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By T. E. Sterner

A laboratory investigation was conducted to determine the applicability of the thermal oil-recovery method as a means of increasing ultimate oil recovery in the Appalachian region. The primary objective was to establish self-sustained combustion in consolidated and unconsolidated sand test assemblies saturated with various amounts of Pennsylvania Grade crude oil and synthetic brine. Tests using blocks of sandstone were unsuccessful because of sealing problems and fracturing of the rock due to thermal stresses. A combustion wave of about 700° F was propagated through the sand-oil-water packs in the linear tube tests. The high air flux used resulted in high oil consumption and rapid movement of the combustion wave through the tubes. Analytical tests on the sand pack showed that sufficient fuel was deposited during the linear tube experiments for a self-sustained combustion wave.

Jan 1, 1967

By A. J. Miscoe, W. M. McKewan

A unique machine for inducing bending fatigue in wire ropes was built at the U.S. Bureau of Mimes' Pittsburgh Research Center. This machine can produce nine levels of degradation through repetitive cycling of long samples of wire ropes of the types used in the mining industry. This report provides an analysis of the results of the first wire rope to be fatigued on this machine. These results have indicated that as a wire rope accumulates bending cycles, its strength first decreases because of wear and then in- creases due to cold working of the wires. As the amount of cold working increases, more and more of the embrittled wires break and the strength of the rope decreases rapidly. In this test, application of several of the regulatory retirement criteria to various sections of the rope suggests that there may be some inconsistency among them.

Jan 1, 1993

By Ned Baker

Details are given of a fuel-induction system for a hydrogen, fueled engine. The induction system was designed to preclude backfire otherwise common in hydrogen fueled engines. System features and test data are presented and discussed in the context of the requirements peculiar to use in an underground mine environment. The fuel induction-system when tested in conjunction with a modified Caterpillar 3304 engine provided backfire free engine operation and satisfactory engine performance.

Jan 1, 1982

By A. T. Coumbe

During 1953-55, 31,105 miles of pipe for transporting crude petroleum and refined products was laid in the United States, a total slightly below that of the prior 3-year period, according to data compiled by the Bureau of Mines, United States Department of the Interior. The added mileage consisted of 22,571 miles of new pipe, representing about 2,163,807 tons of steel, plus 8,534 miles of secondhand pipe, equivalent to about 419,849 additional tons of steel. During the 3-year period, 13,069 miles of pipe was taken up. By January 1, 1956, the Nation had 188,540 miles of pipeline, an 11-percent gain over the 1952 total of 170,504 miles. The greater use of larger diameter pipelines indicated in recent surveys has continued, as the mileage for sizes 12 inches and over increased from 13 percent of the total mileage at the close of 1952 to 16 percent 3 years later. Crude-oil trunklines 12 inches and over in diameter gained 24 percent during 1953-55, and these larger lines represented 30 percent of the total mileage compared with 25 percent at the close of 1952. Furthermore, the mileage for crude-oil lines less than 12 inches in diameter declined 2 percent and made up only 70 percent of the 1955 total against 75 percent n the previous survey. Product Lines 8 inches and over increased 43 percent from 1952 through 1955 and composed 70 percent of the total mileage compared with 66 percent in 1952.

Jan 1, 1956

By J. G. Stuart, A. W. Cipkins, R. H. Thompson

Preliminary safe distances for blasting wiring from typical electromagnetic sources are presented. The distances are for underground coal mines and electromagnetic sources in the frequency range below 10 MHz. The distances are based on measurements performed at 3 locations in each of 3 coal mines. The measurement apparatus, measurement procedure, measurement results and the rationale used in the calculation of the safe distances are all described. The safe distances presented are dependent on the measurement results and their interpretation. More measurements are clearly necessary to establish the validity of the suggested distances.

