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A revolutionary development in blasting materials has occurred in the past 5 years. Ammonium nitrate, sensitized by the addition of a few percent of No. (class) 2 diesel fuel or other suitable petroleum product, has been adopted as a low-cost, efficient blasting agent2/ by a wide variety of mining and heavy construction operations. Moreover, ready availability of the component supplies throughout the Nation and ease of their formulation into effective blasting agents have made it possible for the blaster himself to pre-pare the mixes at the site of use. Although these field-mixed products have partly or wholly displaced explosives such as dynamites in some applications, there are other blasting operations where the fuel-sensitized ammonium nitrate agents do not appear to be well adapted. Important safety characteristics are inherent in ammonium nitrate-based blasting agents prepared under competent factory control, or when field mixed under equivalent close control--as long as the hazards they present are clearly understood and respected. This publication attempts to summarize knowledge of those hazards and to incorporate practical ways to cope with them. A series of safety recommendations thus is presented. These recommendations cover the preparation, storage, transportation, and use of field-mixed ammonium nitrate blasting agents.

Jan 1, 1960

By L. C. IlsLey

"The underlying principle of the storage battery was discovered within the memory of many still living, hence storage-battery locomotives, which were made possible by this discovery, represent a still more recent development. The Journal of the Franklin Institute for December, 1883, gives an account of the building of an electrical omnibus by The Electrical Storage Battery Co., of Millwall, London, This, if not the first, was among the first attempts to, use the storage battery as the source of energy for locomotion.During the period 1888 to 1893 occasional attempts were made to apply this idea practically, but not much was achieved along useful lines.. The progress during the next, 15 years was slow, and little success was achieved. Therefore, the last two decades may be considered the successful development period of the storage-battery locomotive.In coal mining, the application of the storage-battery locomotive is chiefly in gathering service in competition with animal or cable-reel locomotives. The locomotive delivers empty cars to the rooms or working places where the coal is being mined and hauls the loaded cars from the rooms. The length of haul depends upon the distance of, the junction point from the active workings of the mine. As a rule the haulage from the junction point to the foot of the shaft or bottom of the slope is done by trolley locomotives, but in some instances where the grades permit this main-line haulage work it: performed by storage-battery locomotives.As the place storage-battery locomotives occupy with respect to safