Question 1 of 6: Dilution & Recovery, Power & Services, Equipment Costs, Scheduling, Block Caving and Rock Haulage

Question 1: Dilution & Recovery, Power & Services, Equipment Costs, Scheduling, Block Caving and Rock Haulage (40 marks)

Question text not reproduced: the examination questions are © Engineers and Geoscientists BC. Open the official past paper (linked at the top of this page) to read the question, then follow the worked solution below.

1.1 — Dilution and recovery

Dilution is the proportion of barren or sub-economic (waste) rock unavoidably mixed into the ore stream during breaking, mucking and hauling — expressed as $\text{dilution} = \dfrac{\text{waste tonnes mined}}{\text{ore tonnes mined}}$ (or as a percentage of the diluted tonnage). Recovery is the fraction of the in-situ economic reserve that is actually extracted and delivered to the mill, $\text{recovery} = \dfrac{\text{ore tonnes recovered}}{\text{ore tonnes in the reserve}}\times100\%$; ore left in unmined pillars, in a stope's unrecoverable corners, or lost to slough/cave dilution never reaching the mill is recovery loss.

The two values are inversely related in practice, though for different physical reasons depending on the method. In an open-stope method (longhole, VCR), pursuing higher recovery — taking every last metre of ore, mining tight to the contact, delaying backfill — widens the exposed span and invites wall slough, which raises dilution. Conversely, leaving rib or sill pillars to control dilution and ground stability directly removes ore from the extraction, lowering recovery. In selective methods (cut-and-fill, shrinkage), dilution and recovery can move together instead: mining tight to a narrow, irregular contact (resuing, selective blasting) simultaneously minimises waste dilution AND minimises ore left behind in the walls, so a well-controlled selective method can achieve both low dilution and high recovery at once — at the cost of a slower, more labour-intensive cycle.

Mining methods that minimise dilution and maximise recovery together are the filled, selective methods: cut-and-fill (backfilling each lift as it is taken removes the span-driven slough mechanism entirely, and resuing lets the miner follow a narrow, irregular vein contact closely); and shrinkage stoping in a narrow, regular, competent orebody (the broken muck pile itself supports the walls, limiting slough, while the method's narrow-vein selectivity limits waste pickup). Bulk open-void methods (longhole, VCR, sub-level caving) trade this off for cost and rate — they accept higher dilution and somewhat lower recovery in exchange for a much lower unit cost and higher mining rate, which is economic only where the orebody is wide and regular enough that the extra dilution does not overwhelm the value of the ore.

1.2 — Underground power distribution, services and dewatering

Power distribution. High-voltage power (typically 13.8–25 kV Canadian utility supply) is brought down the shaft or a decline on armoured cable to an underground main substation, where it is stepped down (commonly to 4.16 kV or 600 V) and distributed along the main haulage levels via trailing/feeder cable strung on cable hangers or in a dedicated cable raceway. At each working area a portable or skid-mounted substation/transformer further steps the supply down to utilisation voltage (600 V for large mobile equipment, 120/240 V or 550 V for lighting, small tools and fixed pumps); mobile equipment (LHDs, jumbos) is fed either by trailing cable (electric-tram machines) or is diesel-powered where a permanent power run to a mobile work area is impractical. Cable is protected by ground-fault relaying and circuit breakers at every substation, and the whole system is designed to be readily extended/relocated as development advances (portable substations on skids, quick-disconnect cable couplers).

Other services. Alongside power, a producing mine distributes compressed air (for pneumatic tools, air-actuated valves and, in older/smaller operations, air-powered loaders); potable and process water (drilling water, dust suppression, cooling); a fibre-optic or leaky-feeder communications/tracking backbone; and the ventilation system itself (fans, doors, regulators, auxiliary ducting) — all run in parallel with the power distribution along the same main infrastructure corridors.