Jan 1, 1985

Lower costs associated with hands-on training in donning and using a CSE ,self-contained self-rescuer (SCSR). Background In June 1987, the Mine Safety and Heath Administration (MSHA) issued an Emergency Temporary Standard (ETS) requiring that all persons entering underground coal mines must be given bands-on SCSR training. The ETS will shortly be replaced by a new standard requiring this hands-on training. Bureau of Mines research has shown that the most effective means of delivering this training is to have the trainee go through a six-step training procedure while in a kneeling position. Once the miner has looped the neckstrap over his or her head, he or she completes the donning sequence in this order: 1) activate oxygen; 2) insert mouthpiece; 3) put on nose clips; 4) put on goggles; 5) adjust neck and waist straps; and 6) replace miner's cap and prepare to move out. This procedure is repeated until the miner demonstrates that he or she is proficient at the task.

Jan 1, 1988

By E. J. White

Data from oil well logs and core analyses were used to quantitatively evaluate formation porosity and water saturation of heterogeneous carbonates and to study the pore geometry and performance characteristics of cores from these formations. Logs and cores from the Madison limestone from 233 wells in three fields in the Williston basin were used in the investigation. A comparison of log and core analyses shows that log porosities are consistently larger than core porosities, and the differences are correlatable with core porosity ranges. Water saturations from logs are usually smaller than the corresponding core saturations by summation of fluids, but the difference is attributable to mud filtrate invasion into the cores during drilling. Petrophysical studies show that the low permeabilities and high residual water saturations in these reservoir rocks are due to their fine crystalline texture and complex pore geometry (heterogeneity). It is concluded that both log and core analyses must be used for reliable evaluation of these carbonate formations.

Jan 1, 1971

By G. M. Molinda, J. P. Ulery

Since 1970, abnormally high methane emissions have been observed in eastern sections of the Federal No. 2 Mine operating in the Pittsburgh Coalbed in northern West Virginia. The nature and volume of these emissions indicate a source other than the coalbed. An investigation was undertaken to delineate probable methane sources. Gas well information and stratigraphic correlation of corehole data indicates that a sandstone directly above the main coalbed is an additional source of methane emissions. The sandstone body appears to be an ancient stream channel, infilled with fine-grained sand and organic matter. Coalification of the organic matter produced methane as a byproduct. Additional methane presumably migrated into the sandstone unit from the subjacent coalbeds. The methane became trapped in the porous clasticunit by the surrounding impermeable strata. Mine development near this unit creates a significant pressure gradient, triggering abnormal methane emissions into the mine along naturally occurring fractures. Pre-development methane drainage of the sandstone could be accomplished by completing vertical boreholes along its trend.

Jan 1, 1984

By Donald P. Mickelsen

The raining and quarrying trends data shown in this report were reported to the U.S. Bureau of Mines (USBM) by mining and quarrying companies on the Mine Information Supplement survey, as well as by mineral producers on 58 annual USBM production and consumption surveys covering 59 nonfuel mineral commodities. The 1993 data presented in the tables of this chapter, were processed for the first time using the USBM Automated Mineral Information System. Domestic mining of nonfuel mineral materials totaled 3.7 billion metric tons in 1993, including 2.2 billion tons of crude ore nun& or quarried and 1.5 billion tons of mine waste and ore from development. This \vas an overall increase of 11% over that of 1992. These data do not include construction sand and gravel which until 1993 has been surveyed biennially. Starting in 1994, construction sand and gravel and the stones, which were also surveyed biennially, will be surveyed annually.

Jan 1, 1993

In this U.S. Bureau of Mines report, two drill-split mechanical excavation mining system concepts are described, casted, and cost compared with conventional drill-blast mining systems. The concepts for the two drill-split mechanical excavation mining systems presented are (1) a method for steeply dipping narrow veins 0.6 to 1.2 m wide and (2) a longwall method for fiat-lying tabular ores. In the proposed systems, 70 to 80 pet of the ore is disposed of underground as backfill. For sortable ores, face area crushing, sorting, and direct stowage of run-of-mine waste and low-grade ores are proposed. The underground sorting and stowing of waste reduce materials handling, processing, and waste disposal costs by reducing the quality of ore transported to the surface by 70 to 80 pet (by volume). Because many of the tools and techniques proposed for use in these systems have little or no historical mining industry use, cost assignations are heavily influenced by engineering judgments and presented as prefeasibility-level cost estimates. These estimates, presented as discounted-cash-flow-rate-of-return (DCFROR) values, show the proposed systems to be at least as profitable as, and in some cases 50 pet more profitable than, conventional drill-blast mining.

