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By G. Van Weert

Refractory ores and concentrates containing silver tend to form Ag-jarosite during autoclave oxidation. Silver extraction from the oxidation products is therefore low, unless a hot lime treatment step to decompose Ag-jarosite precedes cyanidation. Since silver iodide is very stable but soluble in cyanide solutions, the formation and decomposition of Ag-jarosite in the presence of iodine species was studied. Results indicate that Ag-jarosite formation could be prevented in favour of H3O-jarosite and AgI while oxidizing Ag2S under industrial autoclaving conditions, when the total iodine species to Ag mole ratio exceeded 5. Lower ratios gave proportionally more Ag-jarosite and less AgI. Under these conditions existing Ag-jarosite was also found to decompose in H30-jarosite and AgI.

Jan 1, 1993

By Abdelaziz Baçaoui, Imane Aarab, Abdelrani Yaacoubi, Khalid El Amari, Abderahman Etahiri

Phosphate ores upgraded by flotation are widely used as vital raw materials for the sustainable production of phosphate fertilizers to maintain the food supply. High-grade phosphate reserves become more and more depleted. Low-grade phosphate mining and processing are in the spotlight of the fertilizer industries to cope with the exponential growth in the world’s population and meet the increased demand for fertilizers. As part of a series of research works conducted within a major project aiming at upgrading a low-grade phosphate ore never-exploited before, the present work endeavors to optimize the flotation of this ore using the response surface method (RSM) as a follow-up of a screening study. Unveiling new insights on such a strategy paved the way for improving flotation performances. This work demonstrated that an optimum 30.18% P2O5 with 88.07% recovery was achieved using 63.10 g/t of sodium alginate (NaAL) and 2484.72 g/t of sodium oleate (NaOL) at pH 11.40. This maximum was contrasted with a new formulation of fatty acids (FAM). The 30.76% P2O5 grade with a recovery of 89.72% was attained using 50.94 g/t of NaAL and 1455.40 g/t of FAM at pH 11.68. FAM collector markedly outperformed NaOL at low concentrations proving the efficiency of synergists.

Dec 7, 2022

By Gerald C. Smith

Introduction Newmont Mining Corp.'s Carlin Gold Mining Co. mines gold ore and waste material from four open-pit mines in Northeastern Nevada. Combined production rate is 24.5 Mt/a (27 million stpy). The company is producing gold bullion from the 2.4 kt/d (2600 stpd) Carlin Mill No. 1. and the 2.3 kt/d (2600 stpd) Maggie Creek heap leaching plant. In August 1985, the 6.3 kt/d (7000 stpd) Carlin Mill No. 2 at Gold Quarry came on-line. The total combined throughput of 11.2 kt/d (12,000 stpd) will make Carlin Gold Mining one of the largest US gold producers. Feasibility studies are underway to increase the total gold out-put from Gold Quarry. This will be done by adding a continuous dump-leaching process that will recover gold from lower grade ore. History In 1961, Newmont exploration geologists located claims in the Lynn Creek mining district near Carlin, NV. An exploration drilling program revealed significant gold mineralization. A stepped-up drilling program revealed an ore body with an estimated 10 Mt (11 million st) of ore containing 109 t (3.5 million oz) of gold. Test results in 1963 allowed metallurgists to develop the flowsheet and design for a 1.8-kt/d (2000-stpd) countercurrent decantation (CCD) cyanide plant. Carlin Gold Mining Co. was formed to operate the property. Carlin purchased its first mining equipment in 1965. This consisted of two Bucyrus-Erie 2.7 m3 (3.5 cu yd) shovels and nine 27.2 t (30 st) Wabco Haulpak trucks, two Inger¬soll-Rand track-mounted drills, one Chicago Pneumatic T-650 rotary drill, one rubber tired Michigan 280 dozer, one Caterpillar D-8 36A track dozer, one Caterpillar 16 motor grader, and the necessary ancillary equipment. The fleet was capable of mining 3.2 Mt/a (3.5 mil¬lion stpy) working two eight-hour shifts, five days a week. In 1967, one P&H 1500 3.8 m3(5 cu yd) diesel electrical shovel and three 59 t (65 st) Dart trucks were added to the fleet, along with a Chicago Pneumatic rubber tire mounted rotary drill. This equipment addition increased the mining capabilities to 5.7 Mt/a (6.3 million stpy). Three separate pits, the Main, East, and West, within a 1.6-km (1-mile) radius of the Carlin mill were being mined. The crew consisted of 46 hourly men and three supervisors. This equals 497 t/ total hourly manshift (548 st) or 407 t/total manshift (515 st). Rising gold prices in the early 1970s allowed lower grade deposits to be mined at greater distances from the mill. The Boot-strap pit, 19.3 km (12 miles) north of Carlin, was started in 1973. Two Hough 45.4 t (50 st) front-end loaders, four Wabco Haulpak 45.4 t (50 st) trucks, and a second track type D-8H dozer were added. The higher grade ore was trucked to the Carlin mill using three

