Question 1 of 6: Ground Support, Mining-Method Selection, Ventilation Measurement, Hoisting, Cost Indices and Haulage

Question 1: Ground Support, Mining-Method Selection, Ventilation Measurement, Hoisting, Cost Indices and 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 — Mine ground-support systems

1.1.1 Common anchor (point-anchored, mechanical-sleeve) threaded roof bolt. A steel bar with a threaded collar end and a mechanical expansion sleeve at the toe is inserted into a drilled hole and torqued; torquing expands the sleeve to grip the borehole wall at a single point near the hole bottom, then tensions the bar so a bearing plate clamps the immediate roof skin against the anchor, either suspending a loose bed from a stronger one above (suspension) or clamping several thin beds together into one stiffer composite beam (beam-building). It installs quickly with no cure time, is inexpensive, and is retrievable/re-tensionable, making it the default primary support in moderately jointed, self-supporting ground such as room-and-pillar backs. Because the anchor is a single point rather than the full column, it offers little resistance to shear across a joint plane and can lose tension as the rock relaxes or the anchor corrodes, so it is not relied on for large spans or squeezing/highly fractured ground.

1.1.2 Split set (friction-anchored) bolt. A slotted, C-shaped steel tube slightly larger in outside diameter than the drilled hole is driven in with a jumbo or hand-held leg; the tube compresses to fit the hole and its spring-back generates friction resistance continuously along the FULL embedded length, rather than at one point — it needs no resin cure and no torquing, so it can be installed within seconds of drilling. This makes it well suited to rapid, temporary or advance support immediately behind the face (scaling support, screen-hold-up bolts) and in moderately competent ground, but its friction-only anchorage gives lower ultimate capacity and stiffness than a fully grouted system, and it is not the bolt of choice for permanent, high-load or long-term corrosive-service support.

1.1.3 Cable bolts. Multi-wire steel strand (commonly 7-wire, sometimes bulbed or birdcaged for extra grout keying) is inserted into a long borehole — several metres to tens of metres — and grouted along its full length with cementitious grout, giving very high tensile capacity combined with the long embedment length needed to reach competent ground well above a large open span. Cable bolts are the standard permanent/primary support for large stope crowns, sill and crown pillars, and open-stope backs (VCR, sub-level open stoping) where a conventional 2–3 m roof bolt is far too short to reach beyond the de-stressed/fractured zone around the excavation.

1.2 — Room-and-pillar and VCR: geology, geometry, rock strength

1.2.1 Room and pillar. Geology — the method suits flat-lying to gently dipping, tabular, stratiform deposits of fairly uniform thickness (coal, potash, trona, many sedimentary-hosted base-metal deposits) where a regular grid pattern can follow the orebody without excessive dilution. Geometry — rooms and rectangular or square pillars are laid out on a regular grid; pillar centres and room widths are set from the required extraction ratio, and pillars are sized by tributary-area loading theory (pillar stress = overburden stress × tributary area / pillar area) checked against pillar strength (e.g. Obert–Duvall/Bieniawski formulas). Rock strength — the roof must be competent enough to bolt and span the room width without a beam-building failure, and the pillar rock must carry the tributary load with an adequate factor of safety against spalling/crushing; weak pillar rock forces larger pillars (lower recovery) or partial/sequential pillar recovery, while weak roof rock limits achievable room span regardless of pillar design.

1.2.2 Vertical Crater Retreat (VCR). Geology — suited to massive, competent, steeply dipping to sub-vertical orebodies with fairly regular geometry and strong, well-defined wall rock, since VCR is a bulk open-stoping method. Geometry — large-diameter (150–250 mm) vertical blastholes are drilled from a top-sublevel drill drift down to near the bottom-sublevel undercut, and the stope is blasted upward in successive horizontal crater slices, retreating toward the top sublevel while ore is drawn from drawpoints on the bottom sublevel. Rock strength — because the blasted void is left open (unsupported, or filled well after blasting) for an extended period spanning the full stope height and width, both the ore and the hangingwall/footwall must be strong and self-supporting; poor-quality walls risk progressive slough, dilution and, in the worst case, uncontrolled stope collapse.