Jan 1, 1994

By Graham W. Marks

This report is the fourth of a series concerning the development of spectrochemical methods for the analysis of ore samples and related Metallurgical materials by the total-energy method. In this report, the application of this method to the determination of thallium is discussed. This element lies in group III of the periodic tale; one other Clement of this group; indium, has already been investigated.

Jan 1, 1949

By D. Harrington

"Up to the last 10 or 15 years, responsibility for accidents in mines was generally, though not universally, placed upon the person or persons immediately involved. Present-day thought is rapidly veering toward holding the industry responsible for accidents, with ultimate payment of the bill by the consumer, if financial expenditure should be necessary. While it is undoubtedly correct in principle that responsibility for accidents in and around mines should be shouldered by the employer and while it is probable that as time passes the burden will become even more definitely than now that of the mining companies, on the other hand it is an undeniable fact that large numbers of accidents occur through acts of omission or of commission of some person, very frequently the victim of the accident being the offender.Out of about 3,000 persons killed every rear in our -mines and quarries, nearly half or approximately 1,500 of the fatalities are due to falls of overlying material or of persons, about 500 to haulage, about 350 to fires and explosions, about 125 to explosives, approximately the same number to electricity, and about 300 to various miscellaneous accidents.Generally the accidents from falls involve but one or two persons at a time, and the tendency is now as it always has been to charge the victim or possibly a coworker with the responsibility. Usually this sort of accident occurs at or near a working face (statistics indicate that about 80 per cent of fatalities from falls of roof or rib occur near the face), and as it is Generally agreed that it is the duty of the worker to keep his working place safe, the worker and especially the face worker is given the blame in case of accident."

Apr 1, 1930

By John Nagy

The effect of low percentages of methane (0 to 5 pet.) in an air atmosphere on spark initiated explosions of coal dust was investigated. With low dust concentrations (0.1 to 0.4 oz./cu, ft.), small amounts of methane increase the explosion pressure; with larger dust concentrations (0.5 to 2.0 oz./cu. ft.), small amounts of methane decrease the explosion pressure. The rate of pressure rise of a coal dust explosion increases with increase in methane concentration; about 3 to 4 pct. further addition of methane yields a decrease in the rate of pressure rise. Convection currents created by the air or gas blast used to disperse dust produce an increase in explosion pressure and rate of pressure rise. The minimum explosive concentration of the coal dust decreases about linearly as the concentration of methane increases from 0 to 5 pct. The explosion pressure increases linearly as the initial pressure within the chamber increases. Similarities between coal dust and methane explosions are indicated, INTRODUCTION During coal mining, methane occluded in coal or adjoining strata is liberated into the mine atmosphere,4 Normally, the methane is diluted and swept away by the ventilating air. Under some conditions, through accidental disruption of the ventilation system, sudden outburst, or because insufficient fresh air is supplied to a working place, low percentages of methane may be present during mining operations. If it is permitted to accumulate, methane itself presents an explosion hazard; its limits of explosibility arc approximately 5 and 15 pct. The effect of low percentages of methane on the explosibility of coal dust was studied to provide a broader understanding of the hazards of ignition and explosion and thus to facilitate development of control

Jan 1, 1961

By F. C. Gibson

An electrical method for the continuous determination of propagation velocities in opaque, confined explosives, and propellants is described. A resistance element imbedded in the reactive material generates electrical signals that are displayed oscillographically and indicate the position and progress of a reactive zone associated with progress of a reactive zone associated with burning or detonation. The method is applicable to both solids and liquids.

