The Escobal deposit formed in an intermediate-sulfidation epithermal quartz vein system of probable Upper Miocene to Lower Pliocene age. These deposits are commonly included in the low-sulfidation epithermal class of deposits. Distinguishing characteristics of an “intermediate sulfidation” environment include mineral assemblages indicating a sulfidation state between those of high and low sulfidation types, relatively high total sulfide content of 5 to 10 percent, low-iron “blond” sphalerite, presence of silver sulfosalts, and association with andesitic to dacitic volcanics. Magmatic- associated fluids are implied.

Epithermal deposits form as high-temperature mineralizing fluids rise along structural pathways and deposit quartz and precious- and base-metal minerals in open spaces in response to boiling, which is usually coincident to a release of pressure within the hydrothermal system. This quartz and metal deposition, followed by resealing of the system, is repeated over the life of the hydrothermal system resulting in crosscutting and overprinted breccia and vein textures. Typically, the largest and highest grade deposits are associated with long hydrothermal systems marked by complex overlapping veins.

These deposits are strongly structurally controlled. Mineralizing fluids are directed along structural pathways with high- grade “ore shoots” typically concentrated in open dilatant zones. These dilatant zones commonly form where inflections occur vertically and laterally along the vein.

Metal deposition and zoning in epithermal deposits are related to the level of boiling. Typically, precious metals deposit at or near the boiling level while base metals precipitate below. Boiling may occur at different levels as the hydrothermal system evolves producing an overprint of various episodes.

The Escobal deposit occurs in a similar geologic setting with host rocks, vein characteristics and mineralogy typical of other intermediate-sulfidation systems. Specific definitive features include banded, cockscomb, and drusy vein textures; massive, stockwork and breccia veins; intermediate argillic and quartz-sericite alteration; appreciable base- metal and silver-sulfosalt mineralogy, and associated arsenic and antimony.

Economic mineralization at Escobal comprises silver, gold, lead, and zinc hosted within quartz veins, stockwork zones and hydrothermal breccias. The deposit predominantly comprises sulfide mineralization. Silver, lead, and zinc sulfide mineralization predominates in the Central and West zones though elevated gold values also occur at depth in the Central Zone. In the East Zone, gold-rich mineralization is associated with the upper mixed sulfide-oxide horizon. Silver mineralization in all zones shows a close association with galena and low-iron sphalerite.

A petrographic study of vein samples indicated a fairly simple and consistent paragenesis. Stage I veining consists of banded to massive chalcedony intercalated with quartz and carbonate. This is the volumetrically-dominant vein event and contains the bulk of sulfide minerals. Volumetrically lesser Stage II consists of sulfide-bearing granular chalcedony. Various episodes of post-sulfide quartz, and late barren calcite veining locally cut and/or overprint the main banded vein.

Active mining areas in the Escobal mine are accessed through two main portals, called the East and West portals. These two primary declines provide access to the Central Zone. A third primary ramp is being driven into the East Zone from the Central Zone. Access ramps are driven from the main ramp system to establish sublevel footwall laterals driven parallel to the vein in transverse mining areas on 25 m vertical intervals, with stopes accessed from the footwall laterals. In longitudinal mining area, development is done on vein also on 25 m vertical intervals.

Production from the Escobal mine is achieved via longhole stoping methods. Two variations of this mining method are utilized. Where the vein dimension across the strike is less than 15 m, longitudinal longhole stoping is applied. This method consists of driving horizontal drifts, spaced 25 m vertically, along the strike of the vein and then blasting the ore vertically from the upper level (“over-cut”) to the lower level (“under-cut”).

Breaking slots are established at the extreme ends of the stope to provide a void space for production blasting. These breaking slots are excavated utilizing Cubex drills equipped with V-30 blind bore reaming heads to bore a 30-inch diameter raise between the upper-cut and under cut for each stope. Once the breaking slots are complete, the stope faces retreat towards the accesses by drilling holes between the over-cut and under-cut, charging the holes with explosives, and blasting a ring or row of holes at the end of the stope.

This process continues until the maximum hydraulic radius or design limit of the opening along strike is reached, at which time longhole mining ceases and the void is filled with paste backfill.

In areas where the horizontal vein width exceeds 15 m, measured perpendicular to the strike of the vein, stopes are developed perpendicular to the strike of the vein. This is commonly known as transverse longhole stoping. In this case, 5 m wide by 5 m high footwall laterals are developed approximately 20 m to the south of and parallel to the vein. Access to the footwall laterals is from the primary declines. Five (5) m wide by 5 m high over-cut and under-cut drifts are developed from the south side of the vein to the north side of the vein spaced 25 m vertically. The south side of the vein is typically the footwall but due to local variations in vein geometry can sometimes become the hanging wall.

In the same manner as the longitudinal stopes, breaking slots are established at the extreme ends of the stope to provide a void space for production blasting. Once the breaking slots are complete, the stope faces retreat towards the accesses by drilling holes between the over- cut and under-cut, charging the holes with explosives, and blasting a ring or row of holes at the end of the stope.

The transverse mining method allows for multiple stopes to be in production along strike simultaneously on any given sublevel. Stopes along strike are split into primary stopes and secondary stopes. Each primary stope is separated along strike by a secondary stope.

Ore from the Escobal mine is treated using conventional differential flotation processes producing lead concentrates with high precious metal (silver+gold) grades and zinc concentrates with a lesser precious metal component.

The original design basis for the processing facility is 3,500 tonnes of ore per day (t/d) or 1.28 million tonnes per year; though the installed crushing, grinding, flotation and concentrate processing components were sized for the contemplated increased throughput rate of 4,500 t/d.

The ore is picked up by a front-end loader and dumped into the crusher feed hopper, where it passes through a stationary grizzly to a variable speed vibrating grizzly for initial separation, with oversized material going to the primary crusher and undersized material by-passing the primary crusher.

