonite defining a trend of progressively decreasing sulfidation state of acid hydrothermal fluids with time at any given location within the hydrothermal system. The association of minor amounts of very fine-grained chalcopyrite with tennantite–tetrahedrite as overgrowths to or replacement of enargite–luzonite is consistent with the interpreted decreasing hydrothermal fluid sulfidation state. Sphalerite, locally abundant in association with the tennantite–tetrahedrite, formed about or after luzonite–enargite, also formed as a component of the physio-chemically evolving acid hydrothermal system

• Silver is mostly identified (from electron microprobe analyses and reflected light optical properties) as a component of the complex antimony- and lead-bearing and bismuth-rich sulfosalts which span the enargite–luzonite through to predominant tennantite–tetrahedrite paragenesis. It would appear that silver is poor in early bismuth-rich sulfosalts and rich in the later bismuth-rich sulfosalts that are mostly associated with tennantite/tetrahedrite. Silver mineralization therefore is also genetically associated with the evolving high-sulfidation system. Only minor to trace amounts of argentite are associated with tennantite–tetrahedrite and sphalerite.

Distinctive metal zonation patterns are recognized at Yaxtché. Patterns are broadly defined as a copper–gold assemblage at lower elevations, transitioning upwards into a silver–lead–zinc– barium–antimony metal assemblage at higher elevations. These zonation patterns suggest that physio-chemical gradients had a significant control on localization of silver bearing mineral assemblages. Corbett (2012) proposed that sites of bonanza grade silver mineralization may be a product of fluid mixing along structures as silver-bearing fluids mixed with low pH steam heated waters collapsing down faults.

Preliminary work was performed on four possible mining systems: post-pillar cut-andfill, transverse with pillars, transverse with cemented fill, and sublevel with end slicing.

The post-pillar cut-and-fill mining method was selected for the PEA evaluation.

This mining method relies on using 5 m x 5 m rooms and 5 m x 5 m square pillars. The pillars of one level are planned to align vertically with the next mining level to provide support. Mining starts at the lowest elevation in a mining area and is completed working upward. Some pillars can be extracted when they occur in an area of the stope where there will be no mining above.

Productivity calculations and unit costs for the final selection of a post-pillar cut-and-fill mine plan are based on the following criteria:

• Two 10 hr shifts per day, which yield 16 hr of actual work time
• The ground is considered medium–hard for the drilling and blasting variable
• Drilling of 18 ft (5.5 m) holes (providing 16.6 ft (5.1 m) of advance) with a two-boom production jumbo. Hydraulic drifter (drill) penetration rate is assumed to be 3.5 ft/min (1.1 m/min)
• No standard ground support is indicated; however, 1.5 hr per cycle is allotted for ground control (e.g. rock bolting), spot bolting or other cycle interruptions
• Blasting will be carried out using gelatin-class dynamite and ammonium nitrate/fuel oil (ANFO) as the primary blasting agent. Non-electric (non-el) type detonating systems would be used
• Stope areas will be ventilated with 75,000 ft3/min (2,123.8 m3/min) fans and 42 in (106.7 cm) flexible brattice ducts
• Mucking will be conducted using 7 yd3 (5.4 m3) load-haul–dump (LHD) units. Fill placement will be done by the same mucking units in conjunction with an LHD setup with a rammer for ensuring the fill is tightly compacted before excavating the next cut (above)
• Jumbo availability is assumed to be 85% with a 4 hr repair interruption resulting in a 7.4 hr cycle that includes 2.5 hr for travel and other non-productive time
• LHD availability is estimated at 85% with a 2 hr maintenance and repair period resulting in a 3.3 hr mucking cycle per blasted heading that includes ¾ hr for travel and utilization
• The blasting cycle is seven hours, which includes one hour for travel and lack of utilization.

A typical stope (section) area is predicted to have a production of 1,400 t/d (1,200 t/d of mill feed material and 200 t/d of waste). Backfilling and access development are not on the critical path. The typical stope level will have 98 m of linear development, which results in an average of 200 t/d of internal stope waste development. Analysis of the level plans show that an average of 55% of the room excavations use a horizontal breasting technique.

The drilling task of the excavation cycle will have the longest cycle time of the three components (drill, blast, muck) of a complete production cycle. The drilling cycle dictates the stope output, assuming a single drill unit per stope area (based on the typical stope areas). Drill jumbos selected for the stopes are capable of completing two room cycles per day, and 2.4 breast cycles per day (for a combined 700 t/d of broken material). A single blasting unit will produce 800 t/d. A single 7 yd3 (5.4 m3) LHD unit can produce 1,359 t/d.