Jan 1, 1963

By Kenneth L. Rubow, Virgil A. Marple, Bruce K. Cantrell

"In a cooperative study by the University of Minnesota and the U.S. Bureau of Mines, diesel exhaust and mineral dust concentrations have been measured for aerosols generated in the laboratory and in five underground coal mines. Three mines utilized diesel-powered haulage equipment and two used all-electric equipment. Two source apportionment techniques have been applied to differentiate between the mineral dust and diesel exhaust aerosol concentrations measured. The first technique, using a micro-orifice, uniform deposit impactor (MOUDI) for size selective sampling, is based on modeling aerosol size and the premise that the diesel exhaust portion of the aerosol is predominantly submicron and the mineral dust portion is mostly greater that one micrometer in size. The second technique, Chemical Mass Balance (CMB) modeling, was used to referee the analysis of diesel exhaust and mineral dust aerosol concentrations from the size selective sampling results.The MOUDI size distribution data were modeled to obtain parameters describing the fine and course fractions of the sampled aerosol and to estimate the contributions to each mode from both mineral dust and diesel exhaust aerosol sources. The results showed the size distribution of the mixed aerosols exhibited two definite modes with the minimum between the modes occurring at 0.8 µm. Less than 5 pct of the coal mine diesel aerosol was found in the coarse size fraction. CMB analysis confirms the original premise for using aerosol size to separate diesel exhaust and mineral dust aerosol during sampling.Based on the results from the size selective sampling, a personal diesel aerosol sampler has been developed for measuring diesel aerosol mass concentrations. Th.is sampler uses an inertial impactor to size separate the respirable aerosol at 0.8 µm. All material less than 0.8 µm is collected on an after filter for subsequent gravimetric analysis."

Jan 1, 1990

By D. Harrington

"A tabulation is printed below showing the accident record of the thirty-two mines of the Anaconda Copper Mining Co., in Montana, for the years 1915, 1916, 1917, 1918, 1919 and 1920, with summation for the six years. The table gives the number of shifts worked by each mine with the serious plus fatal accident rate for each 10,000 shifts worked.It is seen that in a total of 16,460,767.5 shifts worked by employees of the company's mines during this six-year period, there were only 2,005 serious and fatal accidents and of these, 200 were fatalities, leaving 1,805 serious accidents (a serious accident being one in which 14 or more days' time are lost). This is at a rate of 9,120 shifts per serious accident and at a rate of 82,304 shifts per fatality. Hence, if 300 days be adopted as a year's work, there would be one fatality per 274.3 years of work in these mines and one serious accident for every 30.4 years work.According to Bureau of Mines Technical Paper 286, (Metal-mine accidents in U. S. in 1919, by W. W4 Adams), during the years 1916, 1917, 1918, and 1919, the metal mines of the United States worked a total of 210,548,640 shifts and had 39,705 serious accidents. During the same years, according to the following tabulation, the mines of the above-mentioned company worked 11,958,978 shifts and had a serious accident rate of 0.995 accidents per 10,000 shifts worked. If the metal mines of the United States, working these 210,548,640 shifts, had this company's serious accident rate of 0.995 serious accidents per 10,000 shifts, the total number of serious accidents would have been only 21,000 or the country's metal-mine serious accidents would have been reduced nearly one-half, and there would be 18,705 fewer serious accidents than actually occurred."

Jun 10, 1921

Papers presented at a U.S. Bureau of Mines (USBM) technology transfer seminar describe the causes of violent material failure in U.S. mines, measurement techniques for monitoring events that result in violent failure, and mitigation techniques for controlling failure. Specific factors contributing to violent failure are identified on the basis of geotechnical monitoring in 16 U.S. hard-rock and coal mines and on statistical analyses of 172 coal bump events. New monitoring and analysis techniques developed as tools for assessing violent failure; geotomographic methods that provide new capabilities for the study of material failure and stress changes over large areas; and seismic methods for determining source locations, calculating energy release, and determining source mechanisms are described. Fair correlations have been established among seismic parameters, elastic stresses, face support load, and violent events. USBM studies have identified the advantages using both yielding and stable pillars for coal bump control. A computer program has been developed as an aid for selecting room-and-pillar layouts. The practical aspects of implementing a destressing program is outlined for coal mines, while the importance of mine orientation and timely support installation in controlling buckling-type failure is identified for hard-rock mines.