Jan 11, 1985

By Jim Richards, Eli Choueiry, Brett Robinson, Amanda Elioff

The Los Angeles Metro Gold Line Eastside Extension (MGLEE) extends Metro’s light rail system to the densely populated and growing East Los Angeles communities. Much of the 6 mile alignment consists of at-grade track, but the project also features a1.7 mile underground segment in cut-and-cover and twin-bored tunnel structures. In light of past experiences with open shield construction, The Metropolitan Transportation Authority (Metro) specified, for the first time, use of Pressure Face Tunnel Boring Machines (TBM’s) to advance tunnels through soft ground. This paper describes the history of the tunnel design, construction management, and construction methods adopted that resulted in successfully completing the tunneling segment.

Jan 1, 2007

By Andreas G. Mueller

Gold-silver and copper-gold deposits with high-temperature wallrock alteration of Ca-Mg-Fe-Al silicates occur in ca. 3.0 Ga old greenstone belts of the Yilgarn Block, Western Australia. The deposits are hosted by amphibolite-facies komatiites, basalts and BIF, controlled by ductile shear zones and located within the contact-metamorphic aureoles of composite granodiorite-granite batholiths. Disseminated sulfides and free gold form foliation-parallel ore zones in a gangue of anhydrous silicates (garnet, clinopyroxene, olivine), hydrous silicates (amphibole, biotite) and calcite. The deposits may be classified on the basis of their dominant economic metals as: 1(1) Gold skarns with substantial silver as by-product, represented by the Marvel Loch mine (production: 2,200,000 t at 3.5 g/t Au), and characterized by the mineral association pyrrhotite, loellingite, arsenopyrite, galena, gold and magnetite. (2) Gold skarns with low silver contents, represented by the Nevoria mine (production:1,850,000 t at 3.4 g/t Au), and characterized by the mineral association pyrrhotite, chalcopyrite, arsenopyrite, gold, Bi-tellurides and magnetite. (3) Copper-gold skarns, represented by the Elverdton mine (production: 850,000 t at .7 5 Cu, 0.81 g/t Au), and characterized by the mineral association magnetite, chalcopyrite, pyrrhotite, pyrite, gold and Ag-Bi tellurides.

Jan 1, 1990

By Hans G. Machel

In low-temperature diagenetic environments (0 c T c 60 - 80°C), elemental sulfur forms by six common processes: via (1) inorganic oxidation of H2S by molecular oxygen, (2) inorganic oxidation of H2S by ferric iron compounds, (3) polysulfide disproportion- ation; and via the microbial metabolic actions of (4) colorless sulfur bacteria, (5) colored sulfur bacteria, and (6) some types of thiobacilli. Economically viable deposits of native sulfur usually are formed by only one process: inorganic oxidation of H2S by molecular oxygen. The amount of elemental sulfur is also limited by three common processes that destroy it in low- temperature diagenetic environments: bacterial reduction, bacterial oxidation, and inorganic oxidation. In high-temperature diagenetic environments (80 - 10o°C c T c 150 - 200°C), particularly in sour gas reservoirs, there are seven common processes forming elemental sulfur: (1) SOdH2S reaction, (2) H2S- hydrocarbon reaction, (3) polysulfide disproportionation, (4) catalytic H2S decomposition, (5) C02-H2S reaction, (6) chlorinification, and (7) exsolution of physically dissolved elemental sulfur. Natural processes that destroy or remove elemental sulfur in such settings are hydrocarbon sulfurization, polysulfide formation, and advection. In most sour gas reservoirs, relatively small quantities of elemental sulfur occur as a free phase (solid or liquid). However, elemental sulfur often precipitates in copious amounts from the reservoir fluids via exsolution and polysulfide disproportionation during production of sour gas reservoirs. By extension, elemental sulfur may precipitate from reservoir fluids in relatively large quantities upon uplift and erosion of deep sour gas reservoirs, or from advection of sulfurous fluids into shallower settings via faults or other high-permeability pathways.

