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By G. P. Hand

The graphite industry as a whole has gone through many changes over the years. Most of the graphite used in major applications since the turn of the century were basic natural or electrode by-product synthetic materials, processed using simple crushing, screening, or grinding techniques. In the mid 1970's new thermal technology was developed which changed the face of the .graphite industry, and opened up a panacea of new products and capabilities. Materials produced through this new technology have helped revolutionize new products and processes in industries including automotive, batteries, fuel cells, electronics, iron, lubricants and steel. This paper outlines the continued improvements in those industries as well as the as well as the outlook for new technology in this rapidly changing area.

Jan 1, 1997

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Shrinkage stope mining can be employed in steeply dipping veins where both the wall and the vein material are sufficiently strong to stand with only minimal ground support. Shrinkage stope mining takes advantage of the swell in volume produced in breaking in-place ore. This expanded volume of ore is allowed to occupy part of the mined-out vein, and in so doing provides ground sup¬port and a working surface for the miners. Only enough of the ore is drawn after each stoping cut to allow work¬ing room. Not until the entire mining block has been broken is the full amount of broken ore drawn from the stope. DEVELOPMENT PRACTICES FOR SHRINKAGE STOPING The vein is developed into stoping blocks, each of which is 76 to 91 m (250 to 300 ft) in length and 61 m (200 ft) between levels. Level development, described later, is either on the vein or parallel to it with crosscuts periodically driven over to the vein. There are usually two exits from each stope. The first entry is a raise at one end driven either conventionally by raise climber, or by a raised borehole. Present practice is that this raise, put in from level to level, is either a 2.1- x 2.1-m (7- x 7-ft) Alimak raise or a 1.2- to 1.8-m (4- to 6-ft) diam borehole. Either raise system will require pro¬visions for an ore chute carried with the stope to the haulage level or the scram drift. At 9- to 12-m (30- to 40-ft) vertical intervals, dogholes are developed from the raise so that when stoping breaks back into these dogholes, timber or steel sets and necessary service facilities can be installed. The second entry to the stope may be either a similar raise at the stope endline, or more economically, a cribbed raise carried with the stope. As previously stated, development of the ore block is achieved by either drifting on the vein or by paralleling the vein with a footwall lateral and then crosscutting over to the vein at specified intervals. In the following discussion, three basic cases are illustrated in Figs. 1, 2, and 3. In Fig. 1, development is in the footwall by a 3- x 3-m (10- x 10-ft) trackless lateral. A heading is also driven on the ore and taken either full width or driven a standard size and slabbed to the vein width. Approxi¬mately 9-m (30-ft) long crosscuts on 7.6-m (25-ft) cen¬ters are driven from the lateral to the vein. For ease in muck handling they are usually put in on a slant rather than at right angles. A single orepass, located off the lateral and in the center of one stope block, can serve not only that stope, but also one on either side, up to a tramming distance of not more than 137 m (450 ft). Raises are put at the stope endlines on 91-m (300-ft) centers. The back of the vein is shot down and just enough broken material is drawn off to permit access for men to work off the broken ore. Development in Fig. 2 progresses by driving a 3- x 3-m (10- x 10-ft) trackless heading on the footwall side of the vein. The ore block is delineated by raises on 91-m (300-ft) centers. Crosscuts are again driven on 7.6-m (25-ft) centers over to the vein. Approximately four rounds are then taken up to the vein and each round is successively belled out so that all the crosscuts (pockets, in this case) connect. In this development, as opposed to that in Fig. 1 where there were no belled

Jan 1, 1982

By H. K. Lin

During slow-cooling crystallization of iron-making blast furnace slags, most rare-earth elements in the feed recrystallize as cerite-calcium while scandium and europium concentrate in a sheet-like "mineral". The mineralogy of this "mineral" is studied with one of the objectives being the extraction of the "mineral" from the slag for recovery of Sc and Eu. This "mineral" has been identified as one of the mica group minerals and can be named as rare-earth-bearing barium fluorophlogopite. The mineralogical characteristics of this "mineral" are investigated and discussed. This "mineral" can be effectively recovered from the slag by flotation with dodecylammonium acetate or a combination of dodecylammonium acetate and sodium oleate as a collector.