Jan 1, 1995

This report presents test results, data evaluations, and typical system performance expectations based on wireless electromagnetic propagation measurements performed in the MAGMA Copper Company's San Manuel mine near Tuscon, Arizona during March 30 through April 2 1981. In particular, and at the Bureau's request, this work was coordinated with testing of MF prototype vehicular radios being developed by A.R.F. Products, Inc. with USBM support under Contract H0308004 as part of a system to be marketed by Mine Safety Appliance Co. This testing also employed a prototype vehicular antenna developed by Terry S. Cory, P.E. with USBM upport under previous Contract P0382223. The work was performed under USBM contract J0308012, Propagation of EM Signals in Underground Metal/Non-metal Mines. The authors wish to thank Mr. Hank Seany, Manager of Mines and Mr. Bob Zerga, Mine Superintendent who so kindly arranged for the visit. Particular thanks are given to Mr. Al Metcalf who arranged the logistical details of our underground work, and whoa forward thinking program toward improving general mine communications at MAGMA provided the test-bed context for evaluation of the prototype equipment. The authors also wish to thank Benny Jiminez, Paul Ayala, Lorenzo De La Cruz, and Kent Billhartz for their day-by-day support of the testing. Mr. Jiminez's effort was instrumental in performing the testing, and he also provided a detailed understanding of the existing multi-level, multi-frequency trolleyphone system at MAGMA. the authors are also indebted to Dr. Larry Stolarczyk of A.R.F. for numerous detailed consultations on MF technology, and who was present and participated in the majority of the testing. 1.1 GENERAL BACKGROUND ON COPPER MINING AT MAGMA the San Manuel mine is the largest copper mine in the U.S., efficiently extracting copper and molybdenum sulfide from the low-grade San Manuel ore body.This ellipsoidal ore-shell body, with major workings located 2000-3000 feet mbelow the surface, has ore shell thickness varying between 100 and 1000 feet over a longitudinal extent of some 8000 feet, and with major & minor crossectional axes of about 5000 and 2500 feet respectively. The San Manuel mine operation employs in excess of 1000 men/shift with gross ore production between 60,000 and 70,000 tons/day. The mining operation employs the block caving technique; actively being implemented on three levels (2075,2375,2675 feet for haulage, and 2015,2315,2615 product/grizzly levels). The ore haulage is largely via D.C. trolley trains to four (3A,3B,3C,3D) product shafts for skipping to the surface. An additional three shafts at two locations are used to convey men, materials, and developmental rock extractions into and out of the mine. On each haulage level, the trains carry ore an average of two miles one-way from the ore raises to the dump points. 1.2 TOPOLOGICAL BRIEF OF THE HAULAGE LEVELS The 2375 level, on which all the testing was conducted, is illustrated in

Jan 1, 1981

By James A. Powell

A seismic system which utilizes in-mine geophones to detect trapped miners has been developed. Tests at the Bureau of Mines Safety Research Mine at Bruceton, Pa., and at two operating mines indicate such a system has a maximum detection range in excess of 1,000 feet. The system is fieldworthy and portable and requires less than 30 minutes to set up and check. Real-time detection is possible.

Jan 1, 1976

By W. V. Kuster

The investigation described in this report is part of a survey by the Federal Bureau of Mines of the titanium resources of the United States. The purpose of this project was to study a typical coastal deposit of beach sands in the Middle Atlantic States, particularly in the area surrounding the Delmarva Peninsula. The field investigation consisted of a reconnaissance survey of the Eastern Shore of Maryland, followed by surface sampling and drilling of the sands on Assateague Island, off the Atlantic coast of Maryland. The study covered the period January-December 1957. Detailed analyses were made of 5-foot sections of eight drill holes along the Maryland section of the island. The heavy-mineral fraction of these samples was separated by sink-float tests, using acetylene bromide as the heavy liquid. The minerals comprising the heavy fraction were examined microscopically, and mineral counts were made on selected samples.

Jan 1, 1959