Jan 1, 1992

Mine Terex articulated dump trucks are working on a large restoration project of an abandoned open cast coal site in Blacklaw, central Scotland. The project is part of the construction of a 42-turbine wind farm complex. During the scheduled 18-month duration of the project, three 28-t(31-st) Generation 7 TA30 dump trucks and the six 36.5-t (40-st) TA40articulated trucks are estimated to move about 1Mm3 (1.3 million cu yd)of material. The trucks are working12-hour shifts, 5.5days a week.

Jan 1, 2005

By Dan Wang, Qiuju Li, Xiao Guo

The chemical states and ions interaction were significant interest to understanding columbite mineral/acid solution interface dissolution and reaction. The present research work was focus to explore the specific cations ( Nb5+, Fe2+, Fe3+) distribution near the ferrocolumbite mineral/acid solution interface, and determine these cations leaching sequence in sulfuric acid solution. X-ray photoelectron spectroscopy had the potential to provide detailed dissolution information by resolving chemical state and ion distribution of ferrocolumbite during sulfuric acid leaching process. The binding energy decreased in sequence as Nb5+- O < Fe2+- O < Fe3+- O, and increased in sequence as Fe3+- SO4 < Fe2+- SO4 < Nb5+- SO4 which revealed Nb in ferrocolumbite mineral more easily leached comparing to Fe. The ratio of Fe2+- O/Fe3+-O on leaching residue decreased, and Fe2+- SO4/Fe3+-SO4 ratio increased as sulfuric acid contraction increasing suggested that the Fe2+ oxidized to Fe3+ and released Nb from mineral and dissolved into leaching solution. The ratio of Nb/Fe was lower 0.07 with sulfuric acid contraction increasing which meant Nb dissolution promoted by oxidation of iron. The dissolution mechanism of Fe element in ferrocolumbite mineral was Fe2+- O broken and generated Fe2+. Fe2+ was oxidized to Fe3+ in concentrated sulfuric acid leaching solution. The dissolution mechanism of Nb in ferrocolumbite was that Nb5+- O was directly digested by H+, and the released Nb5+ complex with SO4 2− ion formed cation − anion pairs. The coordination of formation Nb5+- SO4 pairs and oxidation Fe2+ in mineral could promote ferrocolumbite dissolution.

Aug 5, 2023

Remote methane detection State and federal regulations require frequent measurements to verify that safe methane levels are maintained in underground coal mines. Ventilation control is the primary means of maintaining safe methane levels. In the immediate face area, where the highest methane levels occur, methane measurements are required every 20 minutes during active mining. To perform these measurements, temporary roof supports must be erected. This allows a miner to reach the face. There a hand-held methanometer is used to make the measurements. This procedure is time consuming and potentially hazardous. To facilitate the measuring process and reduce the miner&apos;s exposure to roof falls, the US Bureau of Mines has developed a prototype optical device. It can measure methane remotely from distances up to 12 m (40 ft). The device uses two infrared signals, reflected from the coal face, rib, or roof. One signal is absorbed by the methane, the other is not. The ratio of the two signals determines the average methane concentration in the face area. The prototype detector has also been successfully tested at the Lake Lynn, PA Laboratory. Instantaneous responses were obtained for methane in the ventilation airflow and for methane layered near the roof. Conventional methanometers required from five to 25 seconds to respond to the methane in the ventilation airflow. They did not respond to methane in the roof layer. For more information about the Bureau of Mines&apos; remote detection research, contact Richard W. Watson, US Bureau of Mines, Pittsburgh Research Center, Box 18070, Cochrans Mill Road, Pittsburgh, PA 15236 or call (412) 892-6630. Ignition suppression system The incidence has increased of fractional spark ignition at the mine face. One reason is the advent of mechanical mining. Another is the more difficult ventilation conditions associated with deeper and more extensive underground workings. To cope with the problem of frictional ignition, the Bureau of Mines initiated research to develop a system that would provide rapid detection of incipient methane-air ignition. Then, rapid dispersion of a quenching agent would ensure that the methane flame was extinguished while it was still relatively small. This concept was successfully demonstrated by suppressing explosions in the Bureau&apos;s Bruceton, PA experimental mine. But the bulkiness of the early storage-dispersal system proved impractical to retrofit onto face equipment. In a cooperative effort by the Bureau and the Mine Safety and Health Administration, a more compact system has been developed and is being tested. A continuous mining machine has been equipped with four 1.8-m-long (6-ft-long) and two

