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By D. R. Tesarik, R. W. McKibbin

Material properties affecting slope stability were measured in a large 50-year-old, partially consolidated rockdump located in an active open-pit mine. Field tests included single-ring infiltration and density. In addition, a nuclear depth-moisture gauge was used to measure water content in six stainless-steel-cased drill holes on the crest and an upper bench of the rock dump. Precipitation, evaporation, wind speed and direction, and temperature data were collected at a weather station installed on the dump’s crest. Laboratory tests included particle-size distribution, specific gravity, Atterberg limits, and water content. By measuring material properties of a rock dump presumed to be stable, the safety of miners working on or at the toe of old rock dumps constructed of similar material and located in a similar climate can be assessed.

Jan 1, 1999

By G. W. Eiger

Bureau of Mines laboratory tests show that ilmenite concentrates with titanium dioxide contents as low as 33 wt--pct can be beneficiated to produce a crystalline substitute for natural rutile suitable for direct chlorination. The synthesized product prepared by a three-step procedure was comparable with natural rutile in titanium content. Reduction smelting of ilmenite removes iron as a pig iron byproduct, and the remaining slag phase is greatly enriched in titanium. Oxidation of the slag in the presence of phosphorus¬bearing slag additive converts the titanium-bearing phases to discrete rutile crystals and a glass phase containing most of the associated impurities. Rutile is separated from the glassy impurities by attrition scrubbing ground slag in dilute phosphoric or sulfuric acid solutions. Tests made in a laboratory fluid-bed reactor show that synthetic rutile responds to chlorination at the same rate as natural rutile and that more than 95 pct of the titanium contained in the feedstock was extracted as titanium tetrachloride.

Jan 1, 1974

By Edward L. McHugh, Jami M. Girard

While less than 1% of reported accidents are associated with slope stability problems, slope failures were responsible for about 15% of all U.S. surface mine fatalities between 1995 and 2001. Small rockfalls, which may involve hand-sized rocks weighing only a few kilograms, can cause fatal injuries to workers away from the protection of large machinery. Massive highwall failures containing a million cubic meters of material or more can be fatal even for heavy equipment operators. As part of an ongoing study at the Spokane Research Laboratory of the National Institute for Occupational Safety and Health, several remote-sensing technologies are being evaluated as tools to monitor slopes for hazards and to assess slope stability. Field tests of a hyper spectral imager were conducted to assess its value for improving geologic maps of potentially unstable alteration zones on mine slopes. An interferometric radar device capable of detecting very small displacements on slopes awaits final assembly and field testing. Computerized monitoring methods using images from digital and video cameras are being assessed for application to mine slope surveillance. These and other techniques will eventually provide new tools to augment current methods for monitoring ground control hazards in mines.

By Joan C. Biordi

[Composition profiles for atomic, radical, and stable species, as well as temperature and area expansion ratio profiles, have been determined for a nearly stoichiometric CH4-02-Ar flame and one to which 0.3 percent CF5Br inhibitor had been added. Net reaction rate profiles were calculated for all the observed species. For the normal flame, these and the mole fraction profiles gave rate coefficient information about the elementary reactions in the methane flame, for example, CH. + 0 4- H2CO 0 H, k4 = 1.05 X 1014CM3 mole-1sec-1 for 1,550§T 0,725' K; CO + OH 4 C02 + H, k6= 4.7 X 1011 for 1,350{T X1,750° K. Comparison between the inhibited and normal flame showed that A] and [CH3] were significantly reduced at the lower temperatures in the inhibited flame even though in the hot gas region the [E] [OH!, and [0] were essentially the same in both flames. The CF3Br disappears very early in the flame, relative to the fuel, and the reaction primarily responsible for its disappearance is H v CF3Br 4 HBr + CF3, where k7 is found to be 2.2 X 1014 exp(-9460/RT) for 700-1,550° K. Reaction of the inhibitor with methyl radicals provides for the relatively small amounts of CH3Br observed, but CH3 + Br2 - CHs Br + Br must also occur. The HBr formed reacts rapidly with H atoms to form H2 and Br, but the reaction is soon "balanced" in the flame as may be demonstrated by calculation of the equilibrium constant at various temperatures. The fluoro-carbon fragment produced in reaction (7) also reacts rapidly, in part with methyl radicals to give the observed elimination product CH2CF2 Later in the flame, above -1,400' K, F2CO is formed from the reaction CH2CF2 + 0 P F2CO + CH2 and k., -1.5 X 1013 at 1,600° K. The rather slow decay of carbonyl fluoride is attributed to reaction with H atoms according to the sequence F2CO I H - HF + FCO and FCO + H HF + CO.]