Jan 1, 1996

By B. J. Scheiner

A ferrous chloride-oxygen leaching system was investigated by the Federal Bureau of Mines as a means of re-covering copper and silver from a sulfide concentrate that had been pretreated to remove a majority of the antimony. The pretreated concentrate contained Cu, Ag, Pb, Zn, Sb, and As. Parameters affecting metal extraction, such as temperature, pressure, amount of ferrous chloride, and initial hydrogen ion concentration, were studied. Copper extraction of 98 pct was achieved by leaching 4 hours at 40 pounds per?square inch gage (psig) and 1000 ?C. Copper was recovered from leach solutions by cementation with iron, and the resulting ferrous chloride was recycled. Leaching the residue with cyanide followed by electrowinning recovered 99.7 pct of the silver. Iron, Sb, As, Pb, and S reported to the tails.

Jan 1, 1978

By Robert G. Beverly

INTRODUCTION Standards limiting the annual exposure of United States uranium miners to radon daughters were established in 1967 at 12 Working-Level-Months (WLM). The standard was reduced by a factor of three, to 4 WLM, in 1971. Currently, the standard is again being examined to determine if it should be changed. Since 1967, Union Carbide has calculated individual monthly exposures in company and contract-operated mines located on the Colorado Plateau. Although it has been possible, by extensive ventilation control measures and accurate routine sampling, to meet the current exposure standard, there are many miners whose exposures closely approach the 4 WLM standard for any given year. However, it was noted that for miners who work for any extended period of years the [average] exposure was much less than the standard. The primary purpose of this paper is to show that, in effect, any annual exposure standard to radon daughters results in a long-term exposure considerably below that standard. Further, most miners, due to their job assignments and/or employment habits, only receive a small fraction of the standard. HISTORY OF EXPOSURE STANDARDS Prior to 1967, radiation protection in uranium mines was fundamentally based on a radon daughter concentration guide. In 1960, the American Standards Association published mine and mill radiation protection standards (ASA-1960). The Colorado Department of Mines, in 1961, adoped a standard which followed the ASA Standard and provided that if concentrations exceeded 10 Working Levels (WL), the area was to be shut down until corrective action was taken; if between 3 and 10 WL, corrective action was to be initiated; between 1 and 3, additional samples were to be taken and individual exposures evaluated; and if below 1 WL, conditions were considered to be controlled. In 1967, the U.S. Department of Labor issued the first exposure standard which called originally for limiting annual exposures to 3.6 WLM but which was later changed to 12 WLM. The complicated regulatory developments leading to this standard have been described elsewhere (Beverly-1969, Rock & Walker-1970). Effective July 1, 1971, this exposure standard was lowered to 4 WLM per year, which is the current standard. Over the past year, there has been speculation about the potential risk to uranium miners working at the present standard. A recent NIOSH Study Group Report (NIOSH-1980) concluded: "There is also strong evidence that a substantial risk extends to and below 120 WLM of exposure." The 120 WLM corresponds to a miner working in uranium mines for 30 years, a rare occurrence, at an exposure rate of 4 WLM per year, an even rarer occurrence. On the other hand, the General Accounting Office, in a recent Report to the Congress (GAO-1981), was very critical of reports by NIOSH on general low-level radiation risks. The GAO recognized that”...important questions remain unanswered about the cancer risks of low-level ionizing radiation exposure;" and recommended that Congress enact legislation giving statutory authority to an interagency committee to coordinate Federal research on health effects of ionizing radiation exposure. The International Commission on Radiation Protection at its March, 1980 meeting recommended limiting the inhalation of radon daughters to 0.02 J per year, equivalent to 0.4 WL, which on an annual basis would be 4.8 WLM and noted it is common to reduce this figure by 20% for allowance in the case of uranium miners for external and/or dust exposure(Sowby-1980). This is essentially equal to the present standard of 4 WLM. As earlier uranium miner exposure studies are reevaluated, and as new studies are conducted, it is important that the relationship between regulatory standards and the resulting actual exposures be recognized. UNION CARBIDE URANIUM MINING EXPERIENCE Union Carbide started mining Colorado Plateau uranium-vanadium ores in the late 1920s for the contained vanadium values. In the early 1950s, the Atomic Energy Commission contracted Union Carbide to produce uranium at mills located in Uravan and Rifle, Colorado. The company now has over fifty years of mining experience in the area. Some mines are operated as company mines and others are operated by private mining companies under a contractual arrangement. Ventilation, sampling, and exposure calculations are carried out the same in contract mines as in company-operated mines. Data presented in this report do not differentiate between company or contract employees and include all employees who worked underground any portion of a year in Union Carbide mines from 1967 through 1980. At the peak of uranium mining activities in 1970, there were 577 miners employed at year end (285 company employees and 292 contract) and 52 mines in operation (8 company-operated and 44 contract mines). Contract mines varied from two-man operations up to 15 employees. Company mines were generally the larger operations and employed from 20 to 100 miners.

