Advanced Mining interview questions

125 questions with answers (MIN-301 to MIN-425). Try each one aloud before reading the answer.

Mining Geology & Exploration

MIN-301 · A gold vein is 200 m along strike, 100 m down dip, 2 m true thickness, density 3.0 t/m³ and grade 6 g/t. Estimate tonnage and contained gold.

Volume = 200 × 100 × 2 = 40,000 m³. Tonnage = 40,000 × 3.0 = 120,000 t. Gold = 120,000 × 6 = 720,000 g = 720 kg, or about 23,150 troy oz (÷ 31.10 g/oz).

MIN-302 · How would you design an exploration programme for a porphyry copper target concealed under 50 m of cover?

Start with regional magnetics and gravity to map intrusions and structures, then IP and resistivity to detect chargeable disseminated sulphides beneath cover. Use geochemistry of the cover or groundwater where possible, then drill scout holes on coincident anomalies, followed by systematic core drilling on a grid to define the deposit.

MIN-303 · Calculate the composite grade over a drill intersection of 1.5 m at 8 g/t, 2.0 m at 3 g/t and 0.5 m at 12 g/t.

Length-weighted grade = Σ(length × grade) ÷ Σ length = (12 + 6 + 6) ÷ 4.0 = 6.0 g/t over 4.0 m. If densities differ, weight by length × density.

MIN-304 · What does Gy's sampling theory say about sample preparation?

The fundamental sampling error variance is proportional to the cube of the top particle size and inversely proportional to sample mass. Samples must therefore be crushed finer before each mass reduction step, and a sampling nomogram is used to design the crushing and splitting protocol.

MIN-305 · How do you handle a strong nugget effect in a gold deposit?

Use larger samples (bulk samples, large-diameter core, screen fire assays), closer drill spacing, duplicate analysis, top-cutting of outliers and estimation methods such as kriging with a high nugget or indicator kriging. Grade control in the pit or stope is critical to reconcile estimates.

MIN-306 · Why and how are high-grade outliers capped (top-cut)?

A few extreme assays can overstate the resource if spread across large volumes. A cap value is chosen from log-probability plots, histograms or metal-at-risk analysis, and assays above it are reset to the cap before compositing and estimation.

MIN-307 · How do explicit wireframe and implicit geological modelling differ?

Explicit modelling digitises ore boundaries section by section and joins them into wireframes, which is slow and subjective. Implicit modelling builds 3D surfaces mathematically (e.g. radial basis functions) directly from drill data, making it faster, reproducible and easy to update.

MIN-308 · What is twin-hole drilling and why is it done?

A new hole, usually core, is drilled very close to an existing hole (often RC) to check whether grades and intervals are reproduced. It validates older or cheaper drilling data before it is used for resource estimation.

MIN-309 · Why are downhole surveys important?

Drillholes deviate with depth due to rock foliation and drilling conditions. Without gyroscopic or magnetic downhole surveys, intercepts are plotted in the wrong position, distorting ore body shape, thickness and tonnage estimates.

MIN-310 · How is in-situ bulk density determined and why does it matter?

By weighing core samples in air and in water (with wax coating for porous samples), or by gamma–gamma density logs. Tonnage is directly proportional to density, so a 5% density error gives a 5% tonnage error.

MIN-311 · How does structural complexity affect coal exploration?

Faults, dykes, sills, washouts and seam splitting reduce continuity, so closer borehole spacing, seismic surveys and detailed structural mapping are needed. These features also drive mine layout, panel size and the choice between opencast and underground methods.

Mine Development

MIN-312 · A skip carries 20 t per trip with a 90 s cycle time and the shaft hoists 20 hours a day. What is the hoisting capacity?

Trips per hour = 3600 ÷ 90 = 40. Hourly capacity = 40 × 20 = 800 t/h. Daily capacity = 800 × 20 = 16,000 t/day, or about 5.3 Mt/year at 330 days.

MIN-313 · For a 2 Mtpa ore body at 800 m depth, how would you choose between shaft and decline access?

At 800 m a decline would be about 6–7 km long, so truck haulage cost, cycle time, heat and ventilation become large. A shaft has higher capital and lead time but lower operating cost at this depth and rate. Many operations use a decline for early access and development and a shaft for production hoisting.

MIN-314 · Size a circular ventilation shaft to carry 300 m³/s at a maximum velocity of 10 m/s.

Required area A = Q ÷ v = 300 ÷ 10 = 30 m². Diameter D = √(4A/π) = √(38.2) ≈ 6.2 m, before adding area for fittings, ladders or conveyances.

MIN-315 · A decline must reach 400 m vertical depth at 1 in 8. What is its length?

Horizontal length = 400 × 8 = 3,200 m. Slope length = √(3200² + 400²) ≈ 3,225 m, plus additional length for curves, stockpile bays and level accesses.

MIN-316 · What must be considered when designing ground freezing for a shaft?

