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By Jon C. Volkwein

The U. S. Bureau of Mines investigated a method of using inert gas to prevent the formation of explosive gas mixtures in auger highwall mining of coal. A combination of gasoline and diesel engine exhaust gases was introduced into the auger drill hole using a short section of pipe located at the collar. Gas samples were taken and analyzed on site with infrared detectors for oxygen, carbon dioxide, methane, and carbon monoxide. Evacuated bottle samples were also taken and analyzed by gas chromatography at the Pittsburgh Research Center. These gas results were analyzed for explosibility. Personal exposure to carbon monoxide was also monitored. The highest methane level observed was 9.55 pct. The Inert gas levels, (carbon dioxide and nitrogen) were sufficiently high to prevent any ignition of the methane. Results showed that for all conditions during mining, gas concentrations were non-explosive. The maximum personal time weighted average sample for carbon monoxide was 20 ppm. This system provides a safe, inexpensive, simple method for preventing explosions during auger mining. INTRODUCTION The auger highwall mining method is an effective method to recover coal from a reserve when removal of the overburden by surface mining equipment becomes uneconomical. In this method of mining, a horizontal auger enters the coal seam from the surface mine bench under the highwall and the coal is drilled in a series of parallel holes. Historically, coal mined from the surface is relatively shallow, and over time, methane associated with the coal has dissipated through the surface. In most circumstances, little methane has been found associated with auger mining. However, mining technology has enabled surface mining of deeper reserves of coal. Furthermore, environmental constraints have forced the highwall extraction method to be used to remove coal under wetlands, further increasing the chances of encountering methane. Recently incidents of methane explosions at a few auger mining operations have resulted in injuries and increased testing for methane at the collars of auger holes. The fuel source of the reported explosions was not necessarily limited to methane, but may also have involved coal dust. The Mine Safety and Health Administration (MSHA) met with the Bureau to discuss what technology might be available to enable the safe resumption of mining. The discussion included the difficulty of ventilating through the solid shafts of the augers, that steel bits probably created the ignition source, and that perhaps inerting the holes with low oxygen and high carbon dioxide concentrations from the machine's diesel exhaust was a potential solution. Considering the ventilation aspects of the problem, it was not clear If ventilation could be reliably established. If some degree of ventilation to the front of the mining head is achieved, it may combine with methane to bring the hole atmosphere from a rich, nonexplosive mixture to an explosive mixture. Furthermore, it may not prevent a dust explosion in such a mining configuration. Lack of access through the shafts of auger type mining machines further limits the ability to add water or air to cool bits to prevent an ignition source from developing. Either of these approaches would also be expensive. The process of mining coal In an inert atmosphere has been considered in the past, but to our knowledge, never implemented (Department of Interior, 1970). Clearly, implementation in underground mining would be more complicated. On a mine bench open to the atmosphere, however, adding inert gas to the mining head could provide a quick, feasible method to prevent explosions at auger highwall mining operations. Also the problem of how to move the inert gas to the cutting head of the machine had to be considered. Preventing explosions on auger mining machines using inert gas requires three primary considerations: first is the source of inert gas; second, placing the inert gas at the cutter head; and third, monitoring the hole atmosphere. Any gas source having an effective inert gas concentration of 34 volume pct or greater will prevent methane from Igniting (Zabetakis, 1965). Sources of inert gas considered for this application included liquid nitrogen, modified shipboard inert gas generators (for hydrocarbon shipping and transfer), jet turbine engine (Paczkowski, et. al., 1982), the auger's diesel engine and a gasoline engine. Operation cost, purchase cost and availability limited our testing to the diesel and gasoline engines. This work tested each engine, separately and combined. To ensure effectiveness, both company and enforcement personnel need to know how to monitor the condition of the inerted hole. Measurements at depth inside the hole are possible by remote sampling through rigid tubing, but this method is Impractical for routine monitoring. Continuous monitoring of the exhaust gas stream is an alternative. The U. S. Bureau of Mines evaluated an inert gas system at an auger mining operation at a surface mine near Owensboro, KY. Coal was mined from the Number 9 Coalbed in Henderson Co. KY. Tests were conducted in January and March of 1992.

Jan 1, 1993

By Keith S. Olsen

Any discussion of barite mining is virtually impossible without considering, almost in the same breath, the many other variables, such as beneficiation, transportation, infrastructure and location that impact the economics of a particular barite orebody. My comments will be limited to just the mining aspects of barite orebodies and these observations must be considered in proper context with all other parameters affecting exploitation of a barite property prior to final decision on economic feasibility. In many instances low-cost mining situations are more than offset by high-cost beneficiation requirements. Likewise, there are cases of relatively high-cost mining associated with no or low-cost beneficiation requirements thus permitting economic exploitation. Barite orebodies and occurrences are almost as varied as the number of occurrences. As in most mining situations, the design of a mine plan and implementation of a proper exploitation program must be approached on a case by case basis. Proper and sufficient geologic control of a barite orebody must precede the development of a mine plan. Generally speaking, barite orebodies can and, at many times, do "strange" things; pinch with depth; almost never increase in size with depth, overturn, fault-off, change grade, etc. Due to the extreme variable nature of most barite orebodies, a mine plan and its implementation must be as flexible as possible. The ability to quickly adapt to variable conditions will greatly enhance the economics within a particular barite mining environment.