Material from the vibrating grizzly feeder drops into the primary crusher feed chute that feeds directly into the primary crusher. The ore is crushed to 80 percent minus 150 mm and then discharged onto the secondary screen feed conveyor via the jaw crusher discharge chute. The crusher feed opening is 762 mm by 1,397 mm. The design maximum operating capacity of the crusher is 430 dmt/h of oversize ore. The process design for the nominal feed rate to the primary crusher is 219 dmt/h.

Secondary/tertiary crushing reduces the ore to minus 9 mm as feed to the ball mill. Ore coming from the primary crusher first passes a triple deck secondary screen with oversized material going to the secondary 1.73 m diameter, HP400 cone crusher.

The combined crushed ore is then delivered through tertiary screen feed bins that distribute the ore onto the two tertiary screen feeder belts. The feeder belts then discharge the ore onto each of the two tertiary screens. The crushed ore is weighed on the tertiary screen feed conveyor using a belt scale. The weight of crushed ore passing along the conveyor is measured to inform the operator of the actual crushing rate and then totalized for metallurgical accounting.

The final step in the secondary and tertiary crushing system is the crushing of the oversize ore from the tertiary screens. Each tertiary screen discharges the oversize ore into its corresponding tertiary crusher.

There are six conveyors serving the secondary and tertiary crushing system—the crusher discharge conveyor, the tertiary screen feed conveyor, the fine ore transfer conveyor, the reversible conveyor, the fine ore bin feed conveyor, and the fine ore bin reversible conveyor. All of these conveyors are standard belt conveyors with the same basic components.

Crushed ore is fed to the grinding circuit from the two fine ore bins. Two variable speed belt feeders are located under each fine ore bin. In addition to the fine ore, the ball mill receives grinding balls (generally 3- to 4.5-inch diameter) which grind the ore as the mill rotates. Process water is added to form a slurry in the mill. Zinc cyanide and zinc sulfate are added to the mill feed chute to prepare the ground ore for flotation.

Slurry flows from the ball mill through the discharge trunnion into a trommel screen attached to the mill. The ball mill discharge slurry from the trommel undersize goes to the cyclone feed sump. The oversize is collected in a tote box and returned to the crushing plant for recycle. The trommel screen undersize is combined with process water in the cyclone feed sump. The cyclone feed pump delivers the slurry from the sump to the primary grinding cyclones.

The cyclones produce the final grinding circuit product. The cyclone overflow contains the fine solids fraction in the feed and it reports to lead flotation. The cyclone underflow contains the coarse solids fraction that returns for additional grinding in the ball mill. The target slurry densities are 33 percent solids for the cyclone overflow and 70 percent solids for the underflow.

Differential, or selective, flotation is used to recover precious and base metals from the Escobal ore. The slurry exiting the ball mill is first routed through a lead flotation circuit which depresses the zinc and recovers the majority of the silver, gold, and lead. The lead tailing is then processed through a zinc flotation circuit where the zinc is reactivated for recovery with minor amounts of silver, gold, and lead.

The lead flotation circuit consists of rougher flotation cells, cleaner flotation cells and cleaner scavenger cells. The lead flotation system produces a silver-rich lead concentrate which also contains the majority of the recovered gold.

The grinding cyclone overflow is delivered to the lead flotation circuit as slurry containing approximately 30 percent solids by weight. The slurry flows by gravity to the lead rougher flotation conditioning tank, where reagents are added and agitated, before flowing to the lead rougher flotation cells. Lead rougher flotation consists of one row of seven 40 m3 WEMCO self-aspirated cells with 75-kW agitator motors in series. The cells in each line are arranged so that they drop in elevation from one cell to the next. Lead rougher flotation takes place at a pH of approximately 8. Slurry enters the rougher flotation row through a feedbox and flows through the first cell, then passes into the next successive cell, with each cell producing concentrate. This process continues throughout the bank of cells until the slurry discharges by gravity into the tailings launder. The main process flow then passes to the zinc rougher conditioning tank where it becomes the primary feed for the zinc rougher flotation circuit. Overflow from the rougher flotation cells flows to the lead regrind sump.

Overflow slurry from the lead regrind cyclone is the feed to the lead cleaner and cleaner scavenger flotation circuits. The design feed is 22.3 t/h of solids at a slurry density of 24 percent. The reground lead rougher flotation concentrate is agitated with reagents (zinc sulfate and zinc cyanide) in a 5.6 m3 conditioning tank then gravity-fed to the first cleaner flotation cells.

The zinc flotation circuit is similar to the lead flotation circuit, in that it consists of rougher flotation cells, cleaner flotation cells and cleaner scavenger cells. The zinc flotation system produces a zinc-rich concentrate with minor silver, gold, and lead that remained in the lead flotation tailings.

The zinc flotation conditioning tank receives the lead flotation tailings as a slurry of approximately 30 percent solids which are mixed with zinc activator, collector, and promoter reagents and agitated. The zinc rougher conditioning has an effective volume of 39 m3 and a design flow rate of approximately 471 m3/h. The zinc rougher conditioning tank includes a rubber-lined, dual-impeller agitator driven by an 18.65 kW motor. The zinc rougher conditioning tank discharges by gravity into the first zinc rougher flotation cell feedbox where frother is added to the slurry.

The zinc rougher flotation circuit consists of one row of seven 40 m3 WEMCO self-aspirated cells with 75-kW agitator motors in series. Each flotation cell has two dart valves that control the flow of tailings slurry into the next downstream flotation cell. The slurry from the first rougher flotation cell discharges through two dart valves into the next cell and into each successive cell. The nominal rougher concentrate production is 115 t/h of slurry, containing 25 percent solids.