Jan 1, 1981

By J. A. Zimmer, M. S. Hillis, D. Moredock, J. A. Organiscak, A. B. Cecala

Enclosed cabs have been used on surface mining equipment for many years to protect equipment operators from health and safety hazards. The surface mining and quarrying process is a highly mechanized process that generates large quantities of respirable dust. Due to economic factors, more older mining equipment is being used today than in years past. With this older equipment, it is common to observe deterioration of many components on the enclosed cabs, which greatly reduce their protection effectiveness. NIOSH has performed a number of different field studies, as well as a comprehensive laboratory study, evaluating various factors and parameters regarding enclosed cabs. NIOSH has recently completed a cooperative research effort to evaluate the effectiveness of a newly designed unidirectional powered precleaner, filter and pressurizer that brings clean filtered air in at the roof of the cab and extracts all recirculated air near the floor. This research effort provided another example of the enhancement in air quality that can be achieved in an enclosed cab from the implementation of an improved filtration system and by establishing positive pressurization from increased intake air volume and sealing the cab. This testing showed a significant reduction in respirable dust levels in the enclosed cab when comparing baseline versus post-testing results. Since dust levels were similar near the roof and floor of the cab during post-testing, the unidirectional flow pattern did not appear to significantly improve the performance over the conventional design in this study.

Jan 1, 2009

By C. Alan Tapp

Introduction Mine sampling determines the practicality of any mining operation. Improper sampling can result in an incorrect appraisal of present production and future potential. Therefore, the mine department in charge of ore-reserve calculations and mine sampling should be overseen by competent and experienced professionals with technical backgrounds qualifying them to produce accurate results. Sampling is a process by which portions of an ore body are collected and analyzed to estimate the average mineral content of the entire ore body. It is incorrect to assume that a large number of samples eliminates any errors in the sampling method. To obtain unbiased samples, proper sample location with respect to rock type and mineralization is just as important. The sampling procedure must yield correct results for the type of mineral deposit, and careful consideration should be given to whether or not the sampling technique has been developed to an extent sufficient to eliminate as much human error and bias as possible. Only after the ore has been mined and milled is the sampling accuracy known. Sampling also provides information about the bulk composition of the ore for mineralogical and metallurgical tests that determine the economic ore-waste boundaries and the geologic trends for exploration. Actual mining plans can be developed from this information to maximize profits. Accurate sampling is critical, and thus must be approached in a scientific manner. Sampling Techniques Sampling practices and techniques are as varied as the mines in which they are used. The method(s) chosen must be tailored to suit the company and mining needs. For instance, tabular uranium deposits, vein gold deposits, and porphyry copper deposits pose special problems in conducting unbiased sampling. The mine geologist or engineer in charge must develop a sampling method, test it in a sample area, and then critically evaluate the results. If the results from the test area are accurate within the economic limits established by the company, they then may be adopted for general use in the mine. Four routine sampling methods are suitable for specific sampling objectives in the daily mine routine: (1) channel sampling, (2) chip sampling, (3) grab sampling, and (4) bulk sampling. The final sampling results depend upon how the four methods are combined to accurately determine the grade of the ore body. When used in conjunction with each other during different stages of mine development, the channel, chip, grab, and bulk sampling methods provide an in-house check or a comparison by which mining methods and sampling procedures can be evaluated. However, the most valid check is based upon the daily mill production. Channel Sampling Channel samples consist of cuttings collected from a groove cut into the rock about 102 mm (4 in.) wide and 19 mm (0.75 in.) deep. Various tools ranging from a 1.8-kg (4-lb) hammer and moil to a pneumatic chisel can be used to cut the sample. Accessibility and rock hardness determine the applicable sampling tools. Before attempting to take a sample, the rock surface must be cleaned thoroughly; the method of cleaning depends upon the amount of mine dust accumulated on the surface or the degree of alteration of the rock surface. Typical cleaning methods 'employ a wire brush, water, or chipping a fresh surface. Next, the sample outline is marked on the prepared surface, taking care to choose appropriate sample loca- tions. After determining the proper sample outline and length, the sample can be chiseled out, catching the rock fragments on a canvas tarpaulin on the floor, in a powder box, in a canvas bag to avoid contamination, or by some other suitable means to capture all of the sample. Vertical veins present a special problem because the drift's back usually is arched, not square. For example, Fig. 1 shows that if sample intervals are measured from A' to D', the length of the sample interval is greater than the true vein width, resulting in an incorrect calculation of the ore reserves. The correct method would collect three samples at A'-B', B'-C', and C'-D', using either actual measurements in the mine represented by A-B, B-C, and C-D or trigonometric calculations to determine the true ore thickness. Then, each sample can be weighted by grade and true thickness for the vein. Channel lengths usually are a maximum of 1.5 m (5 ft), and it is good practice to divide longer samples into smaller intervals according to structures, changes in rock types, or differences in rock hardnesses. The influence of those features on the mineralization then can be determined. Chip Sampling Chip sampling is a variation of channel sampling used when the rock is too hard to channel sample economically or when little variation in the mineral content indicates that this sampling method will yield results similar to those of channel sampling. Rather than cutting a channel in the rock, small chips are flaked off at regular intervals over the entire face or area being the mine. sampled. Care must be taken to assure that the sample is representative of variations in the rock hardness and type. hi^ method is fast and useful in preliminary evaluations, but it should not be used for quantitative ore-reserve calculations. Grab Sampling Grab sampling is a fast method for double checking either channel or chip-sampling procedures, and, in some instances, mine production can be estimated from carefully taken grab samples. Grab sampling takes equal amounts of material at selected intervals over a mine dump, a muck pile, or from an ore car to estimate its Channel Sampling mineral content. Generally, this method is not considered reliable. Many independent variables can affect this type of sampling process. Thus, if the ore occurs in the softer fraction and a proportional amount of the result-