1.3 — Air velocity measurement and humidity in hot, deep mines

1.3.1 Slow-moving air (e.g. a depleted, closed-off area). Below roughly 0.25–0.5 m/s a rotating-vane anemometer stalls and under-reads, so slow air is measured with a visible tracer — a smoke tube or smoke candle released into the airstream and timed over a measured distance — or with a hot-wire (thermal) anemometer, which senses convective heat loss from a heated element and remains accurate down to near-zero velocity.

1.3.2 Moderate velocity (e.g. supplying a working stope). A rotating-vane anemometer, traversed by hand across the airway or duct cross-section in a systematic grid (to average out the velocity profile across the opening), is the standard instrument in this range; the traverse average, multiplied by the measured cross-sectional area, gives the air quantity Q.

1.3.3 High velocity (a large fan, closed doors). A vane anemometer is easily damaged or gives unreliable readings at high velocity, and the closed doors around a running fan prevent a hand traverse across its face; instead a pitot-static tube inserted through a small access port measures the velocity (dynamic) pressure directly against a manometer or differential-pressure gauge, and velocity is recovered from $V = \sqrt{2\,p_v/\rho}$ — a method that works at any velocity and needs only a small, safe access opening rather than direct access to the airstream.

1.3.4.1 Measuring water content, and why it matters. The water content (humidity) of mine air is measured with a sling (whirling) psychrometer or an electronic hygrometer, which reads the wet-bulb depression — the difference between the dry-bulb (ordinary) temperature and the wet-bulb temperature of a continuously wetted thermometer bulb exposed to the moving air — from which relative humidity and the psychrometric wet-bulb temperature are obtained. Humidity matters because the body's principal cooling mechanism at depth is evaporative (sweat); high relative humidity suppresses evaporation, so at a given dry-bulb temperature a humid airstream removes far less metabolic heat than a dry one. Heat-stress indices used to manage productivity and heat-illness risk in hot, deep mines (e.g. wet-bulb globe temperature, WBGT) are therefore built around the WET-bulb, not the dry-bulb, temperature.

1.3.4.2 Infrastructure to improve productivity. Bulk mine air coolers (surface or underground refrigeration plants chilling the primary air supply) and localised spot coolers at active working faces reduce both dry-bulb temperature and, by condensing moisture out of the cooled air, humidity; chilled service water supplied to drills and to personal cooling garments further offloads metabolic heat; and refuge/rest stations with conditioned air allow scheduled recovery breaks during a hot shift, all of which are standard infrastructure investments once natural ventilation and depth alone can no longer keep the working wet-bulb temperature within a safe/productive range.

1.4 — Mine hoisting fundamentals

1.4.1 Two most important rope properties. Breaking (tensile) strength — the load-carrying capacity that sets the achievable factor of safety for a given duty — and bending-fatigue resistance/flexibility — how many cycles of bending over the sheave and drum the rope's outer wires can withstand before fatigue failure, which is governed by rope construction (e.g. locked coil vs. round strand) and the drum/rope diameter ratio of 1.4.3.

1.4.2 Most important factor in rope-size selection. The maximum static load the rope must carry — the fully loaded conveyance plus the full weight of the longest suspended length of rope — divided by the minimum static factor of safety mandated by the governing hoisting regulation (in BC, the Health, Safety and Reclamation Code for Mines); this is exactly the selection calculation performed for the 425 m shaft in Question 6.

1.4.3 Drum/rope diameter ratio. Every time a wire in the rope passes over the drum or a sheave it is cyclically bent, and the bending stress in the outer wires is inversely proportional to the drum diameter for a given rope diameter; too small a ratio therefore shortens the rope's fatigue life sharply. Typical minimum ratios run from roughly 60–80 for flexible round-strand rope up to 100–120+ for stiffer, higher-fatigue-life locked-coil rope — consistent with the ratio of 108 specified for the locked-coil rope in Question 6.

1.4.4 Overwind. Overwind is the condition where a conveyance (skip or cage) is hoisted past its intended stopping point at the head sheave/collar — a loss-of-control event that can drive the conveyance into the head sheave or headframe structure, or run the rope entirely off the drum, and is one of the most serious hoisting accidents a shaft can suffer.