Jan 9, 1987

By G. M. Wilemon

The U.S. Bureau of Mines has conducted preliminary investigations on the leaching of phosphate ores obtained from deposits in the western United States using sulfuric acid (H2SO4) in the presence of methanol. Phosphate extractions in excess of 80 pct were obtained on an ore sample from Pocatello, ID. Initial leaching studies using the ore as received yielded extraction levels of approximately 35 pct. However, when the sample was ground for as little as 5 min in a rod mill, in methanol, then treated with HzS04, 82 pct extraction was achieved. The R2O3-P2O5 (R2O3 = Al2O3 + Fe2O3) ratio of the starting material was 0.11. Solubilized products after leaching exhibited an R2O3-P2O5 ratio of 0.02, which is well within desired limits. These results illustrate the potential of acid-alcohol leaching for recovering phosphate values from marginal-grade ores and as an alternative to methods involving current beneficiation techniques that yield lower phosphate recoveries.

Jan 1, 1992

By D. J. Simbi, T. R. C. Fernandes, M. Nyamhunga

The recovery of P2O5 from a 90-95% minus 45 microns phosphate slimes sample was carried out in a 5.08-cm internal diameter and 549-cm high flotation column cell. The influence of the feed rate, feed percent (%) solids, and frother dosage on the recovery was investigated following a truncated cubic 3 level by 3 variables experimental design. Two three-variable interactions of feed rate (-l)/feed percent solids (+l)/frother dosage (-1) and feed rate (-l)/feed percent solids (-1)/ frother dosage (+I) gave the region of optimum column performance: recovery (52.60-74.50%), mass yield (15.68-18.80%), and concentrate grade (22.87-31.2196). The collector dosage also showed a significant influence on the column performance. The high reagent consumption relative to the current plant levels points to the need for the removal of ultrafine particles (<10 µm).

