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By William H. Pomroy

The potential for large-scale fires in underground fuel storage and fuel transfer areas prompted the Bureau of Mines to study fire hazards in these areas and devise improved fire safety technology for use in these areas. This report describes the Bureau's research to characterize and quantify fire hazards in fuel storage and fuel transfer areas, develop guidelines for safe and efficient fueling system design, prepare specifications for automatic fire sensing and suppression systems for these areas, test prototype systems in the laboratory and in underground mines, and perform cost-effectiveness evaluations of the devised technology.

Jan 1, 1985

By John C. Machisak

THIS PUBLICATION presents injury experience and employment data for the nonmetal mining industry in the United States for 1960. Data for the mines and mills are presented separately, as shown in the list of tables. The Bureau of Mines is authorized to send questionnaires to the nonmetal mining industry for their voluntary submission of injury and employment data. As the information submitted contributes largely to the promotion of safety in the industry, special effort has been made to make these questionnaires as complete as possible.

Jan 1, 1963

By J. G. Donaldson

Optimum conditions are given for chloridizing pyrite, marcasite, chalcopyrite, arsenopyrite, stibnite, and covellite with chlorine gas. The formation of metal chlorides is affected by the time and temperature of chloridizing, and by particle size of the sample. Chloridizing data also are given for chemically precipitated antimony (III) sulfide and copper (II) sulfide.

Jan 1, 1962

By H. Miller

In recognition of the need for a portable low-profile hard rock crusher, the Bureau of Mines sponsored the development of such a unit capable of moving up as the face advances. By allowing primary ore reduction at or near the mine face, the advantages of conveyor haulage From the face as demonstrated in soft ore and coal mines, could be realized for hard rock Brines as well. This report presents the results of a development and testing program for such a portable low-profile crusher, designed by Eagle Crusher Company and called the M44. The unit has an extremely low vertical profile considering its size and capacity. The crushing jaws, which form the crushing lumber, are inclined, with the lower jaw mounted for eccentric movement relative to the stationary upper jaw. This provides efficient crushing while also forcing material through the crushing chamber. As a result, the inlet feeder can be disposed horizontally. To facilitate correct processing of different materials, the size of the discharge opening and the action of the movable crushing jaw are variable. The unit also incorporates a feeder hopper and discharge conveyor that provides for efficient ore handling through the machine. Tests demonstrated the M44 Crusher's high capacity, efficiency and reliability. During testing, it crushed over 120,000 tons of materials of up to 32,000 psi compressive strength at production rates in the 200-350 tons/hour range. The unit operated smoothly under simulated low-headroom conditions, and Proved readily transportable.

Jan 1, 1983

By E. D. Calvert

The room temperature tensile properties of commercially pure titanium and seven titanium alloys prepared by skull casting were determined in the as-cast and heat-treated conditions. Data were compared with property data from tensile tests on wrought alloys of the same compositions in the as-worked and heat-treated conditions. The alloys evaluated were Ti-5Al-2.5Sn, Ti-6Al-4V, Ti-6Al-4V-lSi, Ti-6A1-2Sn-4Zr-2Mo, Ti-6A1-6V-4Zr-4Mo, Ti-6A1-6V-2Sn, and Ti-11.5Mo-6Zr-4.5Sn. Test specimens were prepared from both static and centrifugal castings and from plate fabricated by forging and rolling vacuum-melted ingots. The data indicate that the alpha and alpha-beta alloys have strength properties in the as-cast condition that are reasonably close to the strength properties of their wrought counterparts in the as-worked condition. An exception to this was the Ti-6A1-6V-4Zr-4Mo alloy for which insufficient data were available to form any firm conclusions. Fine-grained cast structures of Beta III (Ti-11.5Mo-6Zr-4.5Sn) were shown to be generally stronger in the as-cast condition than the wrought alloy in the as-worked condition except for the yield strength of one structure that was unusually low. Heat treatment of this alloy produced properties that were unpredictable. Indications are, however, that optimization of heat treatment parameters for each cast structure could produce strength properties equal to that of wrought materials.