Water table and groundwater flow velocity (flowing water resists freezing), ground thermal properties, freeze-hole spacing and verticality, brine temperature, required ice wall thickness for earth and water pressure, freeze-up time and temperature monitoring holes.

MIN-317 · How would you estimate monthly development advance for a mine with 4 m rounds?

With one round per heading per shift, two shifts a day and 25 days a month: 4 × 2 × 25 = 200 m per heading per month. Real advance is limited by support time, mucking, ventilation re-entry and equipment availability, so several headings are worked in parallel.

MIN-318 · A winding rope has a breaking load of 1,200 kN and the maximum suspended load is 120 kN. What is its static factor of safety?

FoS = breaking load ÷ maximum static load = 1200 ÷ 120 = 10. Dynamic loads from acceleration and braking reduce the effective factor, so these must be checked against statutory requirements.

MIN-319 · How is a protective shaft pillar sized using angle of draw?

Radius of influence at surface ≈ depth × tan(angle of draw). For a 300 m deep seam and 35° angle of draw, the extra radius is 300 × 0.70 ≈ 210 m, added to the radius of the protected area and a safety margin.

MIN-320 · How would you advance a drift through a water-bearing fault zone?

Probe drill ahead to locate the zone and water, pre-grout to reduce inflow, drive short rounds with spiling or forepoling, install shotcrete and steel sets promptly, monitor convergence and provide pumping and drainage. Excavation should be designed to keep the fault zone crossing short and at a high angle.

MIN-321 · How would you design a dewatering system for 500 m³/h inflow at 600 m depth?

Pump power ≈ ρgQH/η = 1000 × 9.81 × 0.139 × 600 ÷ 0.7 ≈ 1.17 MW. Use multistage centrifugal pumps in stages or direct from a main pump station, with standby units, sumps sized for several hours of inflow plus settling, and dual rising mains.

MIN-322 · What are the trade-offs between underground crushing with shaft hoisting and truck haulage up a decline?

Crushing and hoisting gives low unit cost and high capacity at depth but needs a large crusher chamber and shaft capital. Truck haulage is flexible and has low capital but cost, heat and diesel emissions rise sharply with depth. A conveyor decline is a middle option.

Surface Mining

MIN-323 · What are pit optimisation algorithms?

Lerchs–Grossmann (graph theory) and pseudoflow (network flow) find the ultimate pit that maximises undiscounted value of a block model, respecting slope constraints. Software runs them at varying revenue factors to produce nested pits for phase design.

MIN-324 · Ore value is ₹3,000/t, mining and processing cost ₹1,800/t and stripping cost ₹150/m³. What is the break-even stripping ratio?

BESR = (3000 − 1800) ÷ 150 = 8 m³ of waste per tonne of ore. Ore at the pit limit with a stripping ratio higher than 8 m³/t would lose money.

MIN-325 · How does steepening the overall slope angle affect pit economics?

Steeper slopes reduce waste stripping, often substantially in deep pits, raising NPV. The gain must be balanced against higher failure risk, so slopes are designed at bench, inter-ramp and overall levels to a target factor of safety and probability of failure.

MIN-326 · Size a truck fleet for 2,000 t/h using 100 t trucks with a 25 min cycle and 80% efficiency.

Output per truck = 100 × (60 ÷ 25) × 0.8 = 192 t/h. Trucks needed = 2000 ÷ 192 ≈ 10.4, so 11 operating trucks plus spares to allow for availability.

MIN-327 · A truck of 300 t gross weight climbs an 8% ramp with 3% rolling resistance. What resistance must it overcome?

Total resistance = 8% + 3% = 11% of gross weight = 0.11 × 300 = 33 t ≈ 33,000 × 9.81 ≈ 324 kN. The required rimpull is read from the truck's rimpull–speed chart.

MIN-328 · How is a dragline selected for a strip mine?

Required reach, dumping height and digging depth are found from the strip geometry (overburden depth, strip width, spoil angle and swell). The bucket size is then chosen to meet the annual overburden volume at expected cycle time, availability and utilisation.

MIN-329 · Slope radar shows accelerating wall displacement. What do you do?

Increase monitoring frequency, apply the inverse-velocity method to forecast failure time, set and enforce trigger action response plans, evacuate people and equipment from the danger zone, restrict access and investigate the cause (water, undercutting). Re-entry only follows stabilisation and engineering review.

MIN-330 · What are nested pits and revenue factors?

By running pit optimisation with the metal price scaled by revenue factors (e.g. 0.3 to 1.2), a series of nested pits is generated. The smaller pits have the highest value per tonne and guide pushback design and mining sequence.

MIN-331 · How would you manage an opencast coal mine expanding toward a village?

Shift to controlled blasting with electronic detonators and limited charge per delay to keep PPV and airblast within DGMS limits, use surface miners near habitation, monitor vibration and dust, build green belts and noise barriers, and engage the community about rehabilitation and resettlement.

MIN-332 · Why are pit slopes depressurised?