Jan 1, 1982

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

1.1 INTRODUCTION When underground mining involves total extraction, it induces overburden strata movements. If not properly planned it causes surface subsidence and affects surface environmental conditions. Total extraction usually refers to longwall mining and room and pillar mining with pillar extraction. Surface subsidence has long been a subject of intensive research for scientists all over the world and considerable achievements have been obtained. However due to its difficulties and complicated nature, research into overburden movement has been thus far incomplete as compared to that into surface subsidence. Since surface subsidence is a manifestation of the results of overburden movement, the processes and mechanisms of overburden movement must be fully understood in order to establish the mathematical prediction models of surface subsidence. In this chapter the processes of overburden movement and its zones of movement in the overburden will be discussed. 1.2 PROCESSES OF OVERBURDEN MOVEMENT When total extraction is used, it produces a large void in the coal seam and disturbs the equilibrium conditions of the surrounding rock strata. The roof strata bend downward while the floor heaves. When the excavated area (or gob) expands to a sufficient size, the roof strata will cave. As a result, the overlying strata continue to bend and break until the piles of the fallen rock fragments are sufficiently high to support the overhanging strata. At this tie the overhanging strata no longer cave, but bend and rest on the underlying strata. Strata bending and subsidence develop upward until reaching the surface and forming a subsidence basin. The whole overburden strata and the surface subsidence basin will further go through a period of compaction and gradually become stabilized. Current knowledge regarding the process of overburden movement has been derived from several sources. One is the direct observation of the mined sections and their surfaces, another from field monitoring of strata movements in the overburden, and others from computer modeling and scale modeling in the laboratory. For example, Figs. 1.1 and 1.2 show the results of a borehole monitoring experiment (Borehole B-2) using the full profile borehole inclinometer and full profile borehole extensometer for lateral displacements and vertical subsidence, respectively (Conroy and Gyarmaty, 1983). The panel, 400 ft wide by 5000 ft long and 630 ft deep, was extracted from the Pittsburgh No. 8 seam which had an average thickness of 54 in. The face advanced from east to west. Curve A in Fig. 1.1 shows the inclination of the borehole when it was 45 ft ahead of the face at Position A (i.e., on the solid coal side). The total deviation of the borehole was 1/7000. The first shear movement occurred at 130 ft above the coal seam and appeared to have occurred along the bedding plane with large contrast of rock strata on both sides, i.e., sandstone vs. shale. When the face had passed the borehole 79 ft at Position B, the borehole deformed to assume a bow shape with a maximum deviation of 2.8 in. from the center line. Numerous shearing planes occurred and extended further upward. When the face had passed the borehole by 75 ft at position C, strata subsidence was measured. Fig. 1.2 shows that a strata separation as large as 4 ft occurred at 40 to 100 ft above the coal seam. Above this level, the strata subsided more uniformly with bedding separations within 1 to 2 in. Another example is the longwall mining with complete caving in the Soviet Union's Karaganda Coalfield (Kolebaeva, 1968). The coal seam was 6 ft thick and 154 ft deep. The immediate roof was the fine-grained sandstone interbedded with coarse-grained sandstone. Above this, there was sandy shale and shale. In order to monitor the process of overburden movement, 15 stations were established in a borehole (Fig. 1.3). The elevations of those stations were measured when the face was at various distances from the borehole. Fig. 1.3 shows the movement