Jan 1, 1974

By E. G. Davis

This report summarizes an investigation to determine the occurrence and recoverability of heavy minerals in sand and gravel operations in the Southeastern United States. A total of 61 samples from 34 sand and gravel operations in Alabama, Georgia, South Carolina, and North Carolina were studied. Heavy minerals, in amounts ranging from 3 to 400 pounds per ton, were found in all of the samples. Eighty percent of the samples contained 20 to 40 pounds per ton of heavy min¬eral. Valuable heavy minerals, in order of decreasing amounts, were ilmenite, kyanite, zircon, rutile, and monazite. Large-scale concentration tests were made on five samples from four major producers. Gravity concentration by Humphreys spirals and tables resulted in bulk heavy mineral concentrates that graded from 82 to 98 percent total heavy minerals. Mineral recoveries ranged, in percent, as follows: Ilmenite 60 to 95; zircon 96 to 100; kyanite 37 to 77; rutile 89 to 100; and monazite 96 to 100. Separation studies on the gravity concentrates showed they would respond to high-tension and high-intensity magnetic separation techniques to produce individual mineral concentrates. The recovery of each individual mineral from fractionation of the heavy mineral gravity concentrates, in percent, was: Ilmenite 77 to 94; rutile 41 to 68; kyanite 62 to 79; zircon 49 to 83; and rare-earth products, monazite and xenotime, 42 to 96. The grade of each mineral concentrate was: Ilmenite 51.1 to 68.7 percent Ti02; rutile 93.7 to 94.3 percent Ti02; kyanite 55 percent A1203; zircon 58 to 61 percent Zr02; and rare-earth products containing 73.8 to 90.8 percent monazite and 4.8 to 12.7 percent xenotime.

Jan 1, 1971

By P. G. Guest

Many of the gases and vapors used in anesthesia form explosive mixtures with oxygen or air. Sources of ignition for these mixtures always have existed, in operating and anesthetizing areas. When ether was first introduced, candles and gas-lights and coal stoves caused fires and mild explosions of the vapor-air mixtures. Today the chief causes of ignition are electrical. Fires and explosions are still much too frequent; and, because many of the mixtures are rich in oxygen, the explosions sometimes are very violent. Obvious sources of ignition, such as electric-power sparks and arcs in open motors, switches, and receptacles, as well as exposed incandescent wires and hot plates, usually are recognized by hospital personnel and corrected. A source of ignition that is not so obvious to many is static electricity. Until quite recently the presence of this hazard in a large majority of anesthetizing areas was not even suspected. However, numerous explosions -many of them fatal to patients -have been attributed, with good reason, to sparks from static electricity. In this paper the authors present facts to show the prevalence of static electricity in the average hospital anesthetizing area, and the degree to which this easily developed electrical form of energy constitutes an explosion hazard. An effort is made throughout to explain how static electricity behaves, why it is a hazard, and why certain safety measures are needed and recommended. Following the introduction and a short general discussion of static electricity, which incorporates a glossary of pertinent electrical terms, an account is given of inspections and tests made in a number of hospitals. This account includes many suggestions that should be useful to any person who plans to make similar inspections and tests. However, a comparison of materials and equipment used with those described in the report will reveal, in many cases, the potential dangers and obviate the necessity for tests. The hospital work was supplemented by considerable laboratory research designed to determine the spark energies required to ignite anesthetic-gas mixtures and the exact amount of time involved in developing the usual static charges. The results of this research are reported and the implications examined. The paper concludes with a general discussion of static and related hazards in anesthetizing areas and of the most urgent preventative measures.

Jan 1, 1952

Longwall minim; is a highly mechanized, high productivity mining method that is finding increasing applications in coal. Although the method would be highly advantageous in certain hard rock mining applications it cannot be applied in hard rock because no suitable cutterhead exists for such materials. The wedge longwall cutterhead is a, roller cutter device that is wedge-shaped, with cutters arranged to eliminate the very large normal force ordinarily required to cause such cutters to penetrate hard rock. However, the wedge geometry cannot be extended to a point: special nose cutters are needed to open a narrow slot for the wedge. This report describes the design, fabrication, and successful field testing of these critical nose cutters.

