The Pebble deposit is classified as a copper gold-molybdenum porphyry deposit.

Pebble has one of the largest metal endowments of any gold-bearing porphyry deposit currently known. Comparison of the current Pebble resource to other major gold-bearing porphyry deposits shows that it ranks at or near the top in terms of both contained copper. In fact, Pebble is both the largest known undeveloped copper resource and the largest known undeveloped gold resource in the world today.

The Pebble East and Pebble West zones are coeval hydrothermal centers within a single magmatic- hydrothermal system. The movement of mineralizing fluids was constrained by a broadly vertical fracture system acting in conjunction with a hornfels aquitard that induced extensive lateral fluid migration. The large size of the deposit, as well as variations in metal grade and ratios, may be the result of multiple stages of metal introduction and redistribution.

Mineralization in the Pebble West zone extends from surface to approximately 3,000 ft depth and is centered on four small granodiorite plutons. Mineralization is hosted by flysch, diorite and granodiorite sills, and alkalic intrusions and breccias. The Pebble East zone is of higher grade and extends to a depth of at least 5,810 ft; mineralization on the eastern side of the zone was later dropped 1,970 to 2,950 ft by normal faults which bound the northeast-trending East Graben. East zone mineralization is hosted by a granodiorite pluton and adjacent granodiorite sills and flysch. The East and West zone granodiorite plutons merge with depth.

Mineralization at Pebble is predominantly hypogene, although the Pebble West zone contains a thin zone of variably developed supergene mineralization overlain by a leached cap. Disseminated and vein-hosted copper-gold molybdenum-silver mineralization, dominated by chalcopyrite and locally accompanied by bornite, is associated with early potassic alteration in the shallow part of the Pebble East zone and with early sodic-potassic alteration in the West zone and deeper portions of the Pebble East zone. High-grade copper-gold mineralization is associated with younger advanced argillic alteration that overprinted potassic and sodic-potassic alteration and was controlled by a syn-hydrothermal, brittle ductile fault zone located near the eastern margin of the Pebble East zone. Late quartz veins introduced additional molybdenum into several parts of the deposit.

The proposed Pebble Mine will be a conventional drill, blast, truck, and shovel operation with a mining rate peaking at 90 million tons per year and an overall stripping ratio of 0.1 ton of waste per ton of mineralized material.

The open pit will be developed in stages, with each stage expanding the area and deepening the previous stage. The final dimensions of the open pit will be approximately 6,500 feet long and 5,500 feet wide, with depths between 1,330 and 1,750 feet.

Mining will occur in three phases: Preproduction, Production, and Stockpile Reclaim.

The mine operation will commence during the last year of the Preproduction Phase and extend for 14 years during the Production Phase. During this period, 1,100 million tons of mineralized rock and 100 million tons of waste rock will be mined. After the open pit is depleted, the process plant will be fed with mineralized material reclaimed from the LG stockpile. Non-potentially acid generating (NPAG) waste rock will be used in construction of the tailings embankments or mine site roads. The potentially acid generating (PAG) waste rock will be stored in the LG stockpile until closure, when it will be back-hauled into the open pit. Fine- and coarse-grained soils will be stored southwest of the pit and north of the main tailings storage facility (TSF) embankment and will be used for reclamation during mine closure.

The Preproduction Phase consists of dewatering the pit area and mining of non-economic materials overlying the mineralized material, to be stockpiled for initial process plant feed, from the initial stage of the open pit. Dewatering will begin approximately one year before the start of preproduction mining, which will last for one year. Approximately 30 million tons of material will be mined during this phase.

The Production Phase encompasses the period during which economic-grade mineralized material will be fed to the metallurgical process plant that produces concentrates for shipment and sale. The Production Phase is planned to last for 20 years. Mineralized material will be mined for 14 years and be fed through the process plant at a rate of 160,000 tons/day. The open pit will be mined in a sequence of increasingly larger and deeper stages. As the mining rate will exceed the processing rate, surplus mineralized material, selected based on its relative value, will be stored on the LG stockpile and later processed during the Stockpile Reclaim Phase. Approximately 1.2 billion tons of material are planned to be mined during the Production Phase.

