Pebble is a porphyry-style copper-gold-molybdenum-silver-rhenium deposit that comprises the Pebble East and Pebble West zones of approximately equal size, with slightly lower-grade mineralization in the center of the deposit where the two zones merge. The Pebble deposit is located at the intersection of crustal-scale structures that are oriented both parallel and obliquely to a magmatic arc which was active in the mid-Cretaceous and which developed in response to the northward subduction of the Pacific Plate beneath the Wrangellia Superterrane.

The oldest rock within the Pebble district is the Jurassic-Cretaceous age Kahiltna flysch, composed of turbiditic clastic sedimentary rocks, interbedded basalt flows and associated gabbro intrusions.

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.

Mineralization in the Pebble West zone extends from surface to approximately 3,000 ft deep 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. The Pebble East zone mineralization is hosted by granodiorite plutons and dykes, and by adjacent granodiorite sills and flysch. The Pebble East and West zone granodiorite plutons merge at depth.

Mineralization at Pebble is predominantly hypogene, although the Pebble West zone contains a thin zone of variably developed supergene mineralization overlain by a thin leached cap. Disseminated and vein-hosted copper-goldmolybdenum-silver-rhenium 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 Pebble West zone and deeper portions of the Pebble East zone. Rhenium occurs in molybdenite and high rhenium concentrations are present in molybdenite concentrates. Elevated palladium concentrations occur in many parts of the deposit but are highest in rocks affected by advanced argillic alteration. High-grade copper-gold mineralization also 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.

Within the deposit, the Kahiltna flysch is a well-bedded siltstone with less than 10% coarser-grained wacke interbeds; basalt and gabbro are absent. Bedding within the flysch typically dips less than 25º to the east. The diorite sills are found only in the western half of the deposit, whereas some granodiorite sills extend across the entire deposit. Intrusions and breccias of the alkalic suite occupy the southwest quadrant of the deposit.

The deposit is centered on a group of Kaskanak suite intrusions. Olson (2015) describes the sequence and composition of the intrusions within the Pebble deposit as: 1) earliest, voluminous equigranular granodiorite equivalent to the Kaskanak batholith; 2) transitionally porphyritic granodiorite stocks; 3) early-mineral granodiorite porphyry; 4) inter-mineral quartz granite porphyry; and 5) minor late-mineral high-silica quartz granite porphyry. The north contact of the Pebble East zone pluton is close to vertical, and its upper contact dips shallowly to the west; it remains undelineated to the south and has been dropped into the East Graben by the ZG1 normal fault to the east. Contacts of stocks in the Pebble West zone dip steeply to moderately outward.

The Pebble East zone is entirely concealed by the east-thickening cover sequence. The contact between the flysch and the cover sequence ranges from sharp and undisturbed to structurally disrupted with slippage along the contact. The lower half of the sequence comprises a thick basal conglomerate with well-rounded cobbles and boulders of intrusive and volcanic rock types of unknown provenance, overlain by complex, interlayered, discontinuous lenses of pebble conglomerate, wacke, siltstone, and mudstone. The upper half of the sequence comprises volcanic and volcaniclastic rocks dominated by basalt or andesite and intruded by minor felsic to intermediate sills and/or dykes.

The East Graben is filled by basalt flows and lesser sedimentary rocks that have an uncertain relationship to the cover sequence. The graben fill ranges from approximately 4,265 ft thick north of the ZE fault to a thickness of up to at least 6,400 ft to the south. Basalts in the lower half of the graben are cut by two ~65 Ma monzonite porphyry intrusions, which makes them older than the rocks that cover the Pebble East zone.

Eocene rocks are rare within and proximal to the Pebble deposit. Where thus far encountered, they comprise narrow felsic dykes, a pink hornblende monzonite intrusion intersected at depth in the central part of the East Graben, and a rhyolite flow breccia at the top of the East Graben, south of the ZE fault.

Structure
Within the western part of the Pebble deposit, the Kahiltna flysch occurs as an open, M-shaped anticline with axes that plunge shallowly to the east-southeast (Rebagliati and Payne, 2006). The folding predates intrusive activity at Pebble and diorite sills are commonly thicker where they exploited the hinges of the folds. Folding did not affect the cover sequence.