Jan 1, 1974

By M. J. Beus, B. Stewart, S. Iverson

Synopsis U.S. Mine Safety and Health Administration (MSHA) accidents statistics have identified ore-pass hazards as a significant safety problem in U.S. underground metal mines. The statistics show that nearly 75% of injuries are directly or indirectly related to pulling or freeing of ore-pass chutes, the use of hand tools in ore passes, fails of broken rock in ore passes and structural failures of chutes or gates and ore-pass walls. Researchers at the Spokane Research Laboratory of the National Institute for Occupational Safety and Health (NIOSH) are investigating hazards in and around ore passes in hard rock mines. Risk assessment methods, such as fault-tree analysis, have been employed to identify the most probable causes of orepass failures, define research priorities and analyse the factors that result in malfunctioning and unsafe ore passes. Static and dynamic loads are being measured in a laboratory setting with the use of a reduced-scale orepass mock-up. Field tests are being initiated in mine ore passes to determine static and dynamic loads. Data from the test ore passes are being compared with data from the mine to characterize true system behaviour. Computer modelling with the use of closed-form solutions, finite-element analyses and a newly developed particle flow code predicts static and dynamic leads from the flow and the impact of ore and waste in the ore pass. Preliminary results indicate that dynamic impacts from ore and waste rock might be considerably greater than expected. The total static load is substantially less than is typically used fur the structural design of chute and gate support members. It is suggested that damping factors, normal and shear stiffness and mass frictional characteristics have a significant effect on particle flow and resulting impact loads. This is being verified with the use of results from laboratory tests, field tests and particle flow analyses. Ore passes are vertical or steeply inclined openings in a rock mass through which ore and waste are transported and where they mar also be stored. A typical underground mining system is shown in Fig. 1. The arrows indicate the locations of ore passes and loading pockets and their relative configurations. The control of material flow is potentially one of the most hazardous operations in underground mining. According to MSHA data. most of the ore pass-related accidents in the U.S.A. are associated with the loading-unloading cycle and removal of blockages from ore passes. In particular, attempts to free ore hang-ups have resulted in multiple fatalities. The hazards are obvious; there is uncertainty as to when the ore will break loose and miners are unable to escape falling rock to a safe location. Although ore passes and chutes and gate systems for underground metal and non-metal mines roust meet the requirements specified in the U.S. Code of Federal Regulations (CFR). Parts 57 and 75, recent structural failures of ore-pass linings and gates have underlined the lack of adequate ore-pass design standards available to both MSHA enforcement staff and mine engineers. Because of the importance of an ore pass to the safe operation of a mine design criteria need to be assessed and new guidelines developed. Researchers at the Spokane Research Laboratory of the Nationals Institute for Occupational Safety and Health (NIOSH) are investigating hazards in and around ore passes in hard rock mines and studying design criteria for ore passes. The near-term goal is to conduct laboratory and field experiments and computer modelling to evaluate the ore-pass design criteria currently used. Literature searches, field site

By David A. Hill

This volume contains a summary of the results of past analyses on U.S. Bureau of Mines Contract No. 80122061 which are relevant to electromagnetic location in mine rescue situations. The analytical results indicated that detection of a trapped miner equipped with a special transmitter is feasible in most situations. Location of the miner by surface measurements is also generally feasible within certain errors. The feasibility of miner detection and the accuracy of miner location have been examined theoretically for a wide variety of situations. Agreement with experimental results is generally good in cases where measurements are available. Personnel involved in the investigation include J.R. Wait (Principal Investigator), D.A. Hill (Co-Principal Investigator), K.P. Spies, A.Q. Howard, and Lana Slusher (Administrative Assistant).

Jan 1, 1974

By Paul S. Goodman

Volume II The purpose of this study is to explain reasons for the productivity differences in mining crews. The literature on group productivity is examined. Then a new approach to estimating productivity differences is proposed and estimated across seven different mines. The results indicate a fairly robust model in terms of explanatory power. Recommendations on how the model can be used as a managerial tool are presented.