Dewatering. Groundwater inflow, blast/drill water and any water released from hydraulically placed fill must be collected and pumped to surface (or to a permitted underground discharge/treatment point) continuously, or the workings flood. Water is directed by gravity along ditches/sumps at each level to a series of collection (sump) stations, then lifted in stages by multi-stage centrifugal pumps — typically a small "gathering" pump at each level lifting to a larger stage pump at a central pumping station, and a final high-head multi-stage pump lifting from the lowest main sump to surface (or to an intermediate surge pond partway up the shaft on a very deep mine, to keep any single stage's head within a practical range). Pump selection accounts for total dynamic head (static lift plus friction losses), flow rate, and — critically — solids content: mine water often carries fine rock flour and, where hydraulic fill is used, tailings slimes decanted from the fill; handling this abrasive, occasionally high-solids water requires slurry-duty (hard-metal or rubber-lined) centrifugal pumps, upstream settling/clarifying sumps to drop out the coarser solids before the water reaches the pump suction, and more frequent wear-part maintenance than a clear-water duty pump would need. A properly designed decant/underdrain system on the fill itself (filter fabric, perforated decant pipes) is the first line of defence, reducing the slimes load the pumping system must actually handle.

1.3 — Capital and operating costs of mine production equipment

Capital costs are the one-time (or infrequent, major-rebuild) purchase costs of equipment: LHDs, haul trucks, production and development drill jumbos/longhole rigs, the shaft hoist and headframe sheave/drum assembly, ventilation fans, dewatering pumps, the primary crusher, and rail/conveyor infrastructure. Capital cost also includes major mid-life rebuilds (engine or drivetrain overhaul, hoist rope replacement, drum re-lining) that are large enough to be capitalised and depreciated rather than expensed as routine maintenance.

Operating costs are the recurring costs of running that equipment: fuel or electric power consumption; consumable wear parts (LHD bucket teeth and tyres, drill steel/bits, hoist rope replacement on its normal fatigue-life cycle, crusher liners); scheduled preventive maintenance labour and parts (oil/filter changes, greasing, inspection); operator labour; and unscheduled repair. A useful distinction for equipment selection and mine economics is that capital cost is largely fixed once the machine is bought, while operating cost accrues per hour or per tonne moved — so a cheaper-capital machine with a high operating cost per tonne (e.g., an older, less fuel-efficient truck) can easily cost more over its life than a higher-capital, more efficient replacement, which is exactly the trade-off examined by an equipment life-cycle or replacement-economics study.

1.4 — Development timelines and production scheduling

Development timelines. Before any stope can be blasted, a defined sequence of access and preparation headings must be complete: the shaft or decline itself; main haulage drifts and crosscuts connecting the shaft/portal to the ore zone; ventilation raises/drifts establishing a complete air circuit to the new area; a sill drift (or top/bottom sublevel drifts, method-dependent) developed along the orebody at the stope horizon(s); and, for a drilled stope, the ring-drill or slot-raise development needed before production drilling and the initial slot blast can begin. Depending on depth, distance from existing infrastructure and rock conditions, this development chain typically runs from several months (a shallow, nearby extension of an operating mine) to two or more years for a new mine's first production headings (shaft sinking, +) — and it is sequenced explicitly on the mine's development schedule well ahead of the first tonne of ore because each downstream heading depends on the one before it (no sill drift without the crosscut that reaches it; no production blast without the sill drift and slot).

Production scheduling. Beyond simply setting where and when each stope is blasted, an essential production schedule must also plan and sequence: the development-ahead requirement, i.e. maintaining a rolling inventory of fully developed, ready-to-blast stopes several months ahead of the mining front so that a single delayed heading never stops production; grade control and blending, sequencing stopes of different grade so the mill feed stays within its designed head-grade and metallurgical window rather than swinging wildly stope to stope; backfill cycle timing, since an adjacent stope commonly cannot be mined until its neighbour's fill has cured to design strength, making the fill plant's throughput and cure schedule a real constraint on the mining sequence, not an afterthought; ventilation and services capacity, confirming the air, power and dewatering systems can support the planned number of simultaneously active headings; equipment and labour allocation across development and production faces competing for the same drill/LHD/truck fleet and crews; and geotechnical sequencing constraints such as pillar-recovery timing and stress-shadow effects between adjacent stopes, which can force stopes to be mined in a specific order regardless of which is "ready" first.

