Contract mining & earthmoving

SpecificationBench height typically 20 to 50 feet (Penn State MNG 230 teaching range). Bench geometry terms are defined off the bench floor (horizontal surface), bench face (vertical surface), face angle (average angle the face makes with the horizontal), toe (intersection of bench face and bench floor) and crest (intersection of bench floor and bench face); bench width is crest to toe measured along the upper surface. Haul roads are built to a gradient of up to 15 percent. Safety or catch benches are left to collect sliding material from the benches above and stop large rock pieces. A safety berm of broken material is built along the crest as a guard rail. Stripping ratio is computed as waste blocks divided by ore blocks. Typical practice, not a published standard: bench height is picked to suit the loading unit and the drill, and waste:ore ratio rises sharply as the pit is pushed out over a steeply pitching orebody.

Where it falls downStops working when the stripping ratio needed to expose the next increment of ore exceeds what the ore value will pay for; the course notes that as a pit expands over a steeply pitching orebody the overall stripping ratio increases dramatically and unit cost rises with depth because material is hauled farther. Wall angle is set by geotechnical conditions, and under 30 CFR 56.3130 mining methods must maintain wall, bank and slope stability where persons work or travel, with bench width and height based on the equipment used to clean benches or scale walls. 30 CFR 56.3131 requires loose or unconsolidated material to be sloped to the angle of repose OR stripped back for at least 10 feet from the top of the pit or quarry wall; the two are alternatives, not a single stripping requirement.

Specificationaustralian practice, from Mitra and Saydam, Surface Coal Mining Methods in Australia (IntechOpen, 2012), not a US published standard: dragline strips are typically 40 to 90 m wide and a few kilometres long, draglines can excavate a maximum of about 50 to 80 m of overburden owing to reach and dump height limits, and the method suits gently dipping deposits because spoil is unstable on steep dips. Example machine reach and bucket combinations cited by Penn State: 200 ft reach with a 200 yd3 bucket, or 300 ft reach with a 120 yd3 bucket, at a typical cycle time of 45 seconds. Where cast blasting is used ahead of the dragline, up to 40 percent of the overburden is actually thrown or cast into the adjacent strip, and a single cast shot may use a million pounds of ANFO or more.

Where it falls downBounded by dragline reach and spoil-room geometry rather than by grade: once overburden thickness exceeds what the machine can cast and rehandle, the method gives way to truck-and-shovel prestripping or terrace mining. Steeply dipping seams are excluded by spoil instability. Under 30 CFR 816.67 - the SMCRA surface coal mining rule, which does reach this method because it is overwhelmingly coal - flyrock must stay within the area of control and the permit boundary and must not be cast more than half the distance to the nearest dwelling or other occupied structure, which constrains cast-blast energy near boundaries.

SpecificationWhere overburden across a hill is less than 165 feet, outcrop-to-outcrop mining becomes possible. Auger mining recovers around one-third of the coal under the highwall, with maximum recovery less than 35 percent; cutting heads reach as large as 7 feet in diameter and holes may be drilled more than 300 feet deep. A highwall miner can advance 1,200 ft under the highwall versus a typical 300 ft for an auger, giving recovery near 50 percent for the highwall miner against roughly 30 percent for the auger. Terrace mining dimensions are not published as a numeric standard in the source consulted.

Where it falls downAuger and highwall recovery is capped by the web pillars that must be left for stability, which is why recovery is a third to a half rather than the 60 percent-plus of a room-and-pillar panel. Contour mining stops where the highwall height makes further cut uneconomic or unstable. The mountaintop removal variance under SMCRA section 515(c) and 30 CFR 785.14 requires that no highwall remain, which controls where the method can be permitted.

SpecificationVertical quarry walls of 1,000 ft or more can be excavated in dimension stone operations. Bench heights, block dimensions and saleable-block recovery percentages were not given numerically in the sources opened.

Where it falls downStops working where the rock is jointed, weathered or stained beyond the tolerance of the finished-stone market; fracture spacing rather than grade sets the limit. Saw-based extraction rates are an order of magnitude below drill-and-blast, so the method only pays where the block price is high.

SpecificationPlacer dredges have ranged from 1-1/2 to 20 cubic feet of bucket capacity. Bucket size sets digging depth, but the bucket-to-depth pairs previously given here did not match IC 8462's table and have been withdrawn pending a direct read of the original. What the circular does state: the maximum digging depth recorded to date was 162 feet, and 18-cubic-foot dredges in Malaysia dug 130 to 158 feet below water level. Monthly production for a 3-cubic-foot dredge is given as 60,000 cubic yards at slow speed. Target running time is about 21.5 digging hours per day, roughly 90 percent of available time, with actual performance from 75 percent or less in difficult digging to about 93 percent in very favourable ground. A 9 cubic foot class dredge uses about 75 gallons of water per minute and an 18 cubic foot class about 150 gpm. Dredging is most applicable on surface grades up to 2 percent, with isolated cases up to 6 percent. Dragline placer methods are limited to about 60 feet maximum efficient depth.