Jan 1, 1985

By G. B. Clark

One portion of this project was to test a blast shield and to design some features for a novel tunneling method termed the Automated Drill and Blast System (ADBS). The ADBS tunneling concept was developed for the design of an excavation system of greater efficiency than conventional drill, blast, and muck (DBM) systems by minimizing the down time and approaching the non-cyclic efficiency of continuous tunneling by means of small charge blasting. Small charge blasting uses simultaneously detonated, small light charges (less than 200 grams per hole) in four to eight shallow (24 to 36 in. deep) holes usually in a line. Limitation of the total explosive per blast reduced air blast overpressure, fly rock velocity, vibration, and noise. Thus, shielded equipment can remain at the face and a tunnel shield permits support personnel to work outby the shield at all times. The ADBS equipment will be composed of five components: (1) a chassis, upon which is mounted automated mucking equipment, and a frame which carries (2) the hydraulic drill(s), (3) load/blast system(s), and (4) control cab(s). The blast shield may have its own mounting. Ventilation, horizontal, and vertical alignment controls are incorporated in the system. The blast sheild, which protects personnel outby the advance equipment from air blast overpressure, fly rock, and gases, was tested with full scale small charge blasting. A blast shield was designed, fabricated, and tested under conditions anticipated for a working model. The shield was designed to expand to fit the tunnel shape, to seal itself to the tunnel periphery, and to contain air blast, fly rock, and gaseous explosion products. In addition, the shield, which was lined on the inby side with rubber belting, reduced the blast noise level outby the shield to 113 dBA or less, which is very near permissible levels. The test results show conclusively that the shield concept is viable. Based partially on the shield tests, a blast resistant control cab was designed to be located on the equipment chassis inby the blast shield to protect operators during blasting, and to contain all the controls necessary for system operation and safety. The economics of tunnel driving using the ADBS were compared with those of DBM. Four rock types, which determine advance rate, and four cost categories were examined. The results of this analysis are as follows: 1. ADBS advance rate is faster because equipment remains at the face and operation procedes semi-continuously. 2. ADBS bit and explosive cost (which accounts for 4%o of the total advance cost) is higher than for DBM. 3. ADBS equipment cost, on a per foot basis, is lower than for DBM. 4. Due to a reduced labor force, ADBS labor cost is lower than for DBM. 5. Because of the overbreak reduction due to controlled blasting, ADBS tunnel support cost is lower than for DBM. 6. On a per foot basis, total ADBS advance cost is from 17% to 20% lower than for DBM, depending on tunnel diameter and rock type. The small charge method greatly reduces ground vibration, increases personnel safety by limiting excessive fracture, and decreases tunnel support requirements by reducing overbreak. The overall analysis, tests, and results to date show that the Automated Drill and Blast System may be developed into a viable tunnel excavation method. The method also may be adapted to use in the mining of ore. For some types of mining operations, the cyclic nature of conventional drill and blast methods may not be a disadvantage, either where the drilling and mucking may proceed simultaneously or where the operations are carried out in adjacent working faces. The proposed method would find application where continuous methods of excavation are desirable.

Jan 1, 1979

By C. P. Haller, S. J. Aresco

"INTRODUCTION The Bureau of Mines has been active in promoting the purchase of coal for Government use under specifications that define the requirements in terms of heat¬ing value of the coal, expressed in British thermal units, and the composition as shown by proximate analyses. To these, when required, are added the fusing temperature of ash and the free-swelling index. Under most of these specification contracts the bidders guarantee the quality of the coal, and that guaranteed by the successful bidder becomes the standard of his contract. The deliveries are sampled in accordance with instructions./ issued by the Bureau; the samples are analyzed in the Bureau's laboratory to determine whether the coal is of the quality guaranteed by the contractor and, if it is not, a price adjustment is made.Analyses of the delivered coal and ""tipple"" samples (samples collected at mine tipples as coal is being loaded into railroad cars or trucks) provide valuable data for use in evaluating future bids. Moreover, the continuous sampling of coal as delivered is a check on the practical results obtained in burning the coal.The Government purchased approximately 5-1/2 million tons of coal in fiscal year 1951. In connection with these purchases, the Bureau analyzed 10,700 samples. These are published for the use of Government officials and the public. A series of publications s been completed on coal analyses by States, or, where production is small in any State, by groups of adjacent States. In addition to analytical data showing composition and quality of coals of the States, the publications contain brief descriptions of the geological structure of the coal basins, typical mining conditions in the various districts, the principal economic data of the industry, and the relationship of mine or channel samples to delivered and tipple samples. The present publication covers many States but deals only with the analytical data of the coals.The technical papers published by States, with their numbers are: Alabama, 347; Alaska, 682; Arizona, California, Idaho, Nevada, and Oregon, 696; Arkansas, 416; Colorado, 574; Illinois, 641; Indiana, 417; Iowa, 269 and 706; Kansas, 455; Kentucky, 308 and 652; Maryland, 465; Michigan, North Dakota, South Dakota, and Texas, 700; Missouri, 366; Montana, 529; New Mexico, 569; Ohio, 344 (and Bulletin 499); Oklahoma, 411; Pennsylvania Bituminous, 590 (Supplement, 645); Pennsylvania Anthracitic, 659; Tennessee, 356 and 671; Utah, 345; Virginia, 365 and 656; Washington, 491 (Supple¬ment, 618); West Virginia, 405 and 626; and Wyoming, 484."