Pore water pressure reduces effective normal stress and thus shear strength on potential failure surfaces. Horizontal drains, dewatering wells and drainage galleries lower pressures, allowing steeper and safer slopes.

MIN-333 · What are the key considerations in designing an overburden dump?

Foundation investigation, lift height and berm width, overall slope angle, material shear strength, drainage blankets and toe drains, a factor of safety target (usually 1.2–1.5 depending on consequence), and final reclaimed profile.

Underground Coal Mining

MIN-334 · Find the tributary-area pillar stress for 25 m square pillars, 4.8 m galleries, depth 300 m and γ = 0.025 MN/m³.

σv = 0.025 × 300 = 7.5 MPa. Area ratio = (25 + 4.8)² ÷ 25² = 888 ÷ 625 ≈ 1.42. Pillar stress ≈ 7.5 × 1.42 ≈ 10.7 MPa. With a pillar strength of 16 MPa, FoS ≈ 1.5.

MIN-335 · How is the capacity of longwall powered supports selected?

By estimating the load of the immediate roof and part of the main roof that the support must carry, using methods such as the detached block method, caving height from bulking factor, and numerical models, then checking periodic weighting. Support density (t/m²) and yield setting are chosen with a safety margin.

MIN-336 · A longwall emits 20 m³/min of methane. How much air keeps return methane below 1%?

Q = q ÷ (c_limit − c_intake) = 20 ÷ (0.01 − 0) = 2,000 m³/min ≈ 33 m³/s. In practice a higher quantity is provided to allow for peaks and uneven mixing, and drainage may be added.

MIN-337 · Graham's ratio in a sealed goaf is rising steadily. What actions are needed?

Sample sealed-area gases frequently, check CO, H₂, C₂H₄ and Graham's ratio trends, reduce leakage by pressure balancing and sealing cracks, inject nitrogen to inert the goaf, and prepare to strengthen seals and withdraw workers if explosive mixtures or fire indicators appear.

MIN-338 · How would you plan extraction of coal beneath a river?

Predict subsidence and strata cracking, then keep extraction partial (bord and pillar without depillaring, narrow panels) or use stowing so that cracks do not connect to the river. Maintain adequate parting, monitor water inflow and obtain DGMS permission with conditions.

MIN-339 · What is multi-seam interaction and how is it managed?

Pillars or remnant edges in one seam concentrate stress on seams above and below, causing roof and pillar failures. It is managed by superimposing pillars (columnisation), mining the upper seam first where possible and avoiding pillar edges in the adjacent seam.

MIN-340 · What is a coal and gas outburst and how is it controlled?

An outburst is a sudden violent ejection of coal and gas from the face, caused by high gas pressure, stress and weak coal. Control includes gas drainage ahead of mining, measuring gas content and desorption indices, stress relief by protective seams and remote blasting.

MIN-341 · How is a chain pillar designed in longwall mining?

Chain pillars between gate roads must protect the tailgate from front and side abutment loads. They are designed for development load and second-panel loading using empirical methods (e.g. ALPS), the Wilson yield-pillar approach or numerical models, checking road stability rather than the pillar alone.

MIN-342 · What are the hazards of depillaring a thick seam with a large hanging roof?

A strong overlying sandstone may not cave, leaving a large hanging area that can collapse suddenly, causing an air blast and pillar overriding. Controls include induced blasting to cave the roof, limiting the goaf area, monitoring, and keeping workers clear during critical falls.

MIN-343 · For a 10 m thick seam, how would you choose between LTCC and blasting gallery?

LTCC suits large, regular panels and high production with longwall equipment but needs high capital and good cavability of top coal. The blasting gallery method needs less capital, suits smaller irregular areas and existing bord and pillar development, but has lower output and depends on blasting.

MIN-344 · What are CBM, CMM and VAM?

Coal bed methane is gas extracted from virgin seams through surface wells. Coal mine methane is drained from active or abandoned workings. Ventilation air methane is very dilute methane (usually below 1%) in mine exhaust air, which needs special oxidisers to use or destroy.

Underground Metal Mining

MIN-345 · Which method would you choose for a steep 5 m wide vein with a weak hanging wall?

Sublevel open stoping would cause hanging wall failure and dilution. Mechanised cut and fill (or a short-hole retreat with paste fill) supports the hanging wall progressively and controls dilution, at higher cost per tonne. The choice depends on grade justifying the cost.

MIN-346 · What is the Mathews stability graph?

It plots a stability number N′ (from modified Q, stress factor, joint orientation and gravity factors) against the hydraulic radius of each stope surface. Zones on the graph indicate stable, unsupported, supported or caving conditions, guiding stope dimensions.

MIN-347 · Find the hydraulic radius of a stope hanging wall 60 m long and 30 m high.

HR = area ÷ perimeter = (60 × 30) ÷ (2 × (60 + 30)) = 1800 ÷ 180 = 10 m. This value is used with the stability number in the Mathews graph.

MIN-348 · What must be addressed in block cave design?

Caveability (hydraulic radius of the undercut needed to start and sustain caving), primary and secondary fragmentation, drawpoint spacing and layout, undercut strategy (pre- or post-undercut), extraction level stability, draw control, subsidence extent and risks such as mud rush and air blast.

MIN-349 · How is cemented paste fill designed for exposure?

The required strength depends on the height and width of the exposed face, fill density and whether it is a vertical wall or undercut. Analytical methods (e.g. Mitchell's sliding wedge) set strength targets; binder content (often a few percent) is chosen from laboratory curing tests.

MIN-350 · How should stopes be sequenced in a deep, high-stress ore body?

Use a sequence that pushes stress toward abutments away from active areas, such as centre-out or a pyramid sequence without leaving isolated pillars. Avoid creating highly stressed remnant pillars and sequence to keep mining fronts retreating toward stable ground.

MIN-351 · Estimate the hourly output of an LHD with a 10 t bucket, 4 min cycle and 75% efficiency.

Cycles per hour = 60 ÷ 4 = 15. Output = 10 × 15 × 0.75 = 112.5 t/h. Fleet size follows from daily production divided by hours per day and this rate.

MIN-352 · What is equivalent linear overbreak/slough (ELOS)?

ELOS is the volume of unplanned dilution from a stope wall divided by the wall area, giving an average depth of overbreak in metres. It links Mathews stability results to dilution estimates for planning.

MIN-353 · What are the issues in transitioning from open pit to underground mining?

Timing the switch based on stripping ratio and NPV, designing a crown pillar or sill to separate the two, managing water and rock falls from the pit, overlapping production during the ramp-up, and selecting a suitable underground method. India's Rampura Agucha zinc mine made such a transition.

MIN-354 · What challenges arose in very deep mines such as the Kolar Gold Fields?

At over 3 km depth, very high rock stress caused frequent rock bursts, virgin rock temperatures exceeded about 60 °C requiring refrigeration, and hoisting and ventilation costs became very high. Combined with low grades and prices, these led to closure in 2001.

MIN-355 · How can narrow-vein mining be mechanised?

Using small LHDs, narrow-vein jumbos and longhole drills, sublevel retreat with short sublevel spacing, Alimak or raise-bored slots, and resue or selective blasting to separate ore from waste. The aim is to reduce dilution and improve safety and productivity.

Rock Mechanics & Ground Control

MIN-356 · What are rock bursts?

Rock bursts are sudden, violent failures of highly stressed rock around excavations, ejecting rock and releasing seismic energy. They occur in deep or high-stress mines, particularly in brittle rock, near pillars, dykes and faults, and are controlled by sequencing, destressing and dynamic support.

MIN-357 · A circular tunnel is in a field with σv = 30 MPa and σh = 15 MPa. Find the boundary tangential stresses.

At the sidewalls: σθ = 3σv − σh = 90 − 15 = 75 MPa. At the crown and floor: σθ = 3σh − σv = 45 − 30 = 15 MPa. Sidewalls are most likely to spall.

MIN-358 · Using the Q-system, estimate rock bolt length for a 6 m span with ESR = 1.6.

Barton's formula: L = 2 + 0.15 B/ESR = 2 + 0.15 × 6 ÷ 1.6 ≈ 2.6 m. Bolt spacing and shotcrete thickness are then read from the Q support chart using the equivalent dimension De = 6 ÷ 1.6 = 3.75.

MIN-359 · Using the Obert–Duvall formula, find pillar strength for in-situ strength 20 MPa and w/h = 3.

σp = σc(0.778 + 0.222 w/h) = 20 × (0.778 + 0.666) ≈ 28.9 MPa. Comparing with the pillar stress gives the factor of safety.

MIN-360 · How is seismicity monitored and managed in burst-prone mines?

Microseismic arrays of geophones locate events and measure energy and magnitude. Event clustering and rates indicate hazardous areas, leading to re-entry exclusion times after blasting, modified sequencing, destress blasting and dynamic support.

MIN-361 · A 1.5 m thick weak roof layer must be suspended by bolts on a 1.5 m × 1.5 m pattern with FoS 2 (γ = 0.025 MN/m³). What load must each bolt carry?

Dead weight per bolt = 1.5 × 1.5 × 1.5 × 0.025 = 0.084 MN ≈ 84 kN. With FoS 2, each bolt must be designed for about 170 kN, and anchored in competent strata above.

MIN-362 · What numerical modelling methods are used in rock mechanics?

Finite element (FEM) and finite difference (e.g. FLAC) for continuum problems, boundary element (BEM) for tabular stress analysis, and distinct element (e.g. UDEC, 3DEC) for blocky jointed rock. The choice depends on whether discontinuities dominate behaviour.

MIN-363 · How is longwall subsidence predicted?

Empirical methods (e.g. NCB Subsidence Engineers' Handbook) relate maximum subsidence to extracted height, width-to-depth ratio and goaf treatment; for wide caved panels maximum subsidence is about 0.9 of seam height. Influence function and numerical methods are also used.

MIN-364 · What is dynamic (yielding) support?

Support designed to absorb kinetic energy from rock bursts: yielding bolts (e.g. cone or D-bolts) that elongate under load, high-energy mesh and lacing, and fibre-reinforced shotcrete. The system must retain rock while displacing without breaking.

MIN-365 · How are in-situ stresses measured?

By overcoring (measuring strain relief around a gauge in a borehole), hydraulic fracturing (pressure to open and close a fracture), flat-jack tests, and indirect methods such as borehole breakout analysis. Results feed stress inputs to design.

Drilling & Blasting

MIN-366 · How do you estimate the burden for a bench blast using Langefors' approach?

Langefors relates maximum burden to hole diameter, explosive strength, charge concentration and rock constant: roughly B ≈ (d/33)·√(ρ·s/(c·f·E/B)), giving burdens of about 25–40 hole diameters. Practical burden is then reduced for drilling deviation (about 0.05 m + 0.03 m per metre of hole depth).

MIN-367 · A 115 mm hole is charged with emulsion of density 1.2 g/cc. What is the linear charge concentration?

q = π/4 · d² · ρ = 0.7854 × 0.115² × 1200 ≈ 12.5 kg/m. Multiplying by the charged column length gives charge per hole, from which powder factor is checked against burden × spacing × bench height.

MIN-368 · How would you design a blast to control flyrock near a village?

Increase stemming length (about 0.7–1.0 × burden) with crushed angular stemming, check the front-row burden from a face profile survey, avoid over-charging in voids or soft seams, orient the free face away from the village, use blasting mats if needed, and enforce a clearance zone based on a flyrock prediction model.

MIN-369 · A blast's muckpile is too coarse for the crusher. What design changes would you test first?

Reduce burden and spacing or raise the powder factor modestly, use a higher-energy explosive in hard zones, shorten stemming only if flyrock allows, improve drilling accuracy, and use electronic delays to improve interaction between holes. Measure results with image-based fragmentation analysis before changing everything at once.

MIN-370 · What are the advantages of electronic detonators in large blasts?

Millisecond-accurate, programmable delays let you control fragmentation, muckpile shape and vibration frequency; each hole can be logged and tested before firing; and scatter-related misfires and out-of-sequence initiation are reduced. Cost per detonator is higher, so benefits must come from downstream gains.

MIN-371 · How is presplit blasting designed for a final pit wall?

Closely spaced holes (about 8–12 hole diameters) along the final wall are lightly decoupled-charged and fired before the production blast to create a crack plane. The crack reflects stress waves and protects the wall. Linear charge is chosen so borehole pressure exceeds tensile strength but stays below the crushing strength.

MIN-372 · How do you reduce ground vibration when the PPV limit is exceeded?

Reduce the maximum charge per delay (deck the holes, use smaller diameters or electronic timing), increase distance, improve confinement so energy goes into breaking rock, and adjust delays to avoid frequencies that match nearby structures. Monitor with seismographs and refit the site's scaled-distance law.

MIN-373 · Why is a regression of PPV versus scaled distance site-specific?

PPV = K·(D/√Q)^(−β) depends on geology, joints, overburden and charge confinement, so K and β vary widely between sites. A generic law can be unsafe or over-conservative; the site law should be built from monitored blasts and used at an upper confidence line.

MIN-374 · How do you plan a tunnel blast round with a burn cut?

Drill a cluster of large empty relief holes near the face centre with closely spaced charged holes around them to create the opening. Then fire stopers, easers, contour and lifter holes in delay sequence outward. Pull depth is limited to roughly 90–95% of the relief hole diameter-based limit, and contour holes use light charges for smooth blasting.

MIN-375 · How would you investigate a misfire after a blast?

Keep the area cleared for the prescribed waiting time, have a competent person inspect, identify the misfired hole, and either refire it after re-priming or wash out the explosive per procedure; never drill into or near the old hole. Record the cause (cut-off, damaged lead, desensitised explosive) and correct the practice.

Mine Ventilation & Gases

MIN-376 · How do you calculate the air quantity required for diesel equipment underground?

A common rule is a ventilation rate per kW of rated diesel power (for example about 0.06 m³/s per kW, varying by jurisdiction), summed for equipment working at the same time in a district with a utilisation factor. Gas dilution, dust, heat and minimum velocity requirements are then checked, and the largest governs.

MIN-377 · How is the Atkinson equation used in ventilation design?

Frictional pressure loss p = k·L·Per·Q²/A³, so resistance R = k·L·Per/A³ and p = R·Q². It lets you compute airway resistance from friction factor, length, perimeter and area, combine airways in series and parallel, and size fans for the required quantity.

MIN-378 · Two airways of resistance 0.4 and 0.9 Ns²/m⁸ are in parallel. What is the combined resistance?

For parallel airways 1/√R = 1/√R₁ + 1/√R₂ = 1/0.632 + 1/0.949 = 1.582 + 1.054 = 2.636, so √R = 0.379 and R ≈ 0.144 Ns²/m⁸.

MIN-379 · How would you select a main fan for a mine?

Plot the mine resistance curve (p = R·Q²) and match it to fan curves so the operating point gives the required quantity at good efficiency, with margin for mine expansion and stall-free operation. Consider variable pitch or speed control, booster fans, and redundancy for emergencies.

MIN-380 · How do you assess the risk of spontaneous combustion in a coal mine?

Test coal propensity (crossing-point temperature, R70 or adiabatic oxidation), map high-risk seams and gob areas, and monitor CO, the CO/O₂ deficiency ratio (Graham's ratio) and hydrogen trends. Control with rapid face advance, sealing worked-out areas, inertisation (N₂), pressure balancing and reduced leakage through the gob.

MIN-381 · How does ventilation on demand (VOD) reduce energy cost?

Sensors and tracking of vehicles and people let the control system supply air only where it is needed, adjusting fan speeds and regulators. Because fan power varies with the cube of airflow, a 20% airflow reduction cuts fan power by about half while still meeting gas and diesel dilution needs.

MIN-382 · How would you design cooling for a deep hot mine?

Estimate heat loads from autocompression (about 1 °C per 100 m depth for dry air), virgin rock, machinery and groundwater. Then choose surface bulk air cooling, underground refrigeration plants, spot coolers or ice systems, supported by a wet-bulb temperature limit and work-rest regimes.

MIN-383 · What is the purpose of a methane drainage system, and how is it designed?

It captures gas before it enters the ventilation air, allowing higher production without exceeding methane limits. Pre-drainage boreholes, cross-measure boreholes over the goaf and surface wells feed a vacuum pipeline; design considers gas emission forecasts, capture efficiency, pipeline sizing and safe utilisation or flaring.

MIN-384 · How do you calculate the air needed to dilute a methane emission?

Q = q / (C − C₀), where q is methane emission rate, C the allowable concentration and C₀ the intake concentration. For 0.5 m³/s methane, 1% allowable and 0% intake: Q = 0.5/0.01 = 50 m³/s, before applying a safety factor for peaks.

MIN-385 · How is a ventilation survey carried out?

Measure air velocities (anemometer traverses), cross-sections, pressure differences (gauge and tube or barometers), temperatures and gas concentrations at stations through the network. Results give quantities, airway resistances and leakage, used to calibrate a network model such as VentSim for planning.

Mine Machinery & Transport

MIN-386 · A shovel loads a truck in 2.5 min and the truck cycle is 22 min. How many trucks are needed?

Trucks needed = truck cycle time/loading time = 22/2.5 = 8.8. Using 9 trucks gives a match factor of 9 × 2.5/22 ≈ 1.02, so the shovel stays busy with slight truck queueing; 8 trucks would leave the shovel idle about 9% of the time.

MIN-387 · How is rimpull used to estimate truck speed on a ramp?

Total resistance = grade resistance + rolling resistance, expressed as a percentage of gross vehicle weight. Read the truck's rimpull-speed curve at that effective grade to get the achievable speed, then check retarder curves for downhill loaded travel.

MIN-388 · How do you select between in-pit crushing and conveying (IPCC) and truck haulage?

IPCC suits long-life, deep pits with large tonnage and long hauls, cutting diesel cost, emissions and labour. It needs mine plans that keep the crusher relocatable and conveyors flexible. Trucks give flexibility for selective mining, variable benches and short lives. A cost per tonne over the mine life, including relocation, decides.

MIN-389 · How do you estimate the power required to drive an inclined belt conveyor?

Total power = power to move the empty belt + power to move the load horizontally + power to lift the load (m·g·H per second) + losses at pulleys, skirts and trippers, divided by drive efficiency. On steep, long conveyors the lift component dominates; downhill conveyors may regenerate power.

MIN-390 · How would you plan dragline operations to minimise rehandle?

Choose a mining method (simple side-casting, extended bench or extended key cut) that suits the overburden depth and dragline reach, sequence cuts so spoil lands within reach beyond the pit, and keep a safe highwall and spoil geometry. Rehandle is tracked as a percentage of total overburden and reduced by adjusting strip width and bench height.

MIN-391 · What factors decide hoist selection for a deep shaft?

Depth, payload per trip, cycle time and required daily tonnage determine the hoist type (drum, Blair multi-rope or friction/Koepe), rope size and motor power. Koepe hoists suit deep, high-capacity shafts; double-drum hoists suit moderate depths. Rope safety factors and shaft space also constrain the choice.

MIN-392 · How do you plan maintenance to improve fleet availability?

Use preventive maintenance at fixed hours, condition monitoring (oil analysis, vibration, temperature, payload data) for predictive maintenance, and planned component replacement before failure. Track availability, utilisation, MTBF and MTTR, and prioritise bottleneck equipment.

MIN-393 · How is the productivity of an LHD in a sublevel stoping mine estimated?