Jan 1, 1999

By Gordon Miner

GENERAL DESCRIPTION At the Star mine, Hecla Mining Co., located in Burke, Shoshone County, ID, the mining method is horizontal slicing (timbered or untimbered) cut-and-fill using hydraulic sand backfill. Back stoping as well as breast stoping is used; see Fig. 1. The annual produc¬tion from all cut-and-fill mining is 259 816 t (286,400 st). The average depth ranges from 366 to 2408 m (1200 to 7900 ft). A majority of the tonnage comes from an average depth of 2240 m (7350 ft). Average travel time to the stope is 1 hr. GENERAL ORE BODY REQUIREMENTS AND LIMITATIONS Size and Ground Conditions The ore body is 610 to 1219 m (2000 to 4000 ft) along the strike and 2408 m (7900 ft) on the dip, with an average vein width of 2.6 m (8.5 ft). The dip is essentially vertical. The ore body is heavily sheared and broken on the main fault structures; it is more competent on the wall veins but everywhere subject to rock bursting at depth. The hanging and footwalls are strongly fractured on the main fault structures, again more competent but still fractured on the wall veins. All host rock is Revett quartzite. Value and Production Requirements The ore value is $56 per ton. Production require¬ments are 1089 t/ d (1200 stpd), five days per week. Why Cut-and-Fill? The slusher extraction method is the most satisfac¬tory and practical for narrow vertical veins with sheared walls. The method must be quite selective to be com¬mercial. The veins are too narrow for rubber-tired ex¬traction and the grade is not high enough to make resuing practical. DETAILED DESCRIPTION Development Access to the main shaft is by the 3.2-km (2-mile) main haulage tunnel. The level intervals in the vertical timbered shafts are 61 m (200 ft). Vein development is by crosscut and laterals located 30.5 m (100 ft) into the vein wall. Vein crosscuts are on ±61 m (200 ft) centers. Access to the vein is by raises carried up with the stope. The raises are equipped with four wood¬cribbed chutes (two orepasses reinforced with angle irons, one manway, and one timber slide, Figs. 2-6). The initial cut is located 6.7 m (22 ft) above the main level (Fig. 1). Stoping Cycle Detail Sampling: Grab samples are taken from the mine cars. Channel sampling of the faces and backs is done when necessary. Drilling: For stope drilling, Ingersoll-Rand 300M percussion drills are used, with Ingersoll-Rand 300M and Gardner-Denver RB83M stopers, used for back stoping. Breasting down is done with Gardner-Denver S83FM jacklegs. The back stoping rounds are 0.9 to 1.8 m (3 to 6 ft) and the breast stoping rounds 1.5 to 3m (5 to 10 ft). Blasting: For stope blasting DuPont Tovex 100 watergel with fuses and igniter cord is used primarily. Approximately 60% of the blasting is done with am¬monium nitrate-fuel oil (ANFO). For development headings and shaft sinking, DuPont Drivex watergel with Ensign Bickford Nonel Primadet primers is is used. Mucking: Broken ore is removed entirely by slusher in the stopes. These are mostly 3.7-7.5-kW (5-10-hp) Ingersoll-Rand air-powered or 7.5-37.3-kW (10-50-hp) Ingersoll-Rand electric-powered. In the horizontal de¬velopment headings Eimco 12B and 21 and Atlas Copco LM-56 overshot mucking machines and a 0.76 (1 cu yd) hydroelectric Wagner HST-1 LHD (load-haul-dump) unit are used. Rockway for Extraction: The cribbed chutes (Fig. 7) are nearly vertical and hexagonal in shape and are made of spaced 152 x 203-mm (6 x 8-in.) timber, reinforced with varying weights and sizes of angle iron, depending on wear requirements. Rectangular cribbed chutes are used where ground pressure is not a problem. Cribbing is added in 2.4 or 3-m (8 or 10-ft) extensions accord¬ing to the floor height (whether timbered or untimbered). See Figs. 8 and 9. Backfilling: Backfilling is done with hydraulic sand consisting of classified mill tailings at 65% solids. The fill contains approximately 20% 325-mesh fines. The tailings are pumped 3353 m (11,000 ft) at 50-55% solids 22.7 dry t/h (25 dry stph) with a duplex piston pump. Cycloning increases the tailings density to 65% solids. Fill storage is in a 998-dry-t (1100 dry st) agi