1.5 — Block caving: achieving open-pit-like production costs

Block caving achieves unit costs approaching those of a large open pit because it is the only underground method that is genuinely continuous and gravity-driven at bulk scale: once an undercut is established and the column above is induced to cave under its own weight and stress redistribution, ore reports to drawpoints by gravity flow with essentially no drilling or blasting of the ore itself (only the initial undercut and any secondary breakage of oversize/hangups requires drill-and-blast). Very high sustained tonnage (tens of thousands of tonnes per day from a single block), low labour intensity per tonne (a handful of LHD operators servicing dozens of drawpoints, versus a drill-and-blast crew per active heading), and minimal ground support beyond the extraction level itself (the cave column needs none) combine to push the cost per tonne down toward open-pit levels, while avoiding the pit's stripping-ratio burden and depth/slope-angle limitations.

Advantages: lowest unit cost of any underground bulk method; very high sustained production rate once established; minimal drilling/blasting of ore; well suited to large, low-to-moderate-grade, massive orebodies too deep or too low-grade for an open pit. Disadvantages: very high upfront capital investment and long lead time (years of undercut/extraction-level development before first ore); requires a cavable orebody and overlying rock (competent, non-cavable rock hangs up and stops production); essentially no selectivity — internal waste/low-grade zones cave and dilute along with the ore, and grade cannot be controlled by mining sequence the way it can in a selective stope; and once caving is initiated it is very difficult to stop or reverse, so the method commits the operation to a specific, large-scale mine plan for its life. Examples: Newcrest's Cadia East and Rio Tinto/Turquoise Hill's Oyu Tolgoi (both large-scale block/panel caves developed specifically to reach open-pit-like unit costs at depths beyond an economic pit); Freeport's Grasberg Block Cave (successor to the historic DOZ/Ertsberg caving operations) in Indonesia; and De Beers' historic Kimberley/Premier caving operations, among the earliest large-scale applications of the method.

1.6 — Moving ore from the draw point to the shaft loading pocket

Scoop-trams (LHDs). A load-haul-dump loader muck the broken ore directly at the draw point or chute and either trams it a short distance to an ore pass leading down to the haulage level, or, on a short haul, trams it all the way to the shaft loading pocket itself; LHDs are the universal first link in the chain because their compact, articulated design works in the confined, irregular geometry immediately around an active draw point where no other equipment can operate.

Trucks. For longer hauls than an LHD can economically cover (typically beyond a few hundred metres), rubber-tyred underground haul trucks take over — loaded by the LHD at a transfer/loading point — and run the ore along ramps and main haulage drifts to the shaft loading pocket or to a crusher station feeding a conveyor or ore pass system; trucks are the dominant mid-distance link in most modern trackless mines.

Conveyors. Where haul distances and tonnage are large and sustained (a major caving or bulk operation), a fixed or semi-mobile belt conveyor system — fed from a crusher station that sizes the ore ahead of the belt — moves ore continuously from a central collection point to the shaft loading pocket or directly to a skip-loading bin, offering much lower unit power and labour cost per tonne than truck haulage at the volumes a large caving operation produces, at the cost of the fixed infrastructure and crusher investment needed to feed it.

Gravity. Ore passes — near-vertical raises connecting an upper haulage or draw level to the shaft loading pocket level below — move broken rock down through the mine using gravity alone once it has been dumped in at the top, avoiding an entire haulage step; they are used everywhere the mine geometry allows (most caving and many longhole operations feed an ore pass directly from the draw points or LHD dump points) and terminate in a controlled draw-gate/chute at the loading-pocket level that meters rock into the skip loading system, exactly as at the shaft-bottom loading pocket described above.