history of each station. For convenience of comparison, the movement history of each station was plotted on a common reference point as shown in Fig. 1.4. Clearly, the movements of strata from Station # 1 to #6 were nearly simultaneous, i.e., they behaved as a single unit. The strata from Station #7 to # 15 had differential separations, the maximum of which was 40 in. As the face passed by and moved away, strata separations in general reduced gradually. But strata separation between Stations #12 and #13 remained the same until the end of the monitoring period. The total bed separation between Station #7 and # 15 reached 11 in., approximately 18% of seam thickness, when the face was 132 ft beyond. Based on the above observations, the author developed a conceptual model of overburden movement due to underground longwall mining (Fig. 1.5). The above-mentioned two case studies illustrated that subsidence in the overburden strata propagates upward and subsidence velocity decreases from the bottom to the top. When the subsidence velocity at the surface reaches the maximum value, the subsidence velocity at the bottom portion of the overburden strata has decreased considerably and the strata have begun to compact. Bed separation occurs within a certain distance above the coal seam and reduces from the bottom to the top. When the face has moved away, bed separation reduces gradually, so that eventually some beds completely close and others partially close. Bed separation reaches the highest level when the subsidence velocity at the surface is at its maximum value. For instance in Fig. 1.4 surface subsidence velocity was maximum when the face was 38 to 44 ft past the point. At this time the total bed separation was maximum. Overall the strata directly above the opening are subjected to tension in the vertical direction. But above a certain level, all strata move nearly simultaneously. There is also significant shear movement along some bedding planes. 1.3 ZONES OF MOVEMENT INTHEOVERBURDEN After the extraction of a longwall panel or room and pillar section of sufficient width the strata in the overburden are subjected to various degrees of movement from the bottom to the top. According to the movement characteristics, the damaged overburden can be divided into four zones (Fig. 1.6). Caved Zone. After the extraction of coal, the immediate roof caves irregularly and fills up the void. The strata in this zone not only lose their continuity completely, they also lose their stratified

Jan 1, 1992

By Thomas E. Browne

The Electric Power Research Institute Energy Supply Program is developing coal supply forecasts. By coal supply forecasts I mean the future relationships between coal price and production levels. A representative supply curve is shown in Figure l. It will make the rest of my presentation clearer if at this point I give you some background about EPRI and the Energy Supply Program. EPRI is funded by the nation's electric utility industry to develop new technologies for the production and conservation of electricity. Most of the research is conducted by outside contractors. We participate in joint projects with the Department of Energy and other organizations. The Energy Supply Program is one of three "economic" programs. The other two are Energy Demand and Conservation, and Systems. The Systems Program's main function is to provide the means to integrate the output of the other two Programs. The primary function of the "economic" programs is to provide information pertaining to future energy prices, availability, requirements, etc. to the EPRI personnel planning R&D programs. A secondary role that is becoming more and more important is providing information of direct use to the electric utility industry. I should note that the results of all EPRI's contract work are available to the general public. A "Coal Supply Analysis System" is being developed for EPRI by Charles River Associates. To the extent possible it will draw upon existing work, but much additional work needs to be done in the area of coal supply economics.