Jan 1, 1975

By Harvard Eng

An industrial survey was made for the Missouri Basin Field Committee of the Department of the Interior to determine the type and quantity of water-borne mineral wastes and the water requirements of metal-processing industries in the basin in Kansas and western Missouri. Data are presented on metal-processing plants visited through June 30, 1959. Most of the plants were in the greater Kansas City area. This report is to be supplemented with a similar publication on other industries, including those producing sand and gravel, coal, petroleum, and chemicals. In 1958, 19 million pounds of waste materials was discharged from metal processing industries. Approximately 14 million pounds of this waste was in the form of ferrous sulfate, and 2 million pounds was ferric phosphate. Galvanizing, other hot dipping, and automotive industries contributed the greater part of iron compounds wasted; some of this waste occurred as sludge that was not discharged to water courses but was hauled to refuse dumps, podium was present in nearly all industrial wastes, and the total of more than 900,000 pounds constituted one of the larger material losses. Although small in quantity, electroplating and allied coating wastes contained metals that were expensive when originally purchased.

Jan 1, 1961

By I. F. Avery

The productive capacity of natural gasoline and cycling plants in the United States on January 1, 1956, was 46.4 million gallons per day, according to the Bureau of Mines r biennial survey. This represented an increase of 7.5 million gallons per day during the 2-year period since the last survey. In the previous 2-year period, capacity had increased 4.8 million gallons per day. The total number of plants increased from 535 on January 1, 1954, to 600 on January 1, 1956. The number of operating plants increased from 521 to 571. The major increases in capacity took place in Texas, Louisiana, New Mexico, and Oklahoma, with the absorption-type plants continuing to account for most of the new construction. All types of plants showed construction gains with the exception of charcoal which showed no change.

Jan 1, 1957

By Gregory M. Molinda

The U.S. underground coal miner faces a continuing hazard from the fall of roof. At the root of many injuries and fatalities are weak or defective roof strata. Throughout mining history, millions of miles of entry have provided exposure of every conceivable geologic roof hazard. This report describes the geologic origin, association, and potential danger from the most common hazards. Discussions of weak rock include drawrock, rider coals, head coal, stackrock, and stream valley effects. Discontinuities, or roof defects, are described including, clay veins, slickensides, joints, and paleochannels. A number of examples from U.S. coalfields are used to document geologic structure and associated hazards. Roof fall analysis is a methodology used by NIOSH for hazard recognition and prevention; its application and benefit to the industry are discussed.

Jan 10, 2003

By Eugene T. Sheridan

The annual survey of the fuel-briquet industry by the Bureau of Mines, United States Department of the Interior, revealed that 21 fuel-briquet plants in 11 States were in operation and reported production in 1956. The demand for fuel briquets in the United States has gradually declined since 1948, and the number of operating plants has decreased from 36 to 21. Also, the annual productive capacity of the industry has decreased, although not a proportionate amount because most of the plants that were abandoned were small and in most instances had a designed capacity of less than 25,000 tons per year. In 1956 the 21 active plants had an annual capacity of 3.7 million tons (table 2), which was only 20 percent less than the total capacity of the 36 plants in 1948. This indicates that most of the larger plants have remained active, even though operating at reduced rates, and the average capacity of the industry has increased from 130,000 tons per plant in 1948 to 177,000 tons in 1956. Since 1948 operating rates have decreased from 67 percent to 41 percent of annual capacity.

