Los Verdes contains molybdenum-(copper-tungsten) mineralization in breccias related to felsic intrusive rocks. A number of features of the deposit appear to place it in the category of a molybdenum-bearing breccia pipe related to a larger porphyry copper-molybdenum system. The indicative features are postintrusive breccias with interstices partly filled by fine-grained tourmaline, molybdenite, copper and varying amounts of tungsten and base-metals minerals, and late quartz-pyrite. Although Los Verdes has a tabular form, rather than pipe-shaped, it meets the definition of this type of deposit.

The mineralization at Los Verdes is of magmatic-hydrothermal origin, consisting originally of sulfide minerals, principally chalcopyrite, pyrite, and molybdenite, and apparently related to the granodiorite. Weathering near the surface has converted some of the mineralization to an oxide facies (principally ferromolybdite plus iron oxides). Some of the metals liberated by oxidation of sulfides have percolated downward, and have re-precipitated as secondary sulfides, mainly chalcocite, in the vicinity of hypogene sulfide minerals.

The mineralized zone on the Los Verdes property is about 300 meters northeast-southwest, and 200 meters northwest-southeast. Its eastern edge is at the Moly Pit, where it is cut off by the Buena Vista Fault. The mineralized zone is centered beneath the ridge extending southwest from the Moly Pit.

The Los Verdes mineralization is centered on a lens-shaped body of veining and stockworking within the granodiorite. The lower levels of the drilled portion of the granodiorite may be barren of mineralization but are typically cut by veins and veinlets of quartz with pyrite. This level is referred to a “hypogene granodiorite”.

At higher levels, the veining and brecciation become abruptly greater, with introduction of copper sulfides and molybdenite. Brecciation and filling of voids by tourmaline and chlorite becomes more abundant upwardly.

A distinct supergene zone is present at Los Verdes, overlying the hypogene mineralization. It is readily defined by Cu values much greater than the oxide zone, often over 1% Cu. Mo and W may be elevated compared to the oxide zone, but less notably. Copper-oxide minerals (notably malachite and azurite) often occur at the top of the zone.

The surface of the mineralized zone at Los Verdes is generally underlain by a near-surface oxide zone, typically carrying about 0.1% Mo, mainly as ferrimolybdite, and up to 0.2% Cu, largely as traces of malachite and azurite. While the same copper minerals may occur in both the oxide and the secondarily-enriched zone, the transition of molybdenum minerals is quite abrupt: ferrimolybdite above, and molybdenite below. Jarosite, and ferritungstite also occur, in addition to limonite. These minerals occur as disseminations and in weakly-defined veinlets in granodiorite.

The mine at Los Verdes will be accessed using conventional truck-shovel open pit mining techniques. Open pit mining operations will be based on a 21 hour per day, 6 days per week schedule for a nominal production rate of approximately 3500 tonnes per day. Mineralised material will be drilled and blasted on 5 meter high benches using a staggered blast hole pattern. Wherever possible waste will be drilled using 10 meter high benches using a wider blast hole spacing. All blasting will be performed using bulk ANFO (ammonium nitrate/fuel oil) explosives.

Primary loading equipment will be hydraulic excavators paired with articulated haul trucks. A preliminary waste storage location has been located approximately 1.5 km southwest of the open pit area. Mineralised material will be hauled by the mine trucks approximately 1.5 km to the main processing plant location. It has been assumed that all mining operations will be performed and directly supervised by a capable Mexican mining contractor.

A portable crushing and screening plant processing approximately 3500 tonnes daily is considered in the design. Crushing operations are currently envisioned as a two-stage circuit -- primary jaw crusher and secondary cone crusher. Crushing will be performed on a nominal 6 days per week schedule. All crushing equipment and related stockpiles will be located close to the process plant site.

ROM material is trucked to the crushing plant and dumped to a feed hopper. The 2-stage crushing reduces the rock from a maximum feed size of 400 mm down to 80% passing 18 mm. The crushed material is conveyed to a stockpile which is designed to have a capacity of 10,000 tonnes adequate for 3 days of plant feed requirements.

An external contractor will supply and operate the crushing equipment as well as the transfer of crushed material to the stockpile. All crushing and transfer equipment will be portable in nature and independently powered via diesel drives.