Feb 1, 1953

By Wilbur I. Duvall

This report presents experimental data on crater formation in four rock types. From an analysis of these data a general theory of rock breakage by explosives is deduced. This theory accounts for most of the rock breakage and other phenomena that occur when a concentrated charge is detonated in a drill hole near a free surface. From instrumented crater tests and high-speed motion pictures of the explosions, various experimental checks on the deduced theory of rock breakage are made. The general mechanism of rock breakage by explosives as described in this report is outlined briefly below. The detonation of the explosive charge creates a high gas pressure in the charge hole, which in turn generates a compressive strain pulse in the surrounding rock. This compressive strain pulse travels outward in all directions from the charge hole. Near the charge hole the amplitude of this strain pulse is sufficient to cause crushing of the rock. However, as the strain pulse travels outward its amplitude decays rapidly until no further crushing of the rock is possible. The compressive strain pulse continues to travel outward until it is reflected by a free surface. Upon reflection the compressive strain pulse becomes a tensile strain pulse. As the strength of rock in tension is much less than in compression, the reflected tensile strain pulse is able to break the rock in tension, progressing from the free surface back toward the shot point. In other words, the rock is pulled apart, not pushed apart.

Jan 1, 1957

By Douglas W. Fuerstenau

A detailed experimental study of breakage kinetics in a batch rod mill indicates that while the system is inherently nonlinear, the feed disappearance kinetics for the finer sizes and the coarser sizes at high holdups is well represented by first-order behavior. The set of breakage distribution functions was found to be normalizable and independent of particle holdup in the mill. The size discretized initial breakage rate functions were found to be inversely proportional to particle holdup and almost directly proportional to particle size. A linearized grinding model that results from the first-order Taylor's series expansion of the breakage rate expression was found to give good simulations of the mill product size distributions without recourse to any parameter improvement routines. Material transport through rod mills was evaluated by investigating the effect of operating variables on the holdup and residence time distribution of particulate solids flowing through the mill. Identical experiments performed in a continuous ball mill helped delineate the fundamental differences in particulate transport through these two types of mills. Mechanistic interpretations of the observed transport phenomena are presented, and their implications in the context of tumbling mill analysis and design pointed out.

Jan 1, 1980

By D. R. Corson

The Bureau of Mines instrumented an operating stope in a deep vein mine to determine the amount of wall closure and the magnitude of loads imposed on a backfilled stope. Hydraulic pressure cells used to monitor pressures in the fill) with bolt closure points to measure vein-wall movement and the procedure for their installation) are described) and the results of the laboratory calibration analysis of the reliability of these cells are given. An attempt to monitor density and moisture changes in the backfill with a nuclear density-moisture device was only partly successful) owing to either deformation of access tubing or relative displacements of the vertical vein walls relative to the emplaced tubes. Data from the instrumentation indicate vein-wall closure in excess of 20 inches and pressures within the fill approaching 600 psi. Analysis) incorporating field data with results of laboratory tests of the compressibility of timber and tests defining the constrained moduli of the fill) points out the limited restraint to vein closure offered by timber sets. Ground movement induced by mining activity could be more effectively con-trolled by the use of densely compacted backfill.