Two active stopes are required to satisfy the production requirement of 1,200 t/d. Advance rates of 2.7 m/d of main development (e.g. ramp, accesses) are required to sustain the daily production rate.

The processing facility flowsheet was developed to recover silver from the Yaxtché sulfide deposit. The current design basis is set to process 1,200 t/d of mineralized material from the underground mine for the production of a bulk silver concentrate by conventional crushing (two stages), grinding (single stage) and flotation (rougher [two stages] and cleaners [five stages]) techniques. Testwork completed by DML and JKTech was used as the basis for the design of the process plant. The results of DML’s 2012 locked cycle flotation testwork was Samuel Engineering’s primary data source. This testwork included only two cleaner stages for producing the bulk silver concentrate. Samuel Engineering modeled the mass balance for the process plant to include five cleaner stages in order to produce a marketable, high-grade bulk silver concentrate of about 11.5 kg/t Ag.

Run-of-mine (ROM) mineralized material would be fed to the comminution circuit. Comminution would be accomplished by two stage crushing followed by ball milling to produce a particle diameter of 80% passing (P80) of 45 µm. The ROM material would be dumped by mine trucks into a primary bin equipped with a grizzly feeder (1.2 m by 3.0 m). Oversize material from the grizzly feeder would be discharged to the primary crusher, a 115 kW jaw crusher (0.86 m by 1.12 m) with a rated capacity of 225 t/h. The primary crushed material would be combined with the undersize material from the grizzly and conveyed to the coarse crushed stockpile. The coarse crushed stockpile would have a live capacity of 800 t (or about 16 hrs surge). Coarse material from the stockpile would be reclaimed and conveyed to the secondary crushing circuit.

The coarse material would be pre-screened by the double-deck secondary screen. The screen oversize would be fed to a secondary cone crusher (2.1 m diameter) at an estimated throughput rate of 28 t/h. The secondary crushing circuit would produce a fine product at a rate of about 70 t/h at a nominal P80 of 13 mm. The fine crushed product would be conveyed to a fine crushed stockpile as feed to the ball mill grinding circuit. The fine crushed stockpile would have a live capacity of about 750 t which would provide a surge capacity feeding the grinding circuit of about 13 hr.

The ball mill circuit would operate in closed circuit with cyclones to achieve a grind size at a nominal P80 of 45 µm. The ground slurry from the ball mill (12.5 ft diameter by 15 ft long; 1,500 hp) would discharge to the ball mill discharge sump with about 98 t/h reclaim water to bring the pulp density to about 55% solids. The ball mill cyclone cluster would size the pulp to a P80 of 45 µm for flotation with an overflow of 54 t/h solids in a slurry at a pulp density of 35% by weight solids. Underflow from the cyclones at 170 t/h dry solids (300% circulating load) with a pulp density of about 64% solids by weight would be returned to the ball mill for further grinding.

Cyclone overflow would be pumped to the flotation circuit via the flotation feed conditioning tank. Flotation reagents (collectors, promoters and frothers) would be added to the slurry for conditioning along with recycled process water from the concentrate thickener overflow and tailings reclaim water. The rougher flotation circuit would be done in two stages. Rougher flotation stages one and two would be comprised of four 16 m3 mechanical cells in each stage separated by a second conditioning tank where more reagents are added. The flotation concentrates from both rougher stages would be combined and pumped to cleaner flotation. Cleaner flotation would be done in five stages using 3 m3 and 50 ft3 cells operating in closed-circuit to produce a highgrade silver concentrate. Testwork indicates the silver concentrate would contain elevated levels of arsenic, bismuth and antimony. Reagents would be added to each cleaner stage. The final concentrate from the fifth cleaner stage represents the final bulk silver concentrate. The tailings from the first cleaner stage would be sent to cleaner scavenger flotation with the scavenger concentrate returned to the ball mill and the scavenger tailings to the tailings thickener.

The bulk silver concentrate from the fifth cleaner stage would be pumped to the concentrate thickener where it would be thickened to about 30% solids by weight. The thickener overflow would be returned to the process water tank and the thickener underflow would be pumped to a holding tank ahead of the concentrate pressure filter. The concentrate filter would reduce the concentrate cake to about 10% moisture and the filtrate would be pumped back to the concentrate thickener. The final silver concentrate would be packaged in one tonne super sacks for shipment.