Crushed material is reclaimed from the stockpile on a 7 day per week basis at a rate of approximately 3000 t/d by a front end loader and is dumped into a reclaim hopper and feeder which feeds to the 1st stage grinding mill, a rod mill. A belt scale is installed to weigh and record feed rate to the concentrator. The reclaimed material is discharged to a rod mill with an estimated size of 3.4 m (diameter) by 4.6 m (length) equipped with a 750 kW drive. Rock is ground from particle size (F80) of 18 mm down to P80 of approximately 1 mm. Discharge from the rod mill passes through a trommel screen where oversize materials fall back to the mill while undersizes flow by gravity to a pumpbox where it is combined with the ball mill discharge. The slurry is pumped by a centrifugal pump to a hydrocyclone cluster for classification. Cyclone underflow with 60%wt solids returns to the ball mill which is operated as a wet-overflow-closed-circuit with a 300% circulation load and operates at 60 %wt solids. An overflow discharge ball mill, estimated size of 3.7 m (diameter) by 5.5 m (length) with a 5,000 kW drive is provided. Final cyclone overflow with a particle size of 80% passing 150 microns flows by gravity to a thickener. Thickener underflow with 65%wt solids is pumped to the bulk flotation stage while thickener overflow is pumped to a mill water tank.

A bulk rougher-scavenger flotation line is considered for the bulk flotation stage which produces moly/copper rougher concentrate. The underflow from the grinding thickener is pumped to a conditioning tank where lime and flotation reagents are added at their pre-determined dosages to adjust pH and prepare for flotation. Mill process water is also added to adjust slurry density prior to flotation.

Flotation cells are installed in a stepped arrangement allowing slurry to advance by gravity. The bulk rougher flotation is operated at pH 9.5 by addition of lime slurry. The total retention time in bulk rougher flotation is designed at 3 minutes based on preliminary test work results. Scavenger retention time following the rougher is 12 minutes.

Rougher concentrate is collected via concentrate launders to the concentrate pumpbox from where it is pumped to the 1st stage cleaner in the rougher concentrate cleaning area.

Rougher-scavenger tails is collected in a pumpbox and pumped to bulk flotation tailings thickener for dewatering prior to being disposed to the tailings area.

Overall, the mass pull in the bulk flotation stage is estimated at 11% of mill feed with copper and moly recoveries of +90%. Approximate 14 t/h (dry) of bulk concentrate is generated.

The bulk concentrate needs to be further cleaned to improve the concentrate grade prior to the copper - moly separation stage. A total of four (4) cleaning stages are provided for the primary purpose of removing pyrite in the bulk concentrate. Regrinding of concentrate from the 1st cleaner is included to fully liberate sulphide minerals and thus improve overall metal recoveries.

Rougher concentrate is pumped to the 1st cleaning conditioning tank where it is mixed with flotation reagents added at pre-set dosages. The conditioned slurry flows by gravity to the 1st tank in the first cleaning stage which consists of flotation cells installed in a stepped arrangement allowing slurry to advance by gravity. Tailings from the 1st cleaning stage pass on to the 1st cleaning scavenging. Scavenging concentrate is returned by a pump to the 1st cleaning stage, while the scavenging tailings are pumped to the bulk flotation tailings thickener.

Concentrate from the 1st cleaning stage is pumped to the 1st cleaner concentrate thickener to remove excess water prior to being sent to the regrinding circuit. Overflow from the thickener is pumped to the mill process water tank.

Cyclone overflow flows by gravity to the 2nd cleaning stage. Concentrate is further cleaned from the 2nd cleaning to the 4th cleaning stage.

Moly rougher concentrate goes to the moly cleaning stages for further upgrading. Approximately 2.2 t/h (dry) of copper concentrate at grade of 32% Cu (min.) is produced, which is about 2% of the original feed to the mill.

A regrinding circuit followed by a 5-stage cleaning circuit is designed for cleaning the moly to achieve less than 0.3% copper in the final moly concentrate. Approximately 0.25 t/h (dry) of final moly concentrate at a grade of 55% Mo (min.) is produced.

Bulk moly concentrate is pumped to a pumpbox where it is combined with moly regrinding mill discharge slurry. The mixed slurry is then pumped to a hydrocyclone for size classification. Cyclone overflow goes to a conditioning tank where it is mixed with reagents prior to entering the flotation tanks. The regrinding mill is designed to reduce particle size from a P80 of 30 microns to a P80 of 20 microns. The mill size is estimated at 0.9m (diameter) by 2.1m (length) with a 20 kW drive.