Jan 1, 1977

By J-M M. Rendu

Geostatistics are increasingly recognized as powerful tools for reserve evaluation and grade control. It is also generally accepted that geologic input is required for the results of a geostatistical study to be meaningful. The role of the geologist is significant at all stages, whether it be to oversee the creation of the database, to assist in the statistical and semivariogran studies, to ensure appropriate block evaluation and kriging, or to review the results obtained. However the link between geology and geostatistics is often complex and not well understood. A presentation is made of the areas where geologic input is needed, and of the tools which the geologist should have at his disposition to ensure that his role is fulfilled. Organization of a Geostatisticel Study To be acceptable as the basis for financial investment, a geostatistical reserve estimation should be composed of four phases, the last phase being optional: Phase 1 - Background discussions, Phase 2 - Computation, Phase 3 - Reserve reporting, Phase 4 - Reserve audit. During the background discussions (Phase 1) the role of the geologist is primordial, as it is then that most decisions will be made concerning the geologic factors to be included in the geostatistical model.

Jan 1, 1985

By D. C. Torre

The need for a successful, efficient and reliable continuous face- haulage .system for use with continuous mining machines has been apparent since the first continuous miners were introduced some twenty-five years ago. Some thought and effort were given to the development of continuous haulage over the last twenty years or so. Some initial success was experienced in this area by The Long Company with a system involving bridge conveyors, as far back as the late 1950's. However, extensive success with continuous face haulage for use behind modern, high product¬ion, continuous miners escaped the industry until the early 1970's. It is true that continuous face haulage was successfully introduced into low seam mining, generally in seams ranging under 1.07 m (42") in height, in the 1960's, for use with relatively low production, low seam, auger type continuous miners. However, those available systems which have achieved success in the lower seams, where relatively low capacities are adequate, are not directly applicable for use with today's medium and high seam miners with their standard production capacity of 9.1 - 10.9 mtpm (10 - 12 tpm). Since the late 1960's and early 1970's, several machinery manufacturers, and at least two major coal companies, have devoted considerable expense, time and effort to developing truly successful continuous face haulage for use in coal seams ranging upward from 1.2 m (4') in height. Continuous face haulage equipment has been developed in the last five years employing a variety of transportation media and design concepts, including some which have received wide publicity such as those employing hydraulics, and pleated belt. The truth of the matter is, however, that the only universally applicable and widely successful continuous face haulage equipment introduced to date has been that employing a combination of bridge conveyors and mobile, self-propelled, bridge carriers, connected at the inby end to the tail of the high production continuous miner, and at the outby end attached directly to the panel conveyor (Fig. 1 & Fig. 2) or, on occasion, to a short, self-contained conveyor which, in turn, empties directly onto the panel conveyor. While these bridge type continuous haulage systems are available with both belt and chain conveyor components, the most widely applied and successful system in seams above 1.07 m (42") in height is that offered by The Long-Airdox Company. The standard unit for use with 9.1 - 10.9 mtpm (10 - 12 tpm) continuous miners is normally equipped with belt. However, they have built, and successfully applied, versions of this system employing chain conveyors, (Fig. 3), mainly for operators having extremely difficult mining conditions, including excessive amounts of water