Productivity = bucket capacity × fill factor × cycles per hour × utilisation. Cycle time includes loading at the drawpoint, tramming loaded, dumping into the orepass and returning. Tramming distance usually dominates, so orepass spacing strongly affects LHD output.

MIN-394 · What are the key design checks for a haul road?

Width (typically 3–4 × the widest truck width for two-lane traffic), gradient (often 8–10% maximum), structural thickness from CBR methods, wearing course quality, drainage, safety berms of at least half the wheel height, curve superelevation and sight distances.

MIN-395 · How do autonomous haulage systems change mine operations?

They remove drivers from trucks, increasing utilisation and consistency while reducing tyre wear and fuel. They need high-accuracy GNSS, reliable communications, defined traffic rules, segregation from manned vehicles and strong cyber security. Roles shift to remote supervision and maintenance.

Mine Planning, Valuation & Economics

MIN-396 · How does the Lerchs–Grossmann algorithm find the optimal pit?

It treats blocks as nodes with economic values and slope constraints as precedence arcs, and finds the closure with maximum total value (a maximum closure problem). Running it with varying revenue factors produces nested pit shells used for pushback and schedule design.

MIN-397 · How do you calculate the cut-off grade for an open pit?

Marginal cut-off grade = (processing + G&A cost per tonne) / [(price − selling cost) × recovery], in consistent units. It is the grade at which a tonne that is already being mined just pays for its processing; mining cost is excluded because a block inside the pit is mined anyway.

MIN-398 · What is Lane's theory of cut-off grade?

Lane shows the optimum cut-off depends on which stage (mine, mill or market) limits capacity, and that it should be raised early to bring forward higher-grade cash flow because of the time value of money. Cut-off then declines over the mine life as the NPV of remaining reserves falls.

MIN-399 · A project needs $200 M now and yields $40 M per year for 10 years. Is it viable at 10% discount rate?

NPV = −200 + 40 × [1 − 1.1⁻¹⁰]/0.1 = −200 + 40 × 6.145 = −200 + 245.8 ≈ +$45.8 M. NPV is positive, so it is viable at 10%, though sensitivity to price, grade and capex must be tested.

MIN-400 · How is a sensitivity analysis presented for a mining project?

A spider or tornado chart shows how NPV or IRR changes when key inputs (price, grade, recovery, operating cost, capex, discount rate) vary by ±10–30%. It reveals which variables drive value and where more study or risk mitigation is needed.

MIN-401 · How does geological uncertainty affect mine scheduling?

Kriged models smooth grade variability, so schedules based on them can overstate early recoverable grade. Conditional simulation produces many equally probable models; stochastic scheduling then chooses a schedule that maximises expected NPV while limiting the risk of missing ore targets.

MIN-402 · How do modifying factors convert a resource into a reserve?

Mining factors (dilution, recovery, mining method), metallurgical recovery, costs and prices, and legal, environmental, social and government factors are applied to indicated and measured resources in a mine plan. Only material that is economically mineable after these factors becomes probable or proved reserve.

MIN-403 · How do you decide between open-pit and underground mining for a deposit?

Compare value per tonne of ore after mining costs: open pit has lower unit costs but stripping increases with depth. The transition depth is where the incremental stripping cost exceeds the extra cost of underground mining; also consider recovery, dilution, rock mechanics, environmental footprint and timing.

MIN-404 · How does dilution affect project value?

Dilution adds waste tonnes that are mined, hauled and processed but carry little metal, increasing costs and lowering head grade and possibly recovery. Planning should estimate external and internal dilution realistically (from stope design and blasting) because optimistic dilution overstates reserves.

MIN-405 · How is mine closure cost included in valuation?

Closure and rehabilitation liabilities are estimated from the closure plan (demolition, landform shaping, cover, water treatment, monitoring) and included as discounted cash outflows at the end of life, often backed by financial assurance. Long-term water treatment can dominate the cost.

Mine Safety & Legislation

MIN-406 · How would you conduct a risk assessment for a new underground mining method?

Identify hazards with the workforce (bow-tie or HAZOP-style workshops), assess likelihood and consequence, define critical controls for major hazards such as ground falls, fire, explosion and inrush, and assign owners, verification checks and trigger action response plans. Review after trials and incidents.

MIN-407 · What is a Trigger Action Response Plan (TARP)?

A predefined table linking monitored conditions (gas level, convergence, piezometer level) to escalating trigger levels and the actions required at each level, with clear responsibilities. It turns monitoring data into consistent, timely decisions.

MIN-408 · How do you prevent inrush of water into underground workings?

Map aquifers, old flooded workings and faults; maintain statutory barriers; drill advance probe holes when approaching suspected water; design dewatering and pumping capacity with redundancy; and keep emergency escape routes and refuges above flood levels.

MIN-409 · How should a mine emergency response plan be structured?

It defines emergency types, alarm and communication systems, escape routes, refuge chambers and self-rescuer use, mine rescue team organisation and call-out, command structure, external agency liaison and regular drills. It must be practised, not just written.