The Stockpile Reclaim Phase will commence during the last year of the Production Phase and extend for an additional six years. During this phase, mining activity will be limited to reclaiming material from the LG stockpile to feed through the process plant. The process rate will continue at 160,000 tons/day. At the end of this phase, the LG stockpile will be depleted.

The proposed Project will use the most efficient mining equipment available in the production fleet to minimize fuel consumption per ton of rock moved. Most mining equipment will be diesel-powered. This production fleet will be supported by a fleet of smaller equipment for overburden removal and other specific tasks for which the larger units are not well-suited. Equipment requirements will increase over the life of the mine to reflect increased production volumes and longer cycle times for haul trucks as the pit is lowered. All fleet equipment will be routinely maintained to ensure optimal performance and minimize the potential for spills and failures. Mobile equipment (haul trucks and wheel loaders) will be serviced in the truck shop; track-bound equipment (shovels, excavators, drills, and dozers) will be serviced in the field under appropriate spill prevention protocols.

Electric shovels, each mounted on a base platform equipped with tracks, will be the primary equipment unit used to load blasted rock into haul trucks. Each electric shovel is capable of mining at a sustained rate of approximately 30 million tons per year. Diesel hydraulic shovels, due to their greater flexibility, will be used to augment excavation capacity, depending on the mining application.

Wheel loaders are highly mobile, can be rapidly deployed to specific mining conditions, and are highly flexible in their application. Diesel off-highway haul trucks will be used to transport the fragmented mineralized material to the crusher.

Track-mounted drill rigs are used to drill blast holes into the waste rock and mineralized material prior to blasting. Hole diameters will vary between 6 and 12 inches. Drill rigs may be either electrically powered, as is the case for the larger units, or diesel powered.

This equipment will be supported by a large fleet of ancillary equipment, including track and wheel dozers for surface preparation, graders for construction and road maintenance, water trucks for dust suppression, maintenance equipment, and light vehicles for personnel transport. Other equipment, such as lighting plants, will be used to improve operational safety and efficiency.

Blasted mineralized material from the open pit will be fed to a crushing plant to reduce the maximum particle size to approximately six inches. This crushed material will be conveyed to a coarse ore stockpile, which in turn feeds a grinding plant within the process plant. In the grinding plant, semi-autogenous grinding (SAG) mills and ball mills further reduce the plant feed to the consistency of very fine sand. The next step is froth flotation, in which the copper and molybdenum minerals are separated from the remaining material to produce concentrates. The concentrates are then filtered for shipment. Gravity concentrators will be placed at strategic locations to recover free gold, which will be shipped off site for refining.

The copper-gold concentrate will be loaded into covered bulk shipping containers; the molybdenum concentrate will be packaged in bulk bags and loaded into shipping containers. Other economically valuable minerals—palladium and rhenium—will be present in the concentrates and may be recovered at the refineries.

The concentrate containers will be transported by truck to the Amakdedori Port on Cook Inlet. The contents of the copper-gold concentrate containers will be directly unloaded into the holds of Handysize bulk carriers for shipment, while the molybdenum containers will be loaded onto barges or other ships.

Over the life of the Project, approximately 1.1 billion tons of mineralized material will be fed to the process plant at a rate of 160,000 tons/day. On average, the process plant will produce approximately 600,000 tons of copper gold concentrate per year, containing approximately 287 million pounds of copper, 321,000 ounces of gold and 1.6 million ounces of silver, and approximately 15,000 tons of molybdenum concentrate, containing about 13 million pounds of molybdenum.

CRUSHING AND GRINDING:
Mineralized material from the open pit or LG stockpile will be delivered by 400-ton haul trucks to primary gyratory crushers located adjacent to the rim of the open pit. The crushers will reduce the mineralized material to a maximum size of six inches. The crushed mineralized material from both crushers is delivered via a single, covered, overland conveyor to the coarse ore stockpile.

The coarse ore stockpile is contained within a covered steel frame building to minimize fugitive dust emissions and control mineralized material exposure to precipitation. The stockpile provides surge capacity between the crushers and the process plant, improving the efficiency of the latter and enabling it to operate if the feed from the crushers is not available.