Jan 1, 1958

By T. O. Llewellyn

The Bureau of Mines, as part of its goal to develop technology that can conserve domestic mineral resources investigated beneficiation methods for recovering phosphate minerals from two Southern Florida deposits with high magnesium content. Chemical analyses of sized fractions indicated that no coarse pebble concentrate could be produced from either deposit. Conventional fatty acid-fuel oil batch flotation tests on two phosphate samples, analyzing 6.8 and 8.5 pct P205, yielded concentrates containing 28.97 to 29.80 pct P20S at recoveries ranging from 72.4 to 76.9 pct of the flotation feed. The MgO content ranged from 1.20 to 1.36 pct. Bubbling S02 gas through an aqueous concentrate slurry, further reduced the MgO content of these phosphate concentrates to 0.5 and 0.8 pct, but with a concomitant loss in the phosphate recovery. Settling characteristics of slimes from both phosphate samples showed a slow settling rate, which is a typical characteristic of other slimes in the region.

Jan 1, 1981

By John J. Sammarco

Mining has one of the highest annual average fatality rates among major US industries. Health and safety dangers have been inherent to mining since the early days of picks and shovels. Even though miners' health and safety has improved over the years, mining is still one of the most dangerous occupations. Mining was traditionally a low tech industry. It is now driven by competitive pressures to go high-tech by using programmable electronics (PE) for machine control, atmospheric monitoring and material processing. The industry's experience with the functional safety of PE is limited compared with other industries. Thus, PE is an emerging technology for mining that can potentially create or worsen hazards. The US National Institute for Occupational Safety and Health (NIOSH), Pittsburgh Research Laboratory in Pittsburgh, PA is addressing the safety of this new technology. NIOSH has a proactive project to generate recommendations for addressing the functional safety of PE-based mining systems before the technology proliferates. The recommendations take the form of a safety framework encompassing the entire life cycle for a PE-based mining system.

By Michael G. Pojar, John K. Ahlness

"The copper industry has had a long and interesting history associated with leaching, involving vat, dump, heap, and in situ methods. The Bureau of Mines has also had an interest in copper leaching and has researched the techniques of solution recovery, particularly concentrating on in situ methods. In situ mining is a relatively low cost method and has been proven commercially successful by nine operations. It has been most commonly used for final recovery in old workings at the conclusion of conventional mining operations.This report brings together information about 10 commercial in situ operations as well as 14 experimental projects. These 24 sites comprise most of the in situ copper mining activities that have taken place in the United States. Background information, geology, ore preparation, solution application, and recovery and processing are provided for each operation. Production data and tables summarizing the engineering statistics for each operation and an extensive in situ mining bibliography are included. INTRODUCTIONThe purpose of this Bureau of Mines report is to summarize the documented in situ copper leach mining activities that have taken place in the United States. It was prepared to provide a source of engineering and operating data upon which future in situ copper mining activities can be based. Twenty-four different sites are discussed. These are all of the sites that are known to the authors from the extensive in situ copper leach mining bibliography developed at the Bureau's Twin Cities (MN) Research Center (Appendix B).In situ copper leach mining activities fall into the following categories:1. Commercial operations with ore body preparation.2. Commercial operations in old mine workings.3. Experimental programs.Ore body preparation includes such activities as blasting, block caving, and hydraulic fracturing (hydrofracing) specifically to fragment the ore body. Operations in old mine workings include those done in open pits, worked out block caved areas, and backfilled stopes where leaching is done as an afterthought to conventional mining. Experimental programs are small-scale tests of the feasibility of a commercial operation. Some of these programs did not progress to the point where leach solutions were applied to the ore.Each operation (locations shown in figure 1) will be discussed in terms of background information, geology, ore preparation, solution application, recovery and processing, and engineering and production data. Information was collected from published reports, communications with company personnel, and site visits."

Jan 1, 1982

By W. H. Young

The thickness of the coal seam is one of the more important items influencing the productivity and general overall efficiency of operation 1 of a coal mine. The average thickness of coal seams mined in the United States decreased 2 inches between 1950 and 1955 to an average of 62 inches, which also is 3 inches less than the 65 inches in 1945. The decrease in thickness is due partly to the increased production from strip mines with thinner seams and partly to the increased percentage of total production at underground mines in certain States having thinner seams. The thickness of seams at auger mines was measured for the first time in 1955 and was 53 inches. Though bituminous coal and lignite is mined today from seams of exceedingly variable thickness -less than 2 feet to ore than 50 feet -a large majority of the production, 65 percent in 1955, came from seams 3 to 6 feet thick. This is not surprising since the coal seam of moderate thickness presents the least mining difficulties.