Jan 1, 1982

By A. T. Yu

Two major events highlight recent developments in US ore and coal ports: completion of the last series of modern taconite pellet transshipment facilities on the Great Lakes; and modernization and construction of coal ports, particularly on the East and Gulf Coasts. The New Taconite Transportation System To reduce raw material transportation costs, a fleet of new generation high-capacity 304-m (1,000-ft) self-unloading vessels were built to carry iron ore pellets from the Minnesota-Michigan iron ranges to the steel plants on the lower Great Lakes. Existing port facilities had to be modernized, revised, or completely rebuilt to accommodate these large vessels. Some of these were Burlington Northern&apos;s Allouez, WI, loading dock; Duluth, Missabe & Iron Range Railway Co.&apos;s Two Harbors, MN, facility; Republic Steel&apos;s Lorain, OH, facility; and Chessie&apos;s Toledo, OH, dock. Allouez and Two Harbors receive taconite pellets from unit trains and load them onto large vessels either after dumping or via a large stockpile and reclaim system. The Lorain facility receives iron ore pellets from self-unloading vessels&apos; discharge boom conveyor and reloads the pellets into rail cars or small vessels destined to inland steel mills. The Toledo facility receives Armco pellets from vessels, stockpiles them through the winter, and reloads into unit trains destined for Armco&apos;s mills along Chessie&apos;s rail tracks. Burlington Northern&apos;s $75-million Allouez pellet dock, completed in June 1977, was built to receive pellets produced by Hibbing Taconite Co. and National Pellet Plant in Minnesota, stockpile them through the winter when the lakes are frozen, and load them into 304-m (1,000-ft) vessels in the shipping season. As much as 10 Mt (11 million st) of pellets may be stockpiled within the loop track. A 6-km (4-mile) long conveyor system connects the stockpile area and the dock. Thirty-six new concrete silos were built on the dock to house 2 kt (2,000 st) each of pellets before shiploading. The $35.5-million expansion of the Two Harbors transshipment facilities began shiploading in July 1978 after ground breaking in Aug. 1974. Particularly noteworthy is the first application of the Orboom system-a breakthrough in technology for the modernization of the century-old pocket docks on the Great Lakes to accommodate the new generation of super vessels. The pocket-type loading dock has been a standard on the lakes for nearly a hundred years. Bottom-dump rail cars fill the ore pockets on top of a finger pier. Gravity chutes matching standardized ore ship hatch spacings are lowered to fill the holds of a 20.3-kt (20,000-dwt) ship. The construction is simple and the loading swift. In spite of advances in technology elsewhere, most of these docks continue to serve the iron ore and coal trade in the same manner they did in the 19th century. Although performance of the pocket docks on small vessels remains outstanding, the new 304-m (1,000-ft) vessels are beyond the reach of the old docks. After extensive development, the Orboom system (Patent No. 4,065,002) for pocket docks was successfully developed. The heart of the Orboom system is the retractable shuttle loading arm which loads the wide beam vessels. The Orboom shuttles are fed by existing pockets of the dock that, in turn, are charged by a tripper conveyor along the length of the dock. The Two Harbors shiploading system is supported by a 0.9 Mt (1 million st) storage-reclaim network. Unit trains are bottom dumped. Taconite pellets are stacked and reclaimed by bucket wheel reclaiming systems. Lower Lakes New Ore Ports At the lower Great Lakes receiv-

Jan 10, 1982

By W. C. Labys

This chapter reports on recent modelling effects to analyze the US steam coal market with an econometric as well as a programming approach. The background to coal modelling is examined first. The particular aspects of the steam coal market to be modelled are then presented and the corresponding coal model described. Empirical results are presented and analyzed. Conclusions are then drawn regarding potential modelling applications in the policy area.

Jan 1, 1979

By Raja V. Ramani

The importance of adequate mine ventilation for good health and safety conditions has been recognized since miners first ventured underground. All over the world, governmental actions on mine health and safety are directly connected with mine atmospheric environments, their assessment and control. These governmental actions have been, and continue to be, motivated by the repeated experience that potential threats from inadequate control can evolve, either suddenly or in the long-term, into unpleasant reality, with catastrophic health and safety consequences. These governmental actions are also an indication of the remarkable professional progress being made in the field of mine ventilation. Surely, no developed country allows underground mines to operate without strong measures for the control of ventilation under legislation specific to the mining industry. In the United States, the federal government&apos;s responsibility for mine safety has evolved over the years through a series of legislative actions beginning in 1865 and continuing through 1910, 1941, 1947 and 1952. The passage of the Federal Metal and Nonmetallic Mine Safety Act of 1966 (the 1966 Metal Act), the Federal Coal Mine Health and Safety Act of 1969 (the 1969 Coal Act) and the Federal Mine Safety and Health Amendments Act of 1977 (the 1977 Mine Act) are the most important milestones in this long legislative history. The 1977 Mine Act adopted many of the stronger features of the Occupational Health and Safety Act of 1970, perhaps the most significant piece of legislation on workers&apos; occupational health and safety in US history. It is worthwhile to ask, "How far has the United States come in mine ventilation since the enactment of the 1969 Coal Mine Health and Safety Act?" This Act and its associated regulations defined in great specificity many standards for healthful and safe operation of a mine and, in some cases, what constitutes good ventilation practice. Since 1969, the frequency and severity of disaster-type events in US coal mines has decreased markedly. A primary reason for this achievement is the strength of the Coal Mine Health and Safety Act&apos;s regulatory features in such areas as methane concentration control, dust control, intrinsic safety and explosion-proof enclosures, minimum air quantity and quality standards and air distribution and escapeway provisions. These requirements greatly influenced mine ventilation planning, engineering and practice, leading to increasingly safe working conditions.