Jan 1, 1980

By Wolfgang R. Wawersik, Charles M. Stone

INTRODUCTION Hydraulic fracturing was applied in horizontal drill- holes in the Salado salt formation near Carlsbad, New Mexico. Tests were performed approximately 650 m be- low surface to validate the design of a Waste Isolation Pilot Plant (WIPP) for the disposal of radioactive waste from defense activities of the United States. Hydraulic fracturing was performed primarily to determine whether the virgin in situ stress state at the WIPP site is isotropic and whether the magnitude of the virgin in situ stresses corresponds to the weight of the overburden. Beyond these limited objectives, however, measurements are now being reviewed to evaluate the usefulness of hydraulic fracturing in salt formations in general. Such measurements are desirable to determine stresses induced by mining and to monitor time-dependent stress changes around underground excavations in salt masses. Hydraulic fracturing was chosen because it is the only technique that can be readily applied in virgin ground far from existing mine openings. The method also appeared to yield directly the magnitude of the least compressive principal stress, regardless of the mechanical properties of the rock mass. A knowledge of the least principal compression was deemed sufficient to establish the isotropic character of the virgin in situ stress field under two conditions: (1) the Salado formation was mechanically isotropic, i.e., the fracture toughness did not vary normal and parallel to bedding so as to permit preferred fracture paths, and (2) pressure measurements were combined with determinations of the orientations of hydraulic fractures and, more generally, with observations of the characteristic hydraulic fracture patterns. The latter was possible by locating the drillholes for hydraulic fracturing along the axes of drifts that were mined several months later. The success of general stress measurements by hy- draulic fracturing of salt subjected to anisotropic stress conditions was and remains less certain. While the smallest in situ compressive stresses should still be equal to instantaneous shut-in pressures, there may not exist any unique relationships between the greatest and intermediate in situ principal stresses on one hand and the primary breakdown and fracture reopening pressures on the other. This problem would be a consequence of stress relaxation due to salt creep around any drillhole in which hydraulic fracturing tests are conducted. Moreover, lab- oratory measurements suggest that creep and stress relaxation could be so rapid that it might not be possible to interpret the measurements for anisotropic stress conditions even if hydraulic fracturing were performed as soon after drilling as possible. The primary objectives of the stress measurements in the WIPP were met by combining hydraulic fracturing tests with finite element analyses of the effects of salt creep, observations of hydraulically induced fracture pat- terns, and laboratory hydraulic fracturing tests on oriented core. In situ hydraulic fracturing tests in long drillholes were preceded by several trial tests to establish the best test methods and to demonstrate that salt could be fractured. The resolution of anisotropic stresses and special problems, e.g., in the identification of instantaneous shut-in pressures, are currently being addressed in hydraulic fracturing tests in a mine pillar, in tests following different delays between drilling and fracturing, and in tests with different frac fluids. This paper will review the main results of the analyses above and of one group of hydraulic fracturing tests at different distances from existing excavations. The paper will conclude with a brief description of the geometry of the hydraulically induced fracture patterns. Because of space limitations, it will be assumed that the reader is familiar with hydraulic fracturing or will first consult other sources (e.g., Cleary, 1980; Gronseth and Detournay, 1979; Zoback and Haimson, 1983; Nemat-Nasser et al., 1983; Kim et al., 1984)

Jan 1, 1986

By Larry L. Snyder

A compact 4.87 m (16 ft) diameter tunnel boring machine (TBM) is capable of boring a 12.2 m (40 ft) radius curve while gripping horizontally, or vertically, as may be required for the excavation of turnouts. The machine's 5.0 m (16.5 ft) length and a patented machine structure allows steering while boring without causing harmful eccentric loads. An extremely stiff thrust system, rigid front and rear grippers, and the torsional stiffness of a front end electric cutterhead drive combine to create superior machine stability and boring efficiency. A 2.2 m (7.2 ft) diameter prototype, complete with a soft ground shield attachment, was used successfully on a tunnel project in Scotland.

Jan 1, 1989

By Rod A. Levander, Kenneth B. King

INTRODUCTION The Panna Maria Uranium Project is located in South Texas, between San Antonio and Corpus Cristi, approximately 10 kilometers (6 miles) north of Karnes City, Texas. The tailings disposal facilities are located immediately south of the mill area and the central portion of the west pit. The mine and mill drainage pond is located north of the tailings pond. Location of the facilities is shown in Figure 1. The tailings pond is a disposal area for mill process wastes and has a capacity for storage of approximately 9 million metric tons (10 million tons) of tailings. This capacity is adequate for approximately 10 years of milling at a maximum rate of 2,700 metric tons (3,000 tons) per day. The tailings disposal facilities were designed to essentially eliminate discharge of contaminated waters to preclude degradation of surface streams or ground water. Water levels in the ponds will be maintained. through reuse in the mill and by evaporation. WASTE CHARACTERISTICS The sources of waste that is stored in the tailings pond include runoff from the mill area and tailings from the mill processing. The tailings consist of a sand slime slurry; the percentage passing the No. 200 sieve varies depending on the ore material, and it is generally in the range of 10 to 25 percent. The tailings have a density of approximately 1.2 metric tons per cu m (75 lb per cu ft) and an average specific gravity of 2.55. The liquid effluent from the mill consists of waste solutions from the grinding, leaching, extraction, and washing circuits of the mill. These solutions have an initial pH of 1.5 to 2.0 and contain the remnants of the leaching agents used in the mill process, sulfates, some silica as the primary dissolved solid, and trace quantities of soluble metals. This liquid is discharged with the solids into the tailings pond and is diluted by rainfall and runoff from the tailings dam. Runoff from the mill area, drainage from the mine area, and de-oiled water from truck washing facilities is discharged into the mine and mill drainage pond. Runoff from the mill area drains approximately 28 hectares (70 acres) from offices, shops, the mill site, the ore pad, haul roads, and adjacent areas. Mine drainage is collected in the mine area and pumped to the mine and mill drainage pond.

Jan 1, 1980

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

6.1 INTRODUCTION A surface structure will suffer damage when the additional stresses induced by ground deformations associated with surface subsidence, plus the original stress introduced by construction de¬sign, exceed the strength of the structural elements. In this con¬text, there are two methods available for preventing and control¬ling surface structural damage: one is to strengthen the structure and the other is to design the mining operations such that ground deformations at the structure site can be reduced to an acceptable level. Mining operations include panel layout and mining tech¬niques. These methods are detailed in this chapter. It must be noted that most prevention/control methods men¬tioned in this chapter are used in the countries where the reference papers are cited. In the United States, the coal operators are not required to take those measures mentioned in this chapter. Some of the methods described in this chapter cannot be implemented with¬out changes in the current mining practice as permitted by laws. In addition cost of implementing those methods are not considered here. 6.2 PANEL LAYOUT As shown in Figs. 2.9, 2.10, and 2.11, permanent ground deformations in a subsidence basin mainly concentrate near the edges of the underground opening, and can be divided into four zones. A structure located in different zones will be subjected to different types and magnitudes of ground deformations. In laying out the panels, Table 5.1 and Figs. 2.9, 2.10, and 2.11 could be taken into consideration. Attempts could be made to avoid placing the structure on a location where the ground deformation to which that structure is sensitive is at its maximum. Therefore rational design of the panel is the simplest way to reduce structural defor¬mations. Panel design involves the determination of panel dimension, panel edge location, direction of face advance, and use of yield chain pillars. A. Panel Dimension Since longwalls in the US employ a multiple-entry system, where rows of chain pillars are left unmined, subsidence over those chain pillars is usually smaller. Therefore, whenever possi¬ble, the panel dimension could be designed such that a major structure or structures are over those unmined chain pillars, be¬tween adjacent panels, or some distance beyond both ends of the panels. At the center of the supercritical final subsidence basin, a structure will not be subjected to any final or permanent ground deformations. In order to create such a condition, the panel width must be such that the structure will be located beyond the major influence zone of the final subsidence basin, the minimum dimen¬sion of which must be: [ ] where L is the width or length of the final mined out gob, t is the width or length of the structure to be protected, h is mining depth and [ ] s is the angle of full subsidence. B. Panel Edge Location Wherever there is a permanent panel edge, there are large ground deformations induced on the surface on both sides of the permanent panel edge. Therefore whenever possible the panel di¬mension should be designed such that the permanent panel edges could be located in the areas with the least impacts. In terms of permanent edge location, it is best to eliminate any permanent panel edge under a structure or groups of structures. If this cannot be done, the panel should be lengthened to reduce the number of permanent panel edges, or narrower multiple panels advancing in the same direction in a staggered manner could be employed. If the structures are located in Zones II and III, the longer dimension of the structure must be parallel to the nearest perma¬nent edge (Fig. 6.1). But in Zone IV, the longer dimension should be tangential to the corner of the permanent panel edge. If the structure is inclined to the permanent panel edge, it will be sub¬jected to twisting and shearing. C. Direction of Face Advance The direction of face advance should be parallel to the long axis of the structure. But if the structure is to be located at or close to the center of the final subsidence basin, the direction of face advance should be perpendicular to the long axis of the structure. Careful choice of the direction of multiple face advance is the most effective way to reduce structural deformation and thus dam¬age. This applies the principle of overlapping and cancellations of ground deformations, due to multiple face advance, at the right time and at the right intensity, e.g., opposing tilts, concave and convex curvatures, tensile and compressive strains are induced simultaneously on the structure to be protected by two or more faces. D. Use of Yield Chain Pillars In US longwall panels there are generally two or three rows of stiff chain pillars between the panels. The combined width of the chain pillars ranges from 100 to 350 ft(30 to 107m). depending on mining depth. In general, surface movement above the chain pil¬lars after the panels on both sides have been extracted is much smaller, as compared to that in panel center. Thus in order to create critical or supercritical width of opening and eliminate sur¬face bumps over the chain pillars, yield chain pillars may be em¬ployed (Jarosz and Karmis, 1986). However if yield pillars are to be used, it must be designed such that it yields totally right after the panels on both sides have been extracted. Unfortunately cur¬rent yield pillar design techniques cannot predict when and how much it will yield. In summary, whenever possible attempts could be made to lay out the panel in such a way that surface structures are located above chain pillars between panels or above solid coal beyond both ends of the panels. In those areas the surface structures will most likely be unaffected, or if affected, the damage is so minor that no remedial measures are necessary. 6.3 CONTROLLED MINING TECHNIQUES Several mining techniques are available for reducing the sur¬face ground deformations of specific types. Regardless of tech-

Jan 1, 1992

By M. R. Rosen

CARBINDER Polymer 498 is a liquid organic taconite binder which significantly improves fired compression in both fluxed and acid pellets. It has been used commercially in both shaft and straight grate furnaces. This new binder produces an unusual wicking phenomenon during balling wherein the rate of movement of interstitial water is accelerated to the surface of the balls. The phenomenon is proposed to be responsible for the significant improvement seen in fired compression. This paper presents information as to the effect of CARBINDER Polymer 498 on green ball and fired pellet properties. It also describes the wicking phenomenon and discusses a reversal in the normal balling operation procedures required for successful agglomeration.

Jan 1, 1988

By R. W. Henn

Introduction Shotcrete, by the American Concrete Institute definition, is mortar or concrete pneumatically projected at high velocity onto a surface. The term gunite is still common in literature. It is a trade name established by Cement Gun Co. of Allentown, PA, around 1911 for a dry mix of fine aggregate material, hydrated at the nozzle, and pneumatically projected at high velocity onto a surface. This article will discuss shotcrete applications, ingredients, mix proportioning, batching and mixing, admixtures, testing, and steel fiber vs. conventional shotcrete. Applications Shotcrete is used primarily by the mining and underground construction industries to temporarily stabilize an excavated rock surface until a permanent cover can be installed. It may also be used as the permanent covering material. Other uses include refractory material, architectural finishes, slope stability, wearing surfaces, repairing existing concrete structures, fireproofing, and corrosion protection of steel. Wet Versus Dry Process In the wet process, all ingredients, including water but not accelerator, are mixed the same way as normal concrete before its placement by pneumatic projection. The accelerator is added in a liquid state through the nozzle at the point of placement. In the dry process, all ingredients, including acceleration in dry form but not water, are mixed just before placement by pneumatic projection. Water is added through the nozzle at the point of placement. Liquid accelerator may be added at the nozzle in the dry process. Ingredients Shotcrete consists basically of the same ingredients as concrete with some limitations. Type I, II, and III cements are used. However, Type II seems to be used most in underground applications. This can be attributed to ground water at most sites having slight to moderate concentrations of sulfates. Air-entraining cement or air-entraining admixtures are generally not used in dry mix shotcrete. It is generally agreed that it is impossible to entrain air due to the absence of the usual concrete mixing action, and the high velocity of impingement of the material on the application surface. Air-entraining cement or air-entraining admixtures are used in wet mix shotcrete. Their use, however, is not widespread due to the placement method and the fact that high velocity air drives off most of the entrained air. Test results are not conclusive in showing any durability increase of shotcrete to freezing and thawing by using entrained air. In spite of current practice, the trend is toward additional research, with more agencies and owners requiring the use of entrained air. The philosophy behind this is that some entrained air is better than no entrained air to increase durability to freezing and thawing. The aggregates used in shotcrete are the same as those used in concrete. However, the largest coarse aggregate presently used is 19 mm (0.75 in.). A typical gradation requirement is shown in Table 1. The water quality requirement for shotcrete is the same as that for concrete. Accelerators are by far the most important admixture in shotcrete.

Jan 6, 1984

By Ted D. Johnson

INTRODUCTION The report of the workshop on Alternate Energy Strategies estimated that the world's reserve of uranium oxide is about 700,000 tonnes. In order to extract these reserves, it is estimated that between 100 and 200 million tonnes of solid waste and of the order of 1 billion cubic meters of liquid waste will be generated. In addition, radium within tailing will continue to generate radon gas as it decays for many years after closure. Hence, the handling, storage and possible use of uranium mine waste is a matter which deserves serious consideration and planning. Since the properties of the waste will largely dictate the method of handling and storage or use of the waste, the properties of the waste should be determined, to the degree possible, prior to the design and subsequent operation of the mine. Uranium mine waste has both solid and fluid constituents. The solids include soil and rock removed during site preparation, similar material removed during development of the mine and uranium tailing generated by the mill. Soils can be stockpiled for reuse during reclamation or used as earth materials for construction of fills, roads, dams and soil linings. Rock can be stockpiled, used as mine backfill or can be used for earth construction in fills, roads and dams. The tailing is usually placed in a tailing storage area. The fluid constituents are both liquid and gaseous. The liquid waste is generally composed of the mill water; however, additional waste water can be generated by dewatering processes and diversion of surface runoff. The gaseous constituent of the uranium mine waste is radon generated by the radioactive decay of radium in the ore tailing. DEFINITION OF PROPERTIES The following presents a brief definition of some of the more important properties discussed later. The definitions are not intended to be comprehensive, but are useful as a basis for understanding.

Jan 1, 1980

By William C. Cox, Ronald E. Heuer, Jean-Michel Loignon

The Gatineau shaft described in this paper was an excavation of about 38 m diameter and 21 m depth constructed in Gatineau, Quebec, across the Ottawa River just north of Ottawa, Ontario. A large sewage lift pumping station-was to be constructed in this excavation. Subcontractor for the shaft excavation and initial support was Beaver Underground Structures Ltd., of Dorval, Quebec. Mr. William C. Cox of Beaver conceived the idea for the support system used in this excavation, and was responsible for the construction activities. Initial analysis and design concepts for the lining system were developed by Mr. Ronald E. Heuer. Detailed lining design and construction monitoring were the responsibility of Mr. Jean-Michel Loignon. Geologic conditions indicated by two borings within the shaft area are illustrated in Fig. 1. The shaft excavation would be in the sensitive Leda silty clay. The clay is of stiff to very stiff consistency, with an insitu vane shear strength averaging about 115 KPa (2400 psf) and a reported low value of about 87 KPa (1820 psf). Low strength values were from the middle of the clay layer, with higher strengths at the top and bottom. Average shear strength at the level of the base of the excavation was about 137 KPa (2860 psf). Remolded vane shear strength averaged about 20 KPa (420 psf), with a reported low of 13 KPa (270 psf). Based upon average strength values, the clay sensitivity is 5.8. The clay has an average water content of 53% (range 43 to 66%), which is near its liquid limit (average 56%). The clay unit weight is about 1760 kg/m3 (110 pcf). Underlying the shaft site at shallow depth is limestone bedrock within which the piezometric head is approximately equal to the original ground surface elevation. For shaft support, a -cast-in-place concrete wafer system was considered, but was discarded because of time restrictions imposed by concrete placement and curing. The support system selected utilized liner plate with light steel walers closely spaced. The light rib

Jan 1, 1983

By T. J. DeMull

AMAX's Sleeper Mine uses conventional cyanidation milling and heap leaching techniques to produce gold. Since milling operations commenced in March 1986, average ore processing rates have improved from the design rate of 455 tonnes per day to more than 720 tonnes per day. Heap leaching was added to the process facilities in September 1986 at a design rate of 1,500 tonnes per day, and current heap leaching rates exceed 2,680 tonnes per day. Presented are descriptions of the Sleeper processing facilities with emphasis on the crushing and grinding circuits, a review of the selection of crushing and grinding equipment, and a summary of operating performance during the first 16 months of operations.

Jan 1, 1988

By N. D. Jagger

In order to maximize the benefits obtained from flocculating agents when used as settling aids in high rate and conventional thickeners, it is often necessary to dilute the thickener feed prior to flocculant addition. In some cases, no flocculation at all is possible unless dilution water is used. The most common method of diluting thickener feed has been to recirculate some of the thickener overflow by means of a centrifugal pumping system. This represents a significant capital and operating cost. Supaflo Technologies of Australia has developed and patented an internal recirculation method known as Auto Dilution (or "Autodil") which is based on the principle that an hydraulic head differential exists between the inside and outside of the feedwell because of the higher S. G. of the pulp inside the feedwell, By installing non-return ports in the feedwell wall at the appropriate level, supernatant solution flows into the feedwell at a calculated rate and dilutes the feed without any need for external recirculation of overflow.

Jan 1, 1993

Introduction The frigid regions of Canada's high arctic islands are well-known for major oil and gas explorations. But mining leads the way in development. Cominco Ltd.'s Polaris mine is in its second full year of commercial production, and outstripping the rated production capacity foreseen by its planners. The Polaris mine is located on the shore of Little Cornwallis Island, near the settlement of Resolute Bay. The mine property is compact. It occupies about 3 km2 (0.5 sq miles) of the vast Arctic landscape. A number of mining and technological firsts were necessary to bring a mine into production at this 76° North latitude. Development constraints that had to be overcome included climate, remote location, and the limited ice-free shipping season. Construction of the entire process plant and ancillary facilities took place on a barge in the South where labor and supplies are plentiful. The plant was towed through ice-strewn waters to its permanent location at the mine site 1450 km (900 miles) from the North Pole. Freighters moved the first shipments to the South during ice break-up in late summer 1982. Most of the zinc concentrate is sold to European smelters. The remainder is tolled at a custom smelter in Europe, and the resulting metal is sold by Cominco. Lead concentrate is sold to European smelters. The mine was designed to produce 177 kt/a (195,000 stpy) of zinc concentrate and 41 kt/a (45,000 stpy) of lead concentrate. Actual production in 1983 was considerably greater, with both mill throughput and feed grades being higher than forecast. Production was 217 kt (239,300 st) of zinc concentrate and 51 kt (56,300 st) of lead concentrate. Average milling rate was 2.3 kt/d (2500 stpd), well above the designed capacity of 2.1 kt/d of (2300 stpd). Arctic Climate Extremes The climate is typical of the Arctic: very dry with short, cool summers and long, cold winters. From November through February, there is continuous darkness. Continuous daylight prevails from April to August. Precipitation is light, rivaling the Sahara desert at about 127 mm (5 in.) of rain and snow a year. An almost continual wind forms deep snow drifts against any obstacle on the ground. Temperatures range from a winter low of about -50°C (-58°F) to a relatively pleasant 15°C (59°F) during sunny days in August. Freeze-up begins in September and ice thickness increases from less than 0.3 m (1 ft) in early October to a maximum of about 2.1 m (7 ft) in May. The permanently frozen ground extends to a depth of about 427 m (1400 ft). Only a few inches of surface ground thaws in the summer. Exploration Activity Zinc-lead mineralization was

Jan 10, 1984

By Charlie Sons

Pit and underground mine operators typically use a lot of big equipment tires. Just as typically, they spend a lot of money they shouldn't on these tires. That's because tire use and maintenance are not given enough attention. That often occurs because mine management and personnel do not know the basic facts of tire use and care. If a piece of large equipment is damaged or ruined, a good explanation better he forthcoming. A ruined tire is often just a shoulder shrug. But that sort of fatalistic acceptance can cost tens or hundreds of thousands of dollars a year extra for big equipment tires. That does not have to be. There are some basic tire considerations that, if followed, will extend tire life and lower tire costs. Most companies have the inevitable conflict between the production department, which resists any restrictions that may reduce productivity, and the maintenance department. Production success is based on meeting a schedule of yards moved, while maintenance success is based on availability and equipment cost. When production falls behind schedule and equipment speeds are increased, maintenance may soon find its tire costs rapidly increasing, unless everyone understands the tires' vulnerability to damage. Unless the production department understands the problem, or restraint is applied by top management, costs may become exorbitant. At one job site, tire costs alone exceeded $250,000, to move 55 dam3 (72 000 yd3) of dirt on an 11-km cycle. The operation temporarily caught up with its schedule, but tire costs and attendant downtime soon made this a not-for-profit exercise. Because an operator must always attempt to move payload at the lowest cost, there are cases when higher than normal tire costs must be accepted. For example, in pioneering work with tandem engine scrapers, steep grades shorten tire life to less than normal. But the steep grades make shorter haul cycles possible, yielding more trips per hour than alternative earth-moving methods.

Jan 1, 1982

By Gene E. McClelland

4.1 INTRODUCTION Metallurgical testing is required on any ore, waste or tailing material before amenability to heap leaching treatment can be accurately determined. The amount of testing necessary to determine and optimize heap leach amenability and conditions for any feed material is determined by the ore type, grade, size of deposit, and the overall financial commitment a particular company is willing to make. It is very easy to either "under-test" or "over-test" a particular deposit or feed material. A good heap leach testing program will provide the necessary design information without "bogging" down in too much detail. The three basic stages of ore testing include: • Preliminary--bottle roll tests and small column percolation leach tests; • Detailed--larger column tests, different feed sizes, and feed size optimization; and • Pilot-Scale--Column or field heap. The following subchapters describe the stages of heap leach testing necessary to put a viable property into production. 4.2 PRELIMINARY METALLURGICAL TESTING The two types of preliminary tests that can be undertaken include bottle roll or cyanide agitation tests and small column leach tests. Each of these is described below, along with information about what the test results provide in terms of furthering knowledge of ore qualities. 4.2.1 Bottle Roll Tests The first step in a heap leach testing program is preliminary bottle roll tests at a relatively coarse feed size to determine the ore's leachability. Bottle roll tests are usually conducted on reverse circulation drill cuttings samples (= 1/4 inch (6 mm) in size). Bottle roll tests provide initial information about precious metal recovery, recovery rate, and reagent requirements. The recovery results are to be considered maximums because attrition grinding creates fines and liberates recoverable values that would not be liberated from the material during a static leach. Bottle roll tests can be conducted on up to 25 lb (11 kg) ore charges of up to two inches (50 mm) in size. However, the coarser the feed the more attrition grinding occurs during rolling.

Jan 1, 1988

By M. A. Schimmelpfennig, A. D. S. Gillies

M.J. McPherson Having worked on the thermodynamics of air/liquid-water mixtures passing through the surface fans of deep mines, I find this paper of great interest and congratulate the authors on producing it. There are two matters, however, deserving discussion. First, the authors have described the classical theory of fog formation with fogging occurring at "supersaturation." In fact, the process of fog formation begins well below 100% humidity. The more strongly hygroscopic nuclei in the atmosphere will attract water molecules and begin to grow at a relative humidity perhaps as low as 70%. The process is a dynamic one with both condensation and evaporation taking place throughout the mixture on microscopic liquid surfaces. As saturation is approached, the rate of condensation accelerates rapidly producing the familiar reduction in visibility. Hygroscopic nuclei are present in all natural atmospheres and under the appropriate conditions of pressure and temperature, will produce clean fogs. However, if the air is polluted by particulates from combustion or other processes then the resultant coagulation with growing liquid particles may produce dense (and sometimes photochemical) smogs. It is this process that is likely to occur in underground mines when moist air is cooled below dew point. Second, the authors have summarized very well the individual measures that might alleviate the problem. In particular, I agree that neither heating nor refrigerating the air, by themselves, provides a satisfactory solution. However, there is a combination of these that provides a neat and effective control of fog formation. This involves a small, self-contained refrigeration unit within a duct but without the usual external heat rejection facility. The air is cooled below dew point and, hence, dehumidified on passing over the cold evaporator coils. The heat from the condenser is rejected back into the air downstream from the water eliminator, as shown in Fig. 1. The duct configuration can be designed to create good mixing at the outlet. I have used a climatic simulation program to illustrate the effect, assuming a 500-m (1,640 ft) long airway of cross section 60 m2 (645 sq ft), an airflow of 9 m3/s (318 cu ft per sec), inlet conditions of 19/19.25°C (66/67°F) wet bulb/dry bulb temperatures, a virgin rock temperature of 17°C (63°F) and typical rock thermal properties for a hard-rock mine. I have also assumed that 10% of the rock surface is wet. Figure 2 shows the variation in temperatures along the airway if no measures are taken. The relative humidity remains close to 100% and condensation will occur throughout the length of the airway. Before any such installation is designed, the exercise must, of course, be repeated for each location with accurate data. However, this example demonstrates the potential of the system. The idea is not new - the principle is sometimes used in air-conditioning plants for buildings. Figure 3 shows the effect of removing 30 kw of heat by evaporator coils sited 40 m (131 ft) along the airway and rejecting that heat (plus another 10 kw of compressor and fan power) into the air at 60 m (197 ft). Condensate water is produced, at a rate of 0.45 L/min (0.12 gpm). The result of this sequential cooling/dehumidification/heating process is to separate the wet and dry bulb temperatures along the length of the airway. The relative humidity increases from 80% at the duct outlet to 92% at 500 m (1,640 ft). The airway is maintained free from fog.

Jan 11, 1983

By G. Brien, O&apos

Selective in-plant breakage was investigated at a Central Queensland, Australia mine. Component seams and size fractions from in-plant samples were characterized petrographically and probability analysis of vitrinite reflectance data used to track the distribution of seam proportions through the plant. Results show that the 2 seams reported differentially to coarse and fine circuits, mainly as a function of coal type rather than rank. Seams were also preferentially concentrated in different size fractions of the coarse circuit. Breakage in the coarse coal centrifuge generated approximately 4% of -0.5 mm fines, but was not seam specific.

Jan 1, 1996

By R. H. Moore, B. D. Breitenstein

INTRODUCTION The United States Uranium Registry (USUR) tissue program was funded in 1978 by the United States Department of Energy (DOE) through its Human Health and Assessments Division, and registrant enrollment began in 1980. It is operated by Hanford Environmental Health Foundation (HERF), Richland, Washington, with support provided by Pacific Northwest Laboratory (PNL).** The USUR tissue program is the first systematic effort to contact and gain assistance of workers occupationally exposed to uranium and to seek and study uranium deposition in human tissue. This paper describes the program objectives, populations selected for study, procedures for enrolling workers in the program, types of tissue to be studied, and analytical procedures to be used. PROGRAM OBJECTIVES The objectives of the USUR tissue program are to: 1) Determine the distribution and concentration of uranium, if any, in the tissues of occupationally exposed workers. 2) Compare bioassay measurements of exposed individuals with the results of analyses of tissues obtained at autopsy. 3) Seek evidence of histopathologic changes related to any uranium deposition that may be present. 4) Conduct analyses of whole bodies, when available, to obtain more precise data on the uranium burdens, if any, in the body and organs, and especially the distribution in parts of the body, such as the skeleton, that are not usually accessible for sampling. 5) Develop data that will assist in evaluating a) the accuracy of current in-vivo measurement techniques, b) the propriety of existing regulations, and c) the adequacy of current protection programs. SELECTION OF POPULATIONS FOR STUDY Selection of the populations for the USUR tissue program was based on studies of the United States Uranium Registry described elsewhere (0c81). Potential populations for study were quite varied because of occupational exposure to different chemical forms of uranium, different levels and eras of exposure, varying ability to identify populations exposed in the past, and the general interest and cooperativeness of the populations. Nevertheless, certain identifiable populations emerged that were both willing to participate in the study and likely to provide useful information. Among the targeted populations are individuals who were exposed to uranium prior to the existence of modern exposure standards. These individuals are presently being enrolled by the USUR. The reliability of concentration and distribution data and the demonstration of effects are more likely in such individuals than in those who have received less exposure. However, the targeted groups also include individuals who Have been exposed to permissible levels of uranium, and reference employees in uranium plants who have not worked with uranium. The lesser exposures are of immediate interest because of the rarity of heavy exposures under current standards. The sampling of potentially exposed individuals with negative in-vivo measurement results may serve to check the adequacy of in-vivo measurements in the detection of deposition. [ENROLLMENT PROCEDURE] The enrollment of participants follows three patterns. 1) Presently employed workers are enrolled by the cooperative efforts of USUR staff with the medical, industrial relations, and/or health physics personnel of the uranium facility. Volunteers may be enrolled by medical personnel at the time of their physical examination or by health physics personnel at the time of a bioassay or lung counting procedure. Follow-up after initial contacts with currently employed uranium workers is carried out by on-site personnel, but subsequent follow-up is the responsibility of the Registry. 2) Retired workers are contacted by mail, with a covering letter from their former employer supporting the principles of the study but leaving participation in the program on an entirely voluntary basis. 3) An alternate method of enrollment is to obtain a USUR autopsy consent after death from the next of kin. As part of enrollment in the program, permission is requested for access to the individual's medical and exposure data, and a short occupational history is filled out. The autopsy permission agreement signed by the individual is for an initial period of 5 years; the USUR ordinarily seeks renewal of the agreement upon its expiration. While this agreement provides

Jan 1, 1981

The Mine Safety and Health Administration (MSHA) academy in Beckley, WV, received enquiries from various segments of the mining industry concerning alcohol and substance abuse in the mining industry. The MSHA Director formed a committee, the Mining Industry Committee on Substance Abuse (MICSA), to deal with these enquiries. The committee met in April 1985 at the mining academy in Beckley. The MICSA committee expanded its membership to represent labor, management, and government agencies affiliated with the mining industry as well as government agencies dealing with alcohol and drug abuse. This expanded 18-member committee met for the first time in July 1985. At that time, two cochairmen were elected, one from labor and one from management. MICSA has spent time on education by sharing information on the issues and by calling in outside experts for advice and consultation. The committee also initiated several projects. These include a resource manual on substance abuse and employee assistance programs. Mine operators and labor unions can use these programs as guides. MICSA is also assembling statistics on substance abuse, particularly as applied to mining. And the committee commissioned an Education film, Substance Abuse, Is It Our Problem? (The results of a survey on drug use, given in Table 1, suggest that substance abuse may indeed be a problem.) The educational film, produced by MSHA, is an introduction to the problem of substance abuse in the mining environment. The film encourages mine operators, labor unions, and miners to analyze their individual circumstances to determine whether a problem exists. The film also offers ways to deal with the problem. And the film invites the mining industry to request additional information from MICSA. There is no single approach to substance abuse that will fit the need of every mining operation. As with any health, safety, or production problem, remedies must be site specific. MICSA's aim is to channel ideas to those who seek to help employees in a mining environment who seek to rid themselves of alcohol and substance abuse. As stated more formally: The Mining Industry Committee on Substance Abuse represents the efforts of labor, management, and government. The mission of the committee is to make the mining industry aware of the problems resulting from alcohol and other drug abuse and to recommend possible methods to remedy the problem. The function of the committee is to assist the industry in developing programs and resources to recognize and provide help for troubled employees, and thus reduce the risk of accidents, while lower absenteeism and increasing productivity.

Jan 11, 1987

By Thomas G. Hayes

Track-type tractors remain one of the most cost-efficient stripping tools available to coal miners. Whether they are "sweetening up" a dragline operation with prebenching or serving as the primary tool, dozers are important in driving down mining costs. This article reviews the application zone of track-type dozers. It also provides suggestions for site setup, equipping tractors and machine operation. In recent years, the track-type tractor has become more of a production tool in mining. With the advent of bigger machines, bulldozers have become a typical component in mining operations, removing overburden down to the top seam of coal. Large tractors can also be teamed up with draglines and other systems to maximize the effectiveness of the mining system. Bulldozers remain the tool of choice for topsoil removal, spoil reclamation or any application that demands low ground pressure, or short distance. The introduction of the 523-kW (700-hp) track-type tractor to the mining industry has made the bulldozer a low-cost-per-ton means of removing overburden. The large tractors offer enough productive capability to be considered prime stripping tools. And they can lower mining costs by complementing other mining systems (dragline, cast blasting and loader/truck). Four criteria determine whether a track-type tractor is the right machine for a particular application: ? Production requirements - dozer may fit job requirements better than a second dragline. [ ] ? Product support - availability of parts, service and technical expertise. ? Labor situation - how much of the cost per ton goes to the operator. ? Resale opportunity - recover a percentage of capital cost. Mining efficiency is economically controlling the cost of any system. Most operations rate the most economical means of moving material as cast blasting, dragline and the track-type tractor. All three of these systems share one common value: They are all sensitive to distance. Regardless of the application, all mining systems have an economic range. Exceeding this distance will affect cost and productivity.

Jan 1, 1997