Jan 1, 1977

By Frank G. Horino, V. E. Hooker

INTRODUCTION The determination and use of mechanical properties of rock in engineering and rock mechanics are rapidly developing. Many of these properties are determined on intact rock specimens; thus, their application and repre¬sentation of rock mass properties may be limited. How¬ever, relative information often provides useful guidance in the solution to mine design and stability problems. Summarized in this chapter are some of the stan¬dardized techniques and procedures currently used to obtain these mechanical properties. Typical applications of the use of these properties are also presented. Stan¬dardized techniques include those advanced by the American Society of Testing and Materials (ASTM), International Society of Rock Mechanics (ISRM), US Bureau of Mines (USBM), Canadian Dept. of Mines and Technical Surveys, South African Institute of Min¬ing and Metallurgy, and other individual investigators. Information on the mechanical properties of rock and the behavior of the rock under a given system of stresses represents a necessary part of the information for rational engineering design for any given mining op¬eration. The mining method, the type and extent of sup¬port, the extraction ratio, the overall dimensions of the mine, and the orientation of the rooms and pillars are all decisions that are influenced by the mechanical prop¬erties of the ore, roof, and floor material under various stress systems and the magnitude and direction of the in situ stresses (Hooker, Bickel, and Aggson, 1972). Initial mechanical property information regarding a structure or mine property is generally obtained by two basic techniques: (1) static and dynamic property tests are conducted on intact and fractured rock specimens of exploratory drill core, and (2) dynamic properties are obtained by borehole logging techniques. When mining access becomes available, and as the mining horizon is expanded, additional information can be ob¬tained to verify preliminary mine design values. This chapter presents some of the standardized tech¬niques and equipment currently used in obtaining me¬chanical property data in the laboratory. The properties considered are: (1) uniaxial compressive strength of intact rock core specimens, (2) uniaxial compressive strength of rock cores containing planes of weakness, (3) triaxial compressive strength of intact rock core specimens, (4) triaxial compressive strength of cores with a plane of weakness, (5) Young's modulus, (6) Poisson's ratio, (7) density or apparent specific gravity, (8) modulus of rupture, (9) indirect tensile strength, and (10) creep characteristics. Where possible, an at¬tempt will be made to evaluate each property measure¬ment in relation to the problems of rock mechanics and application of results. TEST SPECIMENS The selection and care of drill core for laboratory testing require some consideration. It is recognized that laboratory-determined properties are not necessarily rep¬resentative of an in situ rock mass property. However, relative information between beds or zones of interest is still valuable information in selecting mining horizons and preliminary design criteria. To provide statistical data the number of drill core samples selected to repre¬sent each of the areas of interest should be from a mini¬mum of three to a maximum of ten test specimens. A judgment must also be made on site as to whether the recovered drill core should be wrapped and sealed in plastic to preserve moisture. On the one hand investiga¬tions of air-dried and saturated specimens have shown that moisture significantly affects the elastic properties and strengths of many rock materials (Obert, Windes, and Duvall, 1946; Colback and Wiid, 1965); on the other hand it is apparent that most core drilling is done with water which may saturate the specimen to a greater extent than in the in-situ condition. Whether or not the decision is made to retain the moisture, the core should be delivered to the laboratory as soon as possible after recovery for subsequent specimen preparation and testing. Specifications Shape: The shape of the specimens influences lab¬oratory testing in two ways: (1) time and cost of sam¬ple preparation and (2) strength of the material. Cy¬lindrical specimens of drill core are by far the least time-consuming to prepare for static or dynamic labora¬tory testing. In addition, the cylindrical shape lends it¬self to a more uniform stress distribution throughout the sample than other shapes, such as rectangles and hexa¬gons. The compressive strengths of various shapes have been studied (Grosvenor, 1963, and Price, 1960), and results indicate that the cylindrical specimens usually provide the highest strength for a given height-diameter ratio. However, reduction in strength from a cylindrical shape to a rectangular in situ pillar is not regarded as significant in relation to other considerations such as planes of weakness in a pillar or safety factors in the design process. Length-Diameter Ratio: The length-to-diameter ra¬tio, LID, has a significant effect on the compressive strength. Various recommendations have been made to use standard LID ratios ranging from 2 to 2.5 to 3 (ASTM, 1975c; ISRM, 1972). However, past work by others such as Obert, Windes, and Duvall (1946) has shown that excellent results can be obtained using LID ratios from 2 > (LID) > >/s. In selecting an LID ratio for testing, one should keep in mind the amount of material available for testing. In many instances, this may be limited. Thus, a shorter specimen such as 1: 1 LID may be necessary to provide enough test data for statistical analysis of results. Sec¬ond, it may be desirable to obtain elastic constants dur¬ing the test. This generally requires instrumentation such as linear variable differential transformers (LVDTs) or strain gages near the center of the specimen. In this case, an LID of 2.5 or 3 is desirable so that the instru

Jan 1, 1982

By T. P. Meloy

Like our wives, the comminution of brittle solids is easy to recognize but difficult to describe. However, describing it we must. Describing the fracture mechanism of brittle solids permits the eventual mathematical quantification of the results, process optimization and control strategies. This paper presents a physical model of brittle solid comminution that explains why the slope of the log-log size distribution curves have a slope of one or less in the smaller particle sizes. To develop the hypothesis, modifications of the Gaudin Meloy physical model are made. In brittle solids, comminution takes place by cracks propagating through the material. These cracks may bifurcate as they move. Cracks may terminate either on the original surface of the material, or on a secondary crack within the material formed earlier in the comminution process.