Jan 1, 1993

By D. B. Sikka, R. B. Bhappu

Porphyry-type copper-molybdenum (Cu-Mo) deposits typically occur in the Phanerozoic rocks and have been well-documented. During the past two decades, a number of porphyry-type Cu-Mo deposits that range in age from 1055 Ma to 2970 Ma have been identified in the Precambrian rocks of Australia, Canada, China, Finland, India. Namibia and Zambia. The Malanjkhand porphyry copper deposit M.P., India is estimated to contain 789 Mt (870 million st) averaging 0.83% copper (mill headgrade); 0.004% molybdenum; 0.2 g/t (0.006 oz/st) gold; 6 g/t (0.175 oz/st) silver and a cap of 7.5 Mt (8.2 million st) of oxide ore averaging 0.8% copper. The economic potential of the expanded copper mining milling, smelting and refining complex is expected to yield cathode copper 118.6 kt (130,700 st); molybdic oxide 830 t (914 st); gold 1.9 t (61,732 oz); silver 46.6 t (1.5 million oz); sulfuric acid 240 kt (264,500 st), copper from leach dumps 11.5 kt (12,600 st); nickel 24 t (26,400 st); cobalt 84 t (92,600 st); bismuth 16 t (17.6 st); selenium 48 t (53 st), tellurium 4.8 t (5.3 st).

Jan 1, 1995

By Stanley W. Ivosevic, Mark E. Emerson

Introduction Metals exploration activity is moderately intense in the US despite the 1981-82 recession in the mining industry. Although exploration expenditures are significantly less than the 1981 peak, quality and goal orientation of current exploration are high. Major foreign and domestic mining companies as well as nonmetal companies had been attracted to invest in minerals in the US. Its capital markets, political stability, and open-door policy are attractive for hard mineral explo¬ration, particularly in the West. The approach was to either acquire fully or partially developed ore bodies through merger and acquisition, or to fund large scale exploration programs in search of bulk-tonnage deposits. The acquisition avenue has not proved the panacea envisioned for overcoming lack of substantive mineral development. Nor has the decentralized approach to exploration that was adopted proven cost-effective. The emphasis on gold, to the exclusion of all other commodities, raises questions regarding the direction of exploration in the US for the rest of the 1980s. Overview More than 300 companies are active in metals exploration in the US. These range from established, well-known domestic and foreign corporations to hurriedly incorporated, unknown companies financed by having promoted mineral prospects or occurrences of questionable quality. Junior companies tend to be oriented toward property entrepreneurship, rather than the development of minerals resources. These companies have helped maintain mineral lands from prospect to production by holding blocks of property intact during periods of exploration inactivity. This was done through promotional efforts, even though few properties that attracted major venturers proved to have real merit. Exploration objectives of senior and junior exploration companies currently embrace only a few commodities. Arranged in approximate descending order of priority, these are: precious metals; base metals in monometallic or polymetallic deposits; strategic metals, such as tungsten, tantalum, and tin; unconventional commodities, such as diamonds; ferrous metals; and commodities such as beryl, zeolites, and barite. Prospectors are important in searching for these targets. In the US, where they have been financed by major and junior companies, prospectors have proven a useful and low-cost means of locating gold prospects (e.g. the Alligator Ridge property in Nevada). The tenacity and persistence of independent prospectors have been factors in discovering properties rejected or overlooked by professional geologists (e.g. Lornex mine in British Columbia). This has been realized by the provincial government that grubstakes bona fide prospectors. Exploration interest in commodities has always been cyclic with price. Gold is one of the few metals that has experienced a price increase of such magnitude to warrant attention. The potential for realizing adequate operating margins for the Nevada open-pit "no-see-um" gold deposits is great. The present declining gold price, however, may indicate that this emphasis on gold, to the exclusion of other commodities, is shortsighted. Mining companies have been the last to react to price increases, opening new mines based on unrealistic market expectations. A scenario could develop where plummeting gold prices bring recession to the gold industry, as lately occurred in uranium, molybdenum, and copper. US Exploration Climate The US is attractive for exploration due to its strong mineral po-