A Brittle-ductile fault (BDF) zone was identified on the east side of the Pebble deposit where it manifests a zone of deformation defined by distributed cataclastic seams and healed breccias. It strikes north-northeast, extends at least 1.86 mi along strike, is up to 650 ft wide and is vertical to steeply west-dipping. The BDF is truncated on the east by the ZG1 fault and does not affect the cover sequence. Displacement was dextral-oblique/reverse (S. Goodman, pers. comm., 2008), and correlation of alteration domains across the fault limits post-hydrothermal lateral displacement to less than 1,310 ft.

The ZB, ZC and ZD faults occur in the Pebble West zone and exhibit normal offset of diorite and granodiorite sills of between 50 ft and 300 ft. Normal displacement on the ZJ and ZI faults is not well constrained. The ZA fault has about 100 ft of apparent reverse movement. A minimum of 820 ft of normal displacement occurred across the steeply west-dipping ZF fault, juxtaposing mineralized sodic-potassic alteration in the east against poorly mineralized, propylitic and quartz-sericite-pyrite alteration to the west. Scissors-style, south-side-down normal displacement on the ZE fault increases from around 100 ft on its western end to about 980 ft on the east side of the deposit. The ZG1 fault forms the western boundary of the East Graben and has a well-defined normal displacement of approximately 2,100 ft in the north and 2,900 ft in the south, based on offset of the contact between the deposit and the cover sequence. The ZG2 fault, which is parallel to the ZG1 fault, has between 880 ft and 1,800 ft of normal displacement. The ZH fault and possible parallel structures farther east mark the eastern margin of the East Graben but remain undelineated.

The mining operations are planned to use conventional open pit mining methods and equipment. The open pit mine envisioned for the Proposed Pebble Project would be a conventional drill, blast, truck, and shovel operation with an average mining rate of approximately 70 million tons per year and an overall strip ratio of 0.12 ton of waste per ton of mineralized material.

The open pit would be developed in stages, with each stage expanding the area and deepening the previous stage. The final dimensions of the open pit would be approximately 6,800 ft long and 5,600 ft wide, with depths to 1,950 ft.

The projected mining schedule was generated using five pushbacks and was based on a maximum processing capacity of 180,000 ton/d. Based on the selected ultimate pit, final pit design and the generated production schedule, the Project’s total LOM is 21 years, including 1 year of pre-stripping followed by 20 years of production. Over the 21-year LOM, the pit would produce 1,291 million tons of mineralized material and 153 million tons of overburden and waste rock. The LOM stripping ratio (defined as waste material mined, in tons, divided by mineralized material mined, in tons) is 0.12:1. Approximately 1.4 billion tons of material are planned to be mined during the production phase.

The Project production fleet would use the most efficient mining equipment available to minimize fuel consumption per ton of rock moved. Most mining equipment would be diesel-powered. This production fleet would 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 would 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 would be routinely maintained to ensure optimal performance and minimize the potential for spills and failures. Mobile equipment (haulage trucks and wheel loaders) would be serviced in the truck shop; track-bound equipment (shovels, excavators, drills, and dozers) would be serviced in the field under appropriate spill prevention protocols. Track-mounted electric shovels would be the primary equipment unit used to load blasted rock into haulage trucks. Each electric shovel is capable of mining at a sustained rate of approximately 30 million tons per year.

Diesel off-highway haulage trucks would 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 would vary between 6 and 12 in. Drill rigs may be either electrically powered, as is the case for the larger units, or diesel powered.

This equipment would 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, would be used to improve operational safety and efficiency.

Slope design recommendations are provided in SRK (2012) and summarized as follows:
- Maximum stack height (MSH) = 400 ft;
- At a minimum, a geotechnical berm of 65 ft should be used separate the various stacks;
- Inter-ramp angle (IRA) = variable, depending on kinematics and rock mass stability, ranging from 40º to 55°;
- Bench face angle (BFA) = variable, kinematically controlled, expected break-back angles in the range of 75º to 55°;
- Bench height (BH) = double-benching (100 ft) in all sectors, with the exception of the YGs-Weak rocks which should be single-benched (50 ft). Fault-zones are considered ‘weak’ and need to be single-benched at a rate of one below and three above, and should be further investigated and applied at the feasibility level; and
- Bench width (BW) = is scaled according to the rock mass condition, typically in the range of 30 to 50 ft) for the 25-year pit.