Where it falls downFails on cemented or boulder-rich gravels, on steep surface grades, where bedrock is too deep for the bucket line, and where water is unavailable. Recovery losses are inherent: gold left in depressions and crevices in bedrock, gold lost to a poor mining pattern, and gold lost in the plant when not washed free of oversize or when it adheres to clay. The source illustrates how dilute the feed is: gravel averaging 25 cents of gold per cubic yard at a 35 dollar gold price contains by weight about 1 part gold in 6.5 million parts waste.

SpecificationFive-spot and seven-spot well patterns are common, with wells typically 12 to 50 m (40 to 150 ft) apart. Ore zones lie about 100 to 300 m (328 to 984 ft) deep in many districts, and greater than 560 m (1,840 ft) at one New Mexico deposit. New Mexico deposits run about 0.2 to 0.3 percent U3O8 by weight and Wyoming deposits about 0.1 to 0.25 percent. Alkaline lixiviant chemistry typically runs sodium from under 400 up to 6,000 mg/L, bicarbonate from under 400 to 5,000 mg/L and pH from 6.5 up to 10.5; NRC lists the leach solution as water mixed with oxygen and/or hydrogen peroxide plus sodium carbonate or carbon dioxide. Licensed circulation flow rates typically range from about 15,100 to 34,000 L/min (4,000 to 9,000 gal/min), with a production bleed of about 1 to 3 percent of the circulation rate to hold the wellfield under inward hydraulic gradient. Greater than 95 percent of the uranium in solution is extracted during the ion-exchange step. About 14 ISR facilities exist in the United States (NRC, In Situ Recovery Facilities, page last updated 14 September 2026, https://www.nrc.gov/materials/uranium-recovery/extraction-methods/isl-recovery-facilities.html; this count is not in the GEIS chapter cited below).

Where it falls downRequires permeable sandstone host beds confined above and below by low- or semi-permeable claystones, siltstones, mudstones or shales; without those confining units the lixiviant cannot be controlled. Roll fronts are crescent-shaped, may extend hundreds of metres but may be only a few metres thick, so a thin or discontinuous front gives poor sweep. Groundwater restoration to the standard in 10 CFR 40 Appendix A, Criterion 5(B)(5) is a condition of the method, not an afterthought.

SpecificationExtraction ratios are usually more than 60 percent; NIOSH case histories of collapses record extraction ratios of about 70 to 78 percent in the failed areas. Pillar width-to-height ratio is the key design number: failures occurred at ratios always less than 3 in coal, usually much less than 1 in metal mines and less than about 2 in nonmetal mines, while the prevention approach uses panel pillars with high width-to-height ratios greater than about 3 or 4 and barrier pillars typically greater than about 10. Under 30 CFR 75.220 each coal operator must develop and follow a District Manager-approved roof control plan.

Where it falls downStops where pillar stress exceeds pillar strength: collapses are more prevalent in shallow mines less than 100 m deep where slender pillars were left, and the containment approach deliberately uses low width-to-height panel pillars (0.5 to 2 range) with barrier pillars around them so a collapse cannot run. Minimum panel widths for the containment-by-barrier approach are at least 80 m, with at least five and usually more than 10 pillars across the panel width. Significant dip, very weak roof, or ore too valuable to leave in pillars all push the operator to another method.

SpecificationPanels are typically three to four kilometres long and 250 to 400 metres wide; one worked teaching example is 300 m wide by 2 km long by 3.5 m tall. A plough is preferred for seam heights up to 1.8 m and a shearer for seams from 2.3 to over 6 m. A shearer operating at 16 metres per minute yields a capacity of 5,000 tons per hour. Hydraulic roof supports carry bearing forces up to 1,750 tons. A three-entry gate road system is common US practice, but it is practice and not law: no MSHA standard requires three gate roads or three gate road entries, and two-entry systems operate under approved roof control and ventilation plans. What federal law does require is two separate escapeways from each working section (30 CFR 75.380). Caving may be delayed 10 to 20 or more passes of the shearer at the start of a panel.

Where it falls downNeeds continuity: faults, washouts, thickness changes and igneous intrusions that a room-and-pillar section can mine around will stop a longwall face. Requires that surface subsidence be acceptable, since the roof is deliberately caved. Requires long lead-time gate road development, so the method is committed years before the panel produces.