Jan 1, 1957

By H. L. Riley, B. W. Gandrud

"INTRODUCTION The Bureau of Mines, United States Department of the Interior, with the cooperation of the University of Alabama and coal producers, has made a series of washability studies to obtain technically and commercially important data regarding the productive coal beds of the State. Based on the system followed in these studies, this report consists principally of screen-sizing and float-and-sink data on the Clark and Gholson beds at Boothton, Ala. A previous washability study 4/ has been made on the Clark bed at Marvel, Ala.SAMPLING AT MINESThe coal from the Gholson and the Clark beds was sampled at the tipple which receives the output,the Boothten No. 3 and No. 4 mines. These two mines are on the Gholson and Clark beds, respectively, but the coal. from both mines is mixed on the tipple and washed in the same washery. The samples taken were minus 4-inch raw coal. The plus 4-inch raw coal is hand¬picked and not washed. AS the 4-inch screen had a capacity in excess of the output of the mines, it was stopped at the end of each mine-car trip. By Observing from which mine incline each trip was hoisted, a separate sample from each bed was cut from the discharge stream of an endless belt which conveys the through 4-inch raw coal to the bin ahead of the washer. The output from each bed was sampled uniformly over the period of a working shift. The two washability samples weighed approximately 3,000 pounds each."

Dec 1, 1947

By R. R. Backer

The Bureau of Mines, in cooperation with Washington Irrigation and Development Co. (WIDCO), evaluated the dewatering of fine coal-refuse slurry using chemicals and various types and combinations of drains in natural earth impoundments. About 40 lb of lime and 2 lb of polymer per ton of dry solids were mixed with the underflow from the preparation-plant thickener. Clear, free water was immediately liberated from the slurry. The percent solids by weight of the slurry increased from about 17 to 30 immediately, and after 5 days increased to 45 wt-pct. This indicates a threefold reduction in total volume after 5 days from 172 cu ft/ton solids originally to 56 cu ft/ton solids.

Jan 1, 1981

By Karl C. Dean

The Bureau of Mines studied the precipitation of copper from dilute acidic sulfate solutions using shredded automobile scrap in a rotary drum. The tests established that the scrap in the rotary drum precipitated copper faster and more completely than shredded, detinned, tinplate scrap in a launder, with comparable iron consumption. Although power is required to rotate the drum, the faster precipitation and continuous recovery of cement copper are advantages of the tumbler method. Cost evaluation indicates that the rotary drum method, which uses relatively low-cost and readily available shredded automobile scrap, is competitive with the launder method, which uses relatively high-cost shredded tinplate scrap or detinned shredded cans.

Jan 1, 1968

This report summarizes the results of several field tests conducted by Collins for the Bureau of Mines under contract SO122076 and contract H0232056. The field projects represent a part of Collins continuing effort on a broad study of coal mine communications. These experiments analyze the propagation characteristics of radio signals in a working coal mine environment. The intent of the program is to provide basic propagation loss characteristics by which we may evaluate various approaches to an integrated mine communications system. The goal of such a system is to satisfy both operational and emergency communications needs in the mine. Field testing was conducted during the last week of November 1972, at a coal mine operated by Inland Steel Company, near Sesser, Ill, and in the Robena No. 1 Mine, Waynesburg, Pa, on 12 March 1974. The tests included uhf propagation losses and noise (200 MHz to 1 GHz) in the Sesser, I11 mine location and vlf propagation loss and noise (1 to 200 kHz) in both of the above locations. The basic experimental parameters were frequency, polarization, antenna type/orientation, and distance. The following reports treat the various aspects of the test programs in detail. Since different antenna configurations were used in the two mines, each report contains its own experimental predictions drawn from its corresponding test activities. A test equipment list and block diagrams of typical experimental setups for the tests of section 2 are contained in appendix A.