Primary Crushing
Mineralized material would be delivered by haulage trucks to each of the two 60 ft x 110 ft fixed primary gyratory crushers. The crushers would be set to produce a product P80 of 145 mm. Located underneath each primary crusher would be the crusher discharge vault and an apron feeder that would control the rate of discharge onto the sacrificial conveyor belt below.

The crushing plant is designed for an operating availability of 75%. Each crusher would have a typical operating range of 5,000-6,000 tons/h depending on the ROM material size distribution. Each crusher would discharge onto a common main overland conveyor via a respective transfer conveyor. Each primary crushing station would be equipped with a rock breaker, dust control equipment and sumps for surface run-off collection.

The major primary crushing equipment is as follows:
• two 60 ft x 110 ft primary gyratory crushers; each fitted with a 1,500 kW drive; and
• discharge vault apron feeders and sacrificial belt conveyors.

Stockpile
Primary crusher product would be conveyed by the overland conveyor to the stockpile located adjacent to the grinding and flotation building. The covered stockpile would have a live capacity of 90,000 tons or 12 hours of mill operating time.

Under normal operation mill feed material would be reclaimed by two lines of three apron feeders onto two reclaim conveyor belts to the two grinding lines.

Primary Grinding
Two identical trains of SAG mill, followed by a conventional ball mill and pebble crusher (collectively SABC circuit) would receive reclaimed mill feed material from the coarse ore stockpile (COS). (Note that the term ore in this context refers only to mineralized feed material but is labelled as ore to conform with industry convention for naming of the stockpile; no economic surety is implied.) Each train would have an average throughput of 90,000 short tons per day. The equipment for the two primary grinding lines would comprise:

• two 42 ft diameter x 27 ft effective grinding length (EGL) SAG mills each with 30 MW gearless drive;
• screens, conveyors and feeders;
• one pebble crusher surge bin; and
• three 933 kW pebble crushers.

The reclaimed material would be fed to each SAG mill feed chute at which point process water and lime would also be added. An automatic ball charging system would deliver SAG mill balls when required. Each SAG mill would discharge onto a pair of SAG mill discharge screens. For each SAG mill, the screen undersize would gravitate to the cyclone feed pumpbox, while the screen oversize pebbles would be conveyed to a common pebble crushing plant equipped with a trio of crushers. Crushed pebbles would be conveyed to a surge-bin from where they would be split to one screen for each SAG mill. Similar to the SAG discharge screens, the crushed pebble screen undersize would discharge into the cyclone feed pump-box, while the screen oversize would return to the pebble crushers with the SAG screen oversize.

Secondary Grinding
Each SAG mill would feed a pair of ball mills via dedicated cyclone packs. Each pair of mills would share a common cyclone feed pump-box, which would split the slurry to one cyclone feed pump for each cyclone pack. The ball milling circuits would be designed to operate with a 300% circulating load. The major process equipment in the secondary grinding circuit comprises:
• four 26 ft diameter x 40 ft long (EGL) ball mills, each driven by a 16 MW twin pinion drive; and
• pumps and hydrocyclone clusters for each ball mill.
The grinding circuit product would have a P80 of 135 µm.

Bulk Concentrate Re-grind
The bulk rougher concentrate would flow to the bulk regrind mill pump-box which would deliver slurry to the regrind hydrocyclone cluster. Three 5,000 kW high intensity grinding (HIG) mills would grind the bulk rougher concentrate to a P80 of 25 µm.

Molybdenum rougher product Re-grind
The molybdenum rougher product would be reground with one 130 kW high intensity grinding (HIG) mill producing a P80 of 25 µm product.

The processing plant is designed with a feed rate of 180,000t tons per day. The feed material would be processed to produce two principal products, a copper-gold flotation concentrate and a molybdenum flotation concentrate, as well as a tertiary gravity gold concentrate through the following unit processes:

• primary crushing;
• grinding with SAG and ball mills;
• bulk copper-gold-molybdenum flotation;
• gravity concentration in the regrind circuit of the bulk rougher concentrate, and
• molybdenum flotation to separate a copper-gold flotation concentrate and a molybdenum flotation concentrate.

Bulk Rougher Flotation
The flow from the conditioning tank would be split between the two parallel banks of bulk rougher flotation cells. Each bank would consist of eight 824 yd3 tank cells, totalling sixteen cells in all. The reagents that would be added include lime, fuel oil emulsion (molybdenum collector), sodium xanthate (SEX) and methyl isob ........