MIN-410 · How do proximity detection systems reduce accidents?

Tags on people and machines and sensors on equipment create warning and slowdown or stop zones, preventing collisions between mobile equipment and people or other vehicles. They are a control in the hierarchy below eliminating interaction but above procedures and training.

MIN-411 · How do you investigate a serious mining accident?

Secure the site, gather evidence and statements, reconstruct the sequence, and use a structured method (ICAM, root cause analysis) to identify absent or failed controls and organisational factors. Recommendations should target system causes, not just individual error, and be tracked to completion.

MIN-412 · How is respirable crystalline silica exposure controlled?

By engineering controls first: water sprays, dust suppression on drills and crushers, enclosed cabins with filtered air, ventilation and dust extraction. Then administrative controls and personal respirators, with personal sampling against exposure limits and health surveillance.

MIN-413 · How do you manage fatigue in a 12-hour shift operation?

Design rosters to limit consecutive night shifts, provide adequate breaks, use fatigue risk management systems (education, reporting, sleep opportunity monitoring, in-cab fatigue detection) and adjust high-risk tasks. Fatigue is treated like any other critical hazard.

MIN-414 · What are the key principles of tailings dam safety under the Global Industry Standard on Tailings Management?

Accountability at board level, an independent review, an engineer of record, consequence classification, robust design for credible failure modes, monitoring with TARPs, emergency preparedness with affected communities, and public disclosure.

MIN-415 · How do you control the risk of fire on mobile equipment underground?

Fit automatic fire suppression systems, use fire-resistant hydraulic fluids, maintain hoses and electrical systems, clean fuel and oil build-up, carry extinguishers, and plan escape and refuge arrangements. Fire is a major hazard because smoke spreads quickly through ventilation.

Mineral Processing & Mine Environment

MIN-416 · How do you calculate the circulating load in a closed grinding circuit?

From a size-fraction balance around the cyclone: C = (o − f)/(f − u), where f, u and o are the percentage of a chosen size class in the cyclone feed, underflow and overflow. C is the ratio of underflow (returned to the mill) to fresh feed; ball mill circuits typically run at 150–350%.

MIN-417 · How is Bond's work index used to size a ball mill?

W = 10·Wi·(1/√P80 − 1/√F80) gives specific energy (kWh/t) to grind from feed size F80 to product P80 (in µm). Multiplying by throughput gives mill power, adjusted by efficiency factors for mill diameter, size and circuit type.

MIN-418 · How do you calculate metal recovery from assays?

Recovery R = c(f − t) / [f(c − t)] × 100%, where f, c and t are feed, concentrate and tailing grades. For f = 2%, c = 25%, t = 0.2%: R = 25 × 1.8 / (2 × 24.8) ≈ 90.7%.

MIN-419 · How would you improve flotation recovery in a copper concentrator?

Check liberation and grind size, reagent dosage and type (collector, frother, depressant), pH control, pulp density, air rate, froth depth and residence time. Use mineralogy (QEMSCAN/MLA) to find whether losses are coarse unliberated, fine slimes or locked middlings, and target the cause.

MIN-420 · How would you test whether waste rock will generate acid?

Run static acid-base accounting (acid potential from sulphide sulphur vs neutralising potential), net acid generation (NAG) tests, and kinetic humidity cell or column tests over months. Classify rock as potentially acid forming or non-acid forming and plan dump placement and covers accordingly.

MIN-421 · How would you choose between active and passive treatment for mine drainage?

Active treatment (lime dosing, high-density sludge plants) handles high flows and acidity reliably but needs power, chemicals and staff. Passive systems (wetlands, anoxic limestone drains, bioreactors) suit lower flows and loads, especially after closure. Flow, acidity, metal load, land area and long-term cost decide.

MIN-422 · How is a waste rock dump designed for stability?

Analyse foundation conditions, dump height and slope, material strength, water and seismic loading using limit equilibrium methods with suitable factors of safety. Build in lifts with benches, provide drainage and avoid placing dumps on weak or saturated foundations or across drainage lines.

MIN-423 · How is a water balance used in mine environmental management?

It tracks all inflows (rainfall, groundwater, recycled water) and outflows (evaporation, seepage, process use, discharge) over time and climate scenarios. It sizes storages, predicts discharge needs and supports permit compliance and drought or flood planning.

MIN-424 · How is the binder content of cemented paste fill chosen?

Strength requirements set the binder: exposed vertical walls need higher strength (often around 0.5–1 MPa) than fill that is only confined. Lab tests on the mine's tailings give strength versus binder and curing time, and binder is typically 3–8%. Binder is a major cost, so it is minimised per stope.

MIN-425 · How do you plan progressive rehabilitation of a surface mine?

Schedule landform shaping, topsoil return and revegetation as soon as areas are no longer needed, rather than at closure. Set completion criteria (slope, erosion, vegetation cover, water quality), monitor against them, and integrate the plan with mining sequence and closure cost estimates.

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