Jan 1, 1974

By Uldis Jansons

This report evaluates the economic Mining feasibility of four selected mineral deposit types that can be anticipated to be found in (d)(2) lands in the western Brooks Range, Alaska. The major assumption made for this study is the availability of a rail system for transporting equipment and concentrates for the base metal mines. Commodity prices used are those in existence in July 1976. Mine models were established based on known deposits and then evaluated. A lead-zinc-silver-barite deposit for which grades were available but ore reserves were assumed would require approximately 25% higher ore grades or higher metal prices to yield a 15% discounted cash-flow rate of return an the investment if mined at a rate of 10,000 tons per day open pit methods. Large copper-zinc mineral deposits could be mined economically at rates either of 5,000 tons per day underground or 10,000 tons per day by open pit methods. A high-grade copper deposit require a copper price of over one dollar ryes pound to yield an economic profit to be mined at 2,000 tons per day by underground methods. A small high-grade gold underground mine would require a gold price of about $295 per ounce to yield a 20% discounted cash-flow rate of return on investment if all new mine equipment were used and if existing labor rates were paid: Enterprising mining groups operating as partnerships might be able to operate this mine profitably at a gold price of about $152 per ounce, perhaps slightly less.

Jan 1, 2012

By Louis A. Panek

This report deals with the combined friction and suspension effects in the reinforcement of horizontally bedded mine roof by bolting. Other reports have treated these two effects separately. Friction effect refers to the reduction of bending in a laminated roof due to the clamping action of tensioned bolts. If the beds tend to have unequal deflections, the bolts oppose this tendency, thus creating a suspension effect, which consists of a transfer of part of the weight of weaker or thinner beds to the flexurally more rigid beds within the bolted unit. The combined effects of friction and suspension on the bending of the beds, whether an increase or a decrease, are difficult to generalize because they depend on the bedding sequence and the bolt pre-tension. Some analysis of the bedding sequence usually is required to determine the critical bed within a set of bolted beds. However, if the suspension effect is substantial and if the bolts are pre-tensioned, then the greatest bending stress is likely to occur in the thickest bed. Findings are based partly on theory and partly on tests of mine-roof models in a centrifuge, which were found to verify the theory. Mathematical formulas are obtained for the change in maximum roof bending stress and for the change of roof deflection due to the combined friction and suspension effects. The maximum bending stress is needed for designing a roof-bolting system and the deflection for evaluating its effectiveness.

Jan 1, 1962

By Robert K. Koch

Laboratory-scale studies on defluorination of natural and synthetic fluorspars by pyrohydrolysis at 1,500° C. were made to determine the feasibility of developing a process for producing hydrogen fluoride from siliceous fluorspars. It was found that the reaction (at 1,500° C.) is not promising for process development, even though a silica-free offgas can be generated. The primary difficulty is the low hydrogen fluoride content of the offgases; an HF content of 6 percent was the best achieved. Previous research by the Bureau of Mines has demonstrated that offgases averaging 27 percent hydrogen fluoride can be produced under carefully controlled conditions at 1,280° C. The lower temperature process is a better method for production of HF. However, charges in which the CaF to Si02 mole ratio varied from 0.4 to 2 could be processed with similar results. Also, combined alumina and ferric oxide contents up to 25 mole-percent did not noticeably alter the rate of defluorination. Therefore, virtually any natural siliceous fluorspar would respond in essentially the same manner when exposed to water vapor at 1,500° C.

Jan 1, 1962

By C. W. Smith

Research was conducted by the Bureau of Mines to determine the effectiveness of various types of crushers for use with the Bureau-developed pneumatic concentration technique for mica. Three types of crushers were investigated in this study, a roll crusher, a jaw crusher, and a hammer mill. All the mica concentrates produced contained over 82 percent mica, but recoveries varied considerably. The roll crusher gave the lowest recovery of 46 percent of the mica followed by the jaw crusher with a recovery of 50 percent of the mica. The hammer mill proved the most effective, producing four concentrates with recoveries of at least 70 percent.

Jan 1, 1981