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By Gary B. Glass

Seventeen percent of US coal production now comes from the 19 surface coal mines in the Powder River Basin of northeastern Wyoming. The Powder River Coalfield, which coincides with this topographic and structural basin, is unique to the conterminous US, in that it contains an estimated 934.4 Gt (1.03 trillion st) of in-place coal resources. At least 13.2 Gt (14.6 billion st) is a compliant, strippable reserve base. In 1993, an estimated 162 Mt (179 million st), or 98% of the coal mined in this Wyoming coalfield, came from the thick - average 30.5 m (100 ft) - subbituminous, Wyodak coalbed, or from the Anderson and Canyon coals in the one mine where the Wyodak is split into two separate coalbeds. The other 2.7 Mt (3 million st) of coal mined in the Powder River Basin in 1993 came from the School, Badger, Monarch and Deitz No. 3 coalbeds. In this "world class" coalfield, there are at least 13 other persistent, 3 to 15 m (10 to 50 ft) thick, coalbeds as well as the Big George and Lake De Smet coalbeds, which are each 61 m (200 ft) or more thick in places. None of these latter 13 coals are mined at this time.

Jan 1, 1994

By R. Church

The southern Powder River Basin (PRB) of Wyoming has developed into the largest coal producing region in the United States in the relatively short time span of 22 years. Since the opening of the first non-captive mine in the Basin in October 1972, truck/shovel fleets have played a major part in this rapid growth. In 1994, more than 200 trucks and 50 shovels will be used at 14 mines to help produce more than 190 Mt (210 million st) of coal in the southern PRB. This article presents an overview of the truck/ shovel operations in the Basin. Fleet sizes are reviewed, applications discussed, operating considerations explored, factors looked at that drive high capacity and restrictions listed for future growth. Fleet size Various sizes of trucks and shovels can be found in the southern PRB. However, a clear trend of increasingly larger size units is being established as producers attempt to remain competitive and productive.

Jan 1, 1994

By G. R. Baiden

This paper discusses the application of Load-Haul-Dump (LHD) machine teleoperation in combination with a guidance system for single operator control of multiple IJIDs. A system accomplishing these goals has been tested at Copper Cliff North Mine of Inco Limited. This development uses a high-speed, high-capacity production network to teleoperate an LHD during loading and dumping by using 3D video and radio remote control. Following loading, the operator dispatches the LHD using an automatic guidance system to traverse the drifts to the orepass. Remote dumping is done at the ore pass and then the machine returns to repeat the procedure. This has significant productivity gains and improvements in safety. Productivity is improved because operators are involved only in the loading and dumping part of the process. This frees the operators to utilize their time more effectively, loading or dumping a particular machine, while other machines with guidance systems continue to work travelling and/or automatically dumping ore in the ore passes. Safety is improved as the operator is located on a remote control station. Automatic dumping will be incorporated as a future development.

Jan 1, 1994

By D. D. Harbuck

Previous studies at the US Bureau of Mines (USBM) have shown that only 70% to 80% of the germanium in domestic zinc process residues can be extracted during H2S04 leaching. To increase germanium extraction, the USBM developed and researched the hypothesis that minor amounts of germanium and zinc silicate, present in zinc calcine, dissolve during the leaching stages of zinc processing, forming soluble germanium and silica gel. This silica gel then selectively adsorbs germanium from solution, rendering the germanium insoluble. Testing showed that as the amount of formed silica gel increased, the amount of insoluble germanium in the zinc residue increased. Several processing alternatives are presented for overcoming this problem.