Jan 1, 1994

By T. Zoltai

Asbestos is a generic term for industrially useful fibers recovered from exploitable deposits of the asbestiform varieties of five silicate minerals: chrysotile, anthophyl-lite, actinolite-tremolite, cummingtorute-grunerite, and rie-beekite. Asbestos and asbestiform fibers of some other minerals are known to be carcinogenic. Minerals crystallized in the asbestiform habit-are characterized by unique properties (morphology, strength and flexibility, physical and chemical durability, and defect-free surface structures). The key to the unusual properties, and health effects, of asbestiform fibers appears to be in their unusual surface structure. In contrast, non-asbestiform crystals of the same minerals, including the five so called asbestos minerals, do not have these properties and are not carcinogenic.

Jan 1, 1989

By Patricio Zarate

Due to the intrinsic variability in orebodies, similar metallurgical behaviour of all material sent to the concentrator is expected to have some variability. Still some specific ores have anomalous performance which can severely impact on the concentrator’s production of metal, in this case copper. The study first will show the mineralogical characteristics of “usual ore” and of this specific “abnormal ore” which will have to be processed by the concentrator in the future. After that base conditions on both ores and flotation test procedure will be explained. Later the different alternatives tested regarding collectors and modifiers results will be compared. Finally, after discussing the different results with different reagents and considering operational variables of the concentrator, Floticor PD 702 is recommended to obtain the best copper recovery without increase in excess mass recovery.

Feb 1, 2025

By W. K. Utt

This report discusses U.S. Bureau of Mines' research involving modem vehicle position determination techniques in surface mining. The applications would include haul trucks, power shovels, and drilling rigs. Accurate position determination of these vehicles in near real-time is expected to result in safer and more efficient mining operations. The requirements for a vehicle position monitoring and warning system were determined. After screening several concepts, the global positioning system (GPS) and a radar positioning system were selected for further evaluation. A state-of-the-art GPS receiver was taken to a surface mine to assess the feasibility of receiving the satellite signals. A state-of-the-art radar positioning system was evaluated with regard to accuracy.

Jan 1, 1995

By Charles E. Scott

The molybdenum industry has been in a depression since the beginning of the decade. Real molybdenum prices are significantly reduced from 1979-1980 levels when there was a perceived capacity shortage and continued expectations of burgeoning demand. Primary capacity brought on stream since then has been unable to operate profitably and indeed, in light of high inventories and reduced demand, has been largely unnecessary. Structural changes in the molybdenum industry will be needed in the future to weed out superfluous capacity and to expand demand through new product applications.

Jan 1, 1986

Machine breakdowns and maintenance delays are large contributors to lost production and increased operating costs in the mining industry. The US Bureau of Mines is investigating the applications of expert system technology to diagnostic and predictive maintenance for mining equipment. This will help increase mining productivity and decrease operating costs by reducing the frequency of equipment failures and repair times. The Bureau has begun this effort by applying expert system techniques to the diagnosis of hydraulic system problems on a continuous mining machine. This paper discusses the development of the hydraulic diagnostic expert system.

Jan 1, 1992

By Tuncel M. Yegulalp

In this paper we revisit the well known "Ultimate Pit Limit Problem" and suggest an alternative approach for fast and efficient solution. We follow the approach originally proposed by Johnson and Barnes (1988), but utilize the "Excess Scaling Algorithm" in solving the equivalent maximum flow problem. This algorithm provides significant increases in speed and reduction in memory requirements over the well known Labeling Algorithm of Ford and Fulkerson (1956), enabling one to solve large problems with relatively small computers. We present results of a series: of benchmark tests using this approach against the well known Lerchs Grossman algorithm as well as a brief description of the new algorithm.

Jan 1, 1992

By A. M. Marabini, G. Belardi, E. Spaziani

The objective of this research was to investigate the possible application of integrated circuits to the concentration of talc from the Valmalenco mine, located in northern Italy. The investigation was aimed at producing a high-grade product (low in impurities such as iron and carbonate) that would be suitable for industrial applications. Because of the presence of accessory minerals, the talc concentrate from this ore is not, at the present time, suitable for industrial applications.

Jan 1, 1996