The stockpiled material will be reclaimed by apron feeders mounted below the pile that deliver it onto two conveyor belts feeding the SAG mills. Baghouse-type dust collectors will be provided at each transfer point to control fugitive dust emissions. Water will be added to the process at the SAG mill, thereby eliminating the need for additional baghouses. A sump will be located in each reclaim tunnel to collect any excess water; however, such drainage is likely to be minimal.

The primary grinding circuit will use two parallel, 40-foot-diameter SAG mills and associated ball mills to grind mineralized material to the finer consistency necessary to separate the valuable minerals. Steel balls are added to the SAG mill to aid in grinding the mineralized material. Coarse mineralized material, water, and lime are fed into the SAG mills and the mineralized material is retained within the SAG mills by grates until the particles reach a maximum size of one to two inches.

Discharge from each SAG mill will be screened to remove larger particles ranging from one to two inches (“pebbles”) and sent to the ball mills. The large particles will be conveyed to the pebble-crushing facility where they will be crushed and re-introduced to the SAG mill.

The next grinding step is ball milling. Ball mills have a lower diameter-to-length ratio than SAG mills and use a higher percentage of smaller steel balls compared to SAG mills, allowing them to grind the feed to a finer size. Two ball mills will be matched with each SAG mill.

The slurry from the ball mills will be pumped into the hydro-cyclones, which separate the finer material from the larger material through centrifugal force. The slurry with the coarser material will be recycled back to the ball mills for additional grinding. The slurry containing the finer material will be pumped to the flotation cells. Grinding circuit slurry pH levels will be adjusted to 8.5 by adding lime slurry to minimize corrosion on the mill liners and promote efficient mixing prior to flotation.

CONCENTRATE PRODUCTION:
Copper-gold and molybdenum concentrates will be produced via flotation, which will separate the metal sulfides from pyrite and non-economic minerals. Two tailings streams will be produced: bulk tailings and pyritic tailings.

The rougher flotation circuit is designed to separate the sulfide minerals, predominantly copper, molybdenum, and iron sulfides (pyrite) within the process plant feed from the non sulfide minerals. Slurry from the ball mills is split between two banks of bulk rougher flotation cells. Reagents added to the slurry promote mineral separation by inducing mineral particles to attach to air bubbles created by blowing air through the flotation cells. Additional reagents are added to promote froth bubble stability. This froth, with the mineral particles attached, rises to the surface and is collected as a bulk rougher concentrate for the next phase of flotation.

Bulk rougher concentrate slurry is then routed to the regrind circuit. Material that does not float – the bulk flotation tailings from which most of the sulfide minerals have been removed – will be pumped to two tailings thickeners.

The bulk rougher concentrate is reground to sufficiently liberate minerals and enable the separation of the copper-molybdenum sulfide minerals from iron and other sulfides, thus producing concentrates with commercially acceptable grades. A gravity gold recovery circuit is attached to the regrind circuit to recover free gold that might otherwise be lost.

Reground bulk rougher concentrates will be upgraded through a two-stage cleaning process. The concentrate from the cleaning process will report to copper-molybdenum separation, while the tailings will report to the pyritic tailings thickener for thickening prior to pumping to the pyritic tailings storage cell in the TSF. The same reagents used in the rougher flotation circuit will be used in the cleaning circuit, with additional reagents used to aid in the suppression of gangue minerals. The cleaning stage is operated at an elevated pH through lime addition—to suppress pyritic minerals, which would lower the grade of final concentrates.

Water will be removed from the bulk concentrate in a conventional thickener. This will remove as much of the bulk flotation reagents as possible before the slurry enters the copper gold/molybdenum separation circuit, thus increasing separation process efficiency. Reagents will be recycled to the rougher process with the thickener overflow. The resulting slurry will contain 50 percent solids by weight and will go forward to copper gold/molybdenum separation.

The final flotation process is designed to separate copper-gold and molybdenum concentrates by adding reagents. The concentrate from the separation stage is the molybdenum concentrate, while the tailings comprise the final copper-gold concentrate.

The upgraded copper-gold concentrate will be thickened to 55 percent solids by weight in a high-rate thickener. The thickener overflow will return to various circuits for use as process water. The thickener underflow will be fed to a pressure filter to reduce the moisture to approximately eight percent.

The molybdenum concentrate will be thickened in a high-rate thickener to 55 percent solids by weight.