Jan 4, 1984

By C. Alexander

Stillwater Mining Company is developing a second underground palladium and platinum mine approximately 21 km (13 miles) west of their existing Nye operation in south central Montana. A 5650 meter (18,550-foot) tunnel is required to gain access to the ore bearing horizon (J-M Reef). Access and primary mine development can be accelerated through the use of two 4.6 meter (15-foot) diameter tunnel boring machines (TBM). One TBM was purchased new while the other was purchased used. Machine specifications, site support facilities, portal/launch chamber construction, TBM assembly, startup and operation are discussed in this paper.

Jan 1, 1999

By C. Alexander

Stillwater Mining Company is developing a second underground palladium and platinum mine approximately 21km (13 miles) west of their existing Nye operation in south central Montana. A 5650 meter (18,550-foot) tunnel is required to gain access to the ore bearing horizon (J-M Reef). Access and primary mine development can be accelerated through the use of two 4.6 meter (15-foot) diameter tunnelboring machines (TBM). One TBM was purchased new while the other was purchased used. Machine specifications, site support facilities, portal/launch chamber construction, TBM assembly, startup and operation are discussed in this paper.

Jan 1, 1999

By Robert M. Stevens

INTRODUCTION Truck haulage of muck in underground mining operations has been in use around the world for many years. From the late 1.950s into the early 1970s, haulage by means of load-haul-dump (LHD) units challenged haulage by means of trucks, and, to a large degree, LHD units dominated the trackless concept of moving mate¬rial underground. However, beginning in 1975, the emphasis began to shift back to the use of trucks, with LHD units assuming the secondary role of loading the trucks. Over the years, several manufacturers have been offering trucks with capacities ranging from 4.5 to 45.4 t (5 to 50 st) and with features that specifically address the constraints of underground operation. TRUCK TYPES Three different types of trucks are available for use underground, namely the tip dumpers, telescoping dumpers, and push-plate dumpers. The following para¬graphs describe each of these three types. Tip Dumpers Tip dumpers are designed to lift the truck box, using gravity to dump the load out the rear of the box. The primary application of these trucks is in the haulage of muck to the surface. However, these trucks also may be utilized underground, if a chamber is provided to accommodate the raised truck box at the dumping point. Fig. 1 illustrates a four-wheel drive truck with a 10.0-t (11-st) capacity; this truck is only 1.9 m (6.25 ft) wide and is designed specifically to follow narrow and winding vein courses. Fig. 2 illustrates a two-wheel drive tip dumper with a 24.5-t (27-st) capacity; this truck operates on 12% grades with firm dry surfaces and in modes both loaded-up and loaded¬down. Telescoping Dumpers Telescoping dumpers specifically address the constraints imposed by low back heights, and they represent an optimization of compressed envelope size with maximum payload capacity. Figs. 3 and 4 illus¬trate a typical telescoping dumper with a capacity of 18.1 t (20 st). Fig. 3 shows the box in position for carrying the load, and Fig. 4 shows the same truck with the box in the full-dumping position. Fig. 5 illustrates the loading and dumping cycles. Fig. 6 illustrates an approximation of how an LHD unit with an eject-o-dump (EOD) bucket would be used to load over the rear of this type of truck. Fig. 6 shows a low-profile telescoping truck in the first stage of loading, with a 6.8-t (7.5-st) LHD unit equipped with an EOD bucket. A loader with a smaller capacity and a narrower bucket would allow the bucket to fit inside the truck box and would achieve an even lower loading height than that illustrated. Gen erally, an LHD unit equipped with an EOD bucket and teamed with a telescoping dump truck can be used with a lower back height than could any other front-end loader and truck team. Push-Plate Dumpers Push-plate dumpers are similar to telescoping Bump¬ers, except that they do not have the two-stage box. Instead, push-plate dumpers utilize a single push plate to discharge the load out the rear of the box. Gen erally, these units are limited to payload capacities of about 9 t (10 st). Fig. 7 illustrates a typical push-plate dumper.

Jan 1, 1982