Jan 1, 1994

By Dan Plazak

The authors are to be congratulated for presenting a simple and practical tool for quick estimates of ground water flow rates. At the risk of detracting from the simplicity of the method, I would like to suggest a slight refinement to the equation as presented. As the authors noted, the Jacob-Lohman equation (1952) assumes confined-aquifer conditions. To use confined-aquifer equations in water-table aquifers, Jacob (1963) presented a correction to convert water-table aquifer drawdowns [ ] to equivalent confinedaquifer drawdowns [ ]:

Jan 1, 1994

By Gerhard Kircbner, Boen Tan

"IntroductionThe Key Lake uranium-nickel ore deposits are located in the Athabasca Basin of northern Saskatchewan, Canada (Fig. 1). Many papers have been published on Key Lake in the past (Dahlkarnp and Tan, 1977; Kirchner and Tan, 1977; Kirchner et al., 1980). Immediately after its incorporation in 1968, Uranerzbergbau GmbH, a German holding company, set up exploration subsidiaries in Australia, the United States and Canada. This account reviews those exciting years starting with the concept to explore the rim of the Athabasca Basin in Saskatchewan, and leading up to the discovery of the deposits in 1975 and 1976.A few summary statements can be made:• The discovery of Key Lake resulted from a very systematic exploration approach and the persistence of Uranerz, the operator. • Cluff Lake and Rabbit Lake had been discovered some seven years before, but it was the discovery of the Key Lake deposits which brought Saskatchewan to the forefront to become Canada's foremost uranium producing province.• The experiences gained at Key Lake, which were subsequently published, triggered the discovery of many other deposits of similar, or even larger, size than Key Lake. The spatial relationship of those deposits to graphitic conductors and hydrothermal alteration zones are presently the most effective indicators for locating new deposits - indicators which were first recognized at Key Lake.• Key Lake is, and will remain for some years to come, the largest producing high-grade uranium mine in the world."

Jan 1, 1994

By G. A. Brooks

The problem of treating concentrates with increasingly higher levels of arsenic and antimony is compounded by the trend to stricter environmental guidelines regarding the disposal of smelter wastes. Much of the arsenic and antimony from non-ferrous processes is recovered as oxide fume. The options for handling these fumes include recycling them to the smelter, stabilising and disposing of them, and recovering values from them. Whilst arsenic oxide has little value, there is a market for antimony oxide and this provides an incentive for devising techniques for separating antimony oxide from arsenic oxide before disposal of the arsenic oxide. Arsenic can be separated from antimony oxide by selectively volatilising the more volatile arsenic trioxide from a mixed oxide sample or by selectively condensing the less volatile antimony tetroxide at high temperatures, leaving arsenic trioxide to condense out at lower temperatures. The results of selective volatilisation experiments at 379 to 587 °C indicate that it is difficult to achieve an arsenic content below 5.0 wt.% using air or nitrogen as carrier gas due to the formation of solid solutions between the two oxides. Injection of oxygen into a mixed vapor stream improved separation by selective condensation and antimony tetroxide containing as low as 0.23 wt.% was obtained in laboratory experiments.

Jan 1, 1994

By Allan J. Breitenbach

The heap leaching process has developed into a low-cost method for recovering copper, gold and silver from large and small deposits of low-grade ore. The earliest known heap leaching process with chemicals occurred in about 1752 and is attributed to Spanish miners percolating acid solutions through large heaps of oxide copper ore on the banks of the Rio Tinto (Ran Zyl, et al., 1988). In 1969, US Bureau of Mines research¬ers proposed the use of dilute alkaline cyanide solutions as a low-cost means for recovering gold values from low-grade ore containing micrometer-sized particles of gold (Miners News, 1991). The modern-day heap leaching process typically involves the percolation of leach solutions through controlled lifts of stacked, low-grade ore (crushed, aggomerated or run-of-mine) in a heap for chemical extraction of precious and base metals into solution (Fig. 1). The enriched (pregnant) solution is collected by gravity flow in lined basins, sumps and ponds for processing and recovery of the target metals (copper, gold and silver). The barren solution from the recovery facilities is recirculated to the heap, along with chemical reagents and makeup water, for containment of all process solutions in an enclosed and environmentally sound system. This involves the use of low-permeability liner systems beneath the heap leach stacks to maximize solution recovery and protect the underlying ground water (Fig. 1). This article discusses the typical heap leach liner systems currently used by mining companies in North America. The most successful heap leach pad liner systems from past precious metal operations generally include an underlying, low-permeability soil bedding material; a geomembrane liner, and an overlying protective cover fill.

Jan 1, 1994

By Edward DeLilla

For the past 30 years, surface mines have moved toward more highly mechanized, automatic systems that continuously mine overburden and ore to increase productivity, lower operating costs and minimize manpower needs. To achieve these goals, mine operators must be involved in every step of the machine procurement process. Several factors are involved in making the machines operate economically and efficiently: -Prudent engineering. -Proper planning. -Realistic feasibility studies. -Effective specification. -Thorough training of the mine personnel.

Jan 1, 1994

By Derek J. Fray

The need for on-line sensors is outlined and it is demonstrated that the control of many processes is prevented by the availability of these devices. The use of solid electrolytes is described for the sensing of elements in molten pyrometallurgical phases. In the cases where there is no suitable electrolyte, the addition of a second phase of a compound containing both the element to be measured and the element whose ion is mobile in the electrolyte can be used. For the measurement of high concentrations of an element a method based on current measurement is proposed.

Jan 1, 1994

During 1993, contraction of indus¬trial economies in Germany (-2%), France (-1 %), Japan (-0.1 %), Eastern Europe (-2%) and the former Soviet Union (-13%) dimmed prospects of a robust economic recovery for North America. Some European countries that suffer from high unemployment and large government deficits have non-com¬petitive cost structures that will require long-term economic reforms to solve. Restructuring in the United States also continued in 1993. The sluggish pace of the US economic recovery, in¬troduction of federal tax increases and the persistence of high levels of unemployment did little to inspire consumer confidence. By the end of the year, however, sales of consumer durable and capital equipment began to grow. Depressed demand for minerals by industrial economies in recession was partially offset by continuing rational¬ization or abandonment of non-competi¬tive productive capacity in Eastern Eu¬rope and the former Soviet Union. Al¬though aluminum, lead, titanium sponge and zinc prices fell, increases in copper and steel consumption helped steady their prices along with prices of some alloys. Cash flow to the mining industry - the primary source of exploration funds - was therefore constrained. For many companies, exploration expendi¬tures were reallocated geographically and in terms of commodities sought.

Jan 1, 1994

By William H. Pomroy, Kenneth E. Hjelmstad

An underground mine fire can seriously threaten the lives of all personnel in the mine. This fact makes fire emergency planning a high priority in the mining industry. A fire emergency plan includes a means of warning miners of the existence of a fire, and the need to evacuate or retreat to a place of refuge. The warning alarm must be fast and must reach all miners, regardless of their location or work activity. Means of fire warning are varied, and include such methods as stench, flashing lights, telephones, and word-of-mouth (messengers). Unfortunately, none of these methods are entirely satisfactory. Stench can be slow and seldom achieves total mine coverage, especially in high-back room-and-pillar mines with low ventilation velocity. Telephones are effective only where personnel are close enough to hear the phone ring. A novel fire warning alarm system has been developed by the U.S. Bureau of Mines which enables instantaneous warning of all underground personnel. The warning alarm is sent by an electromagnetic signal which penetrates through mine rock from a transmitter to small person-wearable receivers. Each miner's cap lamp battery could be equipped with a small receiver and when a fire or other emergency occurs the miner could be alerted to the danger by a blinking of the cap lamp. Testing of prototype equipment has demonstrated that the transmitted signal is capable of penetrating through over a mile of mine rock and activating the receiver. This paper describes the theoretical basis for through-the-rock ultra-low frequency electromagnetic transmission, design of the prototype transmitter and receiver, test results, and status of commercial development of systems utilizing this technology.

Jan 1, 1994

By W. Hoffelner

Plasma technology has become increasingly important for the treatment of several types of waste (toxic, nuclear, metal-bearing scrap). Plasma furnaces combine the advantages of a closed hearth process with a very versatile (several plasma gases can be employed) and effective heat source. It can be successfully used for recovery of metals with subsequent purification and/or vitrification of residuals from drosses or filter dusts.

Jan 1, 1994

By R. K. Devala

In this paper, numerical approximation schemes, for solution of the mass transfer problem in mine ventilation have been developed by using three implicit finite difference schemes. Results from these schemes have been compared with the analytical solution for spatial and time increments, using the concept of Root Mean Square error. For implicit finite difference schemes, in general, the modified central difference scheme is better than the two other schemes at lower velocity. Also, at lower air velocities, implicit schemes underpredicted the concentration values. At higher air velocities, the Crank-Nicholson and Laasonen schemes provided better results than the Modified Central Difference scheme.

Jan 1, 1994

By C. Bourgoin

"The factors constraining productivity in the mining of narrow vein ore deposits are numerous and often interdependent. This paper concerns mining techniques capable of reducing some of these constraints in the case of vein deposits whose dip is less than 45 degrees. Modifications of conventional mining methods as well as suggested new developments are presented, as are matters pertaining to the mechanization of operations as a mean of increasing both productivity and work safety. IntroductionThe mining of vein deposits, whose width (thickness) is generally less than 2 metres and whose dip is often variable, calls for the use of varying stoping procedures as well as appropriate infrastructure and equipment. In the case of intermediate dips (15 degrees to 45 degrees), it is generally conceded that the degree of mechanization is limited and that ore removal via gravity is impossible. This paper proposes certain innovations in matters of equipment use and stoping methods, together with certain technical data pertaining to the application of these new procedures Stoping MethodsTo mine vein deposits, one must choose a method that can adapt to irregular ore limits and allow, as well, for good ore recovery with as little dilution as possible. In addition, it is important to minimize development work in waste rock. In principle, one selects a method that is both safe and profitable; this method will be defined mainly by the nature of the vein deposit and its dip, by the competency of the rock formations and by certain technical considerations. Moreover, the stoping procedure in a given deposit may well change with time as a function of new layouts or new equipment."

Jan 1, 1994

By Morris J. Viljoen, David R. Young

"The zone of economic interest at H.J. Joel Gold Mine is the VS5/Beatrix Composite Reef, a quartz-pebble conglomerate of Archean age. The reef occurs at the base of the Eldorado Formation, Central Rand Group, of the Witwatersrand Supergroup. At least three stratigraphically distinct conglomerates form this economic zone. In order of deposition these are the Footwall, Beatrix, and the VS5/Beatrix reefs. The Footwall and Beatrix reefs are equated with the Aandenk Formation with transport directions toward the northwest, whereas the VS5lBeatrix Reef is of Eldorado age, with transport directions toward the south. The change in the drainage direction was caused by a major basin-wide tectonic event that occurred prior to the deposition of the Eldorado Formation. The Beatrix Reef was formed in a proximal, braided, fluvial environment by a process of aggrading gravel bars between channelways. The VS5/Beatrix Reef was formed by the desegregation of pre-existing Beatrix Reef with the introduction of new material containing much shale detritus and is seen to have been deposited in a more -distal, braided, fluvial environment, similar to the Beatrix Reef.The average gold concentrations are different for each conglomerate; the Beatrix at 32.1 ppm Au, the Footwall at 24.4 ppm Au, and the VS5/Beatrix at only 12.4 ppm Au. A sedimentary gold concentrating process related to higher energy braid channels was primarily responsible for these high gold grades. Not only the basal unconformity surface, but also, internal degredation surfaces, were prime sites for concentration. An understanding of the spatial distribution of these differing conglomerates and their braid channels forming the economic reef zone is necessary in predicting ore grades for future mining."

Jan 1, 1994

By Bror Nyrnan

Solvent extraction technology and equipment have been developed within Outokumpu since the early seventies. During this period altogether nine solvent extraction processes for molybdenium, zinc, rhenium, cobalt, nickel, calcium and magnesium have been commercially adopted in Outokumpu's metallurgical plants. During last years a new mixer-settler concept based on a "Vertical Smooth Flow" (VSF) principle has been developed. It has also been found to be a good method for copper solvent extraction, decreasing essentially the crud formation and the organic and aqueous entrainments.

Jan 1, 1994

By Paul D. Tomlingson

Editor's note: This is the first of two articles on maintenance. Part two, "Effective maintenance downsizing techniques," will appear in the October issue of MINING ENGINEERING. Most mine managers are aware that the cost of maintenance rarely goes down unless work is done less often and with fewer people. Fewer people translates into downsizing and seldom means making maintenance more efficient. Even so, mines that have struggled unsuccessfully to improve maintenance and are now pressed to reduce maintenance costs may opt for downsizing. On the other hand, mines with an immediate need to reduce costs may be forced to downsize because they cannot wait for needed improvements to take effect. This can be bad news for maintenance. In any case, the mine must downsize correctly to ensure effective maintenance is sustained.

Jan 1, 1994

The U.S. Bureau of Mines (USBM) has developed the coal Mine Roof Rating (CMRR) to aid in ground control. This paper describes the application of the CMRR at four coal mines operated by a major U.S. coal company. The mines are widely separated geographically, and cover a wide variety of geologic conditions and mining methods. The study found that the CMRR was a useful predictor of roof performance. In mines where the CMRR varied, and other parameters held constant, roof falls were associated with low CMRR values. Important similarities in roof performance were found in mines with similar CMRR values, despite radically different geologic environments. The analyses also indicated that statistics can be used to evaluate ground control performance. For these mines, a "roof fall probability" can be estimated from the CMRR and the intersection span. A spreadsheet program has been developed to facilitate treatment of CMRR data, and it is briefly describe

Jan 1, 1994

By Dean H. Herrick

The production of crushed stone in the United States has exceeded 1.0 Gtpy since 1987. Eighty percent of the crushed stone produced is used in the aggregate industry, accounting for over half of the quantity of natural aggregate consumed in the United States. Approximately 718 of the crushed stone produced in 1989 was limestone and dolomite followed by granite, traprock, sandstone and quartzite, miscellaneous stone, marble, shell, calcareous marl, volcanic cinder and scoria, and slate in order of descending volume (Table 1). Of the remaining 20% crushed stone produced in the United States and not used as aggregate, 12% is used in the manufacture of cement which is an important construction material. Only 8% of crushed stone is used in specialty products not associated with the construction industry. Therefore, this chapter will concentrate on crushed stone as used in aggregates, with occasional reference to other uses or to chapters which discuss specific nonconstruction uses of crushed stone. Construction aggregates are hard, essentially inert materials suitable for being formed into a stable mass by the addition of cementing or binding materials that produce a concrete by compaction, or by natural weight to produce a road or foundation fill. Aggregates are usually from 80 to 100% of the material volume of the construction masses in which they are used. Conditions necessary for a rock deposit to be developed for construction crushed stone include: Quality-should pass rigid specifications for strength and durability; Quantity-adequate volume of rock is needed to have a production life of at least 10 years, preferably 20 or more years; Market-must have adequate market to support a new crushed stone operation; Transportation-transportation cost must be competitive for the intended market; Environmental-the impacts of mining, processing and transportation of crushed stone on the environment at or surrounding the mine site must be within acceptable limits; Permitability-mining, processing and transportation of crushed stone must usually be permitted by governing agency. The following discussion will touch upon all facets of crushed stone operations from the mineral deposit and its geology to technology, uses, economic factors and end with governmental regulations of mining crushed stone. The sections are not designed to answer all questions, but rather to be an introduction to the crushed stone industry. Where more information is required, The Aggregate Handbook by Barksdale (1991) is recommended as the best general source.

Jan 1, 1994