SpecificationCut or slice height is typically at least 10 ft but no more than 30 ft, set by ground control. Stope height is typically 150 to 300 ft, stope length on the order of 200 to 2,000 ft, and stope width at least 6 ft and up to 100 ft. Three backfill types are used: hydraulic fill of tailings at 55 to 70 percent solids with 3 to 4 percent cement and a 10 percent cement mix to top off; paste fill of unclassified tailings up to 88 percent solids with only a few percent cement; and cemented rock fill of waste rock with cement added. Backfill can represent 10 to 20 percent of the mining cost, of which nearly 75 percent is the cost of the cement.

Where it falls downLow productivity and high cost per tonne confine it to high-value metalliferous ores; the fill cycle interrupts production in every stope. It is also the method of choice for recovering pillars left by other methods. Where the ore is weak as well as the walls, the cut has to shrink toward drift-and-fill dimensions.

SpecificationDescribed by Penn State as small slice progression in weak ore zones within the cut-and-fill family, so the slice geometry and fill specifications of cut and fill apply: cuts of at least 10 ft and no more than 30 ft, and hydraulic, paste or cemented rock fill with the cement contents listed under cut and fill. No separate published drift dimension range was found in the sources opened; typical practice, not a published standard, is that the drift is sized to the jumbo and loader.

Where it falls downThe lowest-productivity method in the underground set, because every tonne is mined through a development-sized opening and every opening has to be filled. Only viable where the metal value per tonne is very high or where the alternative is sterilising the block.

SpecificationApplies to narrow veins as thin as 3 ft or thicker veins of 100 ft or so. Stope length typically 150 to 300 ft and stope height 200 to 300 ft. Ore swell on blasting may be as much as 30 percent or even more, and that swell fraction is what is drawn during mining. Ore may remain in the stope for six months to a year before final drawdown.

Where it falls downTies up working capital because the broken ore sits in the stope for months; requires ore that will not oxidise, pack or cement while it sits. Requires walls competent enough to stand with only broken-muck support. Not suitable where the mine needs to blend grades continuously.

SpecificationBig-hole stoping drills on the order of 300 ft in length at approximately 6 inches in diameter. A slice of approximately 10 to 15 ft is blasted with each round. Sublevel spacing for ring drilling is limited by the accuracy of the ring drills. Penn State describes the method as productive and large-scale with low dilution. Crown and sill pillars are left unmined and hold significant ore of considerable value that cannot be economically extracted without collapse risk.

Where it falls downRequires host rock reasonably strong enough to prevent cave-in, because vast open spaces are created and stand unsupported. Irregular ore boundaries defeat it, since the drill pattern is committed before the ore is seen. Ore tied up in crown and sill pillars is a permanent loss unless a later backfilled pillar-recovery campaign is run.

SpecificationDrilled on a grid on the order of 12 ft by 12 ft with holes on the order of 6 inches in diameter, designed to take off a 10 ft slice of ore per round.

Where it falls downSame constraints as shrinkage stoping: the ore remains in the stope for months so its mineral properties must be stable, and the method demands high grade and a valuable commodity that is uniform in quality. Not a blending method. Large-diameter hole deviation over the full stope height limits how tall a stope can be drilled accurately.

SpecificationFrom a global review of 24 sublevel caving operations (21 active plus 3 ceased): most modern operations achieve between 80 and 90 percent ore recovery at draw percentages of 100 to 120 percent, with ore loss typically 10 to 20 percent of the orebody. Measured dilution across the benchmarked mines ranged as low as 4 percent and as high as 43 percent. Sublevel spacing across the surveyed operations ranges from about 15 to 30 m centre to centre. Width of draw generally varies between 9 and 12 m with an average of about 11 m; crosscut spacing ranged 12 to 24.75 m centre to centre. Drive widths across the benchmarked operations ran 4 to 7.5 m. Fan-pattern mines drill and charge to about 35 to 40 m; silo designs use blastholes up to 60 m. Peak production rates ranged from 0.2 Mt/a at the smallest to 27 Mt/a at the largest, with most operations between 1 and 6 Mt/a.

Where it falls downThe whole hangingwall must be allowed to cave and the surface must be allowed to subside. Development intensity is high because every sublevel needs a full drift network, which makes development metres per tonne much worse than block caving. Dilution rises steeply if draw is continued past the arrival of the waste front, which is why draw control is the method's governing discipline rather than an administrative task.

SpecificationOre blocks can be up to 500 metres in height and several hundred metres in length. Undercut drilling uses holes of 2 to 3 inches. Development can take up to several years of advance work before the first tonne is drawn. Penn State characterises it as a high-productivity method with low operating costs and a long development time before production. Drawpoint spacing, dilution and recovery percentages were not given numerically in the sources opened.

Where it falls downThe orebody must cave under its own weight; if the ore is too competent the cave hangs up and the block is lost. Must be massive both horizontally and vertically. Subsidence of the ground surface above the orebody has to be acceptable, with the subsidence zone deepening and widening as extraction continues. The multi-year development lead means the method commits capital long before it produces, and a mis-set drawpoint layout cannot be changed once the cave is running.