Jan 1, 1967

By G. Huang, S. Vytla

Computational fluid dynamics (CFD) is a continually developing numerical technique by which complex fluid flow problems can be solved on computers. To be able to transfer CFD expertise to analyze and design a face-ventilation system, a comprehensive validation study of current CFD tools against mining-related benchmark experiments is required. In the last four years, such studies, funded by National In¬stitute for Occupational Safety and Health (NIOSH) under Grant #R01 CCR415822, were conducted in the Department of Mining Engineering at the University of Kentucky. A wide array of ventilation schemes and mining configurations were considered. Recently, an additional study was performed in cooperation with the NIOSH Pittsburgh Research Laboratory. This study was dedicated to the methane behavior in the empty (containing no equipment) face area with a blowing curtain and a 10. 7-m (35-ft) setback. In this paper, the authors present and discuss the computer-simulation data and compare them with data collected during the laboratory studies at the NIOSH lab.

By H. B. Hill

Virtually from the date of its discovery in December 1928 the Oklahoma City oil and gas field, Oklahoma County, Okla., has attracted national and international attention. The history of the field has been associated closely with rules and regulations pertaining to production control and proration practices. Early development was rapid, and Oklahoma City soon attained the distinction of being the first major field in which the producing wells were over a mile in depth. Before the structural features of the reservoir had been disclosed every outpost well was considered a "wildcat", so complex appeared the underground geology, The wells not only were deep and the reservoir pressures high, but the production from the early wells was a high A.P.I. gravity oil, accompanied by large volumes of gas. Several "blow-outs" and fired occurred during the early development, and the close proximity of these now unusual events to a large city gave the field much publicity, both local and national. Later, when its production threatened to demoralize the market for oil and its products, the Oklahoma City field again attracted wide publicity. The then little understood geology of the underground reservoir and the diversity of ownership of the land overlying the producing structure made orderly development and production difficult to enforce. Proration was enforced with difficulty., and the, rate of production from the field continued to increase until, on August 4, 1931, martial law was declared in all prorated fields of the State, and most of the wells in the Oklahoma City field wore shut n. On October 9, 1931, permission was granted t o reopen and prod e the wells in the Oklahoma City field. The order permitting production of the wells provided that during the period beginning October 10 and extending to October 20, 1931, all wells in the field that hid not produced more oil than specified under the earlier proration orders and that had been shut in for the required 55-day period were permitted to produce 5 percent of their last potential during that 10-day period. On November 7, 1931, the Corporation Commission of Oklahoma issued an order which divided the Oklahoma City field into the Wilcox, Arbuckle, and Lower Simpson zones for proration purposes. Wells producing from the Wilcox zone were allowed to produce 4 1/2 percent of their potential5/ from November 1, 1931, to January 1, 1932; Arbuckle limestone wells were allowed to produce 250 barrels of oil daily; and Simpson sandstone wells which had a potential of over 2,000 barrels per day or which made 3 percent or more water were allowed to produce 200 barrels of oil per day.

Jan 1, 1937

By R. A. Hatch

The United States depends mainly on foreign imports, principally India, for an adequate supply of sheet mica suitable for its many uses in the electrical and electronics industries. Sheet mica has, in fact, played an indispensable role in developing these industries,4/ and, in turn, the products of these industries (radio, radar, electric motors, and many others) have become indispensable to the military forces. Because or these conditions, sheet mica in its better qualities long has been classified as a strategic and critical material, and from this stems the wide-spread interest in substitute materials of which synthetic mica is one potential type. The United States Departments of the Navy and the Army entered into a cooperative agreement with the Bureau of Mines in 1947 to study ways and means of synthe-sizing and growing mica crystals. This is one of a series of reports5/ that describe the results of this program.

Jan 1, 1956

By James M. Link

Underground coal mining is the most hazardous industrial occupation in the United States.. Haulage accidents in underground coal mines contribute substantially to the occupational hazards of this type of work. In an effort to improve mine safety, a completely hydraulic transport system has been designed to move coal from the mine face to the preparation plant. The capital investment for the hydraulic haulage system is great, and the size of the mine developed for this study is large. Mine production for the conventional haulage mine is 1. 1 million tons of clean coal per year, while the hydraulic haulage mine will produce over 1.6 million tons of clean coal per year. Safety statistics for a number of mines of this size are shown.

Jan 1, 1975