1.4.5 Three overwind-prevention methods. (1) A primary depth/position indicator system that continuously tracks drum rotation against the conveyance's actual position and automatically cuts power and applies the service brake as the conveyance approaches its normal stopping point. (2) A mechanical overwind safety device (a crash beam, decelerating rope-catching device, or mechanically actuated dog/wedge arrangement in the headframe or on the drum shaft) that physically arrests conveyance travel if it passes beyond the normal stop, independent of the electrical control system. (3) An independent final (emergency) limit switch, set beyond the normal stopping point and wired directly to the emergency brake circuit, so that even a complete failure of the primary depth-indicator/controller still trips the brakes before the conveyance can reach the sheave.

1.5 — Mine cost-estimating terms

1.5.1 Marshall & Swift Mine/Mill (M&S M/M) cost index. A published capital-cost escalation index, developed by the Marshall & Swift valuation service (now Marshall Valuation Service), that tracks equipment and construction cost inflation specifically for the mining/milling sector, analogous in function to the Chemical Engineering Plant Cost Index used in process industries. Built from a base-year value of 100 and updated on a regular cycle by tracking a representative basket of mining equipment, labour and material costs, it lets an estimator escalate an old, known-year capital cost to the current year simply by multiplying by the ratio of current-year to base-year index — exactly the technique used with the Table 5.3 capital/operating indices in Question 5.3.

1.5.2 The "six tenths" (0.6/0.7 power) rule. A capacity-scaling relationship, $C_2 = C_1(X_2/X_1)^n$ with $n \approx 0.6\text{–}0.7$ for most mining and process equipment, used to scale a known cost at one capacity to an estimate at another capacity without a full re-estimate. The exponent reflects economies of scale: for a roughly cube-shaped structure or vessel, fabrication cost scales with surface area (a squared linear dimension) while capacity scales with volume (a cubed dimension), giving cost ∝ capacity^2/3 ≈ 0.67 — close to the empirically observed 0.6–0.7 range, and visible directly in the sub-unity exponents (0.591–0.946) of every category in Question 5's Table 5.2.

1.5.3 The O'Hara Method (CIM Bulletin, February 1980). A parametric, order-of-magnitude capital-cost estimating method developed by T.A. O'Hara and published in the CIM Bulletin, February 1980, that expresses the major cost components of a Canadian underground mine (shaft, mine development, concentrator, and supporting infrastructure) as power-law functions of mine capacity and, for shaft-related items, depth — calibrated from a historical database of actual Canadian mine costs. Like the Camm (1989) models used in Question 5, it is intended for rapid pre-feasibility/screening-level comparison between design options rather than a bankable estimate, and it was later extended and refined into the Mular & Poulin CapCost system (CIM Special Volume 47, 1998).

1.6 — Methods of hauling rock to surface

Rail haulage. A trolley, battery or diesel locomotive hauls a train of ore cars on fixed track between the working levels and the shaft or a surface portal; it is efficient for long, relatively flat, well-established haul routes but is inflexible — extending or relocating track lags behind mine advance. Typical capacity: roughly 500–2,000 t/day, depending on gauge, locomotive fleet and train length.

LHD + haul-truck (trackless) haulage. A load-haul-dump loader mucks broken rock at the face and either trams it directly to an ore pass/dump, or loads a rubber-tyred underground haul truck for longer runs to the shaft or a ramp to surface; being trackless, it adapts readily to irregular ramp and level geometry and is the dominant system in modern mechanised mines. Typical capacity: roughly 500–5,000+ t/day per unit, scaling with LHD/truck bucket and box size.

Shaft skip hoisting. Rock is dumped into a loading pocket at the shaft bottom, loaded into a skip, and hoisted vertically to surface on a wire rope over a drum or friction (Koepe) hoist — the method used for the deep 425 m production shaft designed in Question 6. It is the highest-capacity method per unit of infrastructure for deep, high-tonnage operations. Typical capacity: several hundred to a few thousand tonnes per hour per hoist (Question 6's example shaft moves 500 t/hr).

Conveyor haulage. A continuous belt, run along a decline, incline shaft or drift, moves broken rock in a steady stream rather than discrete batches; it needs a crusher or sizing station ahead of the belt to control lump size but then offers very high sustained, continuous throughput with low unit power cost, and is the preferred bulk-haulage method for large caving operations. Typical capacity: roughly 1,000–10,000+ t/day continuous, well above the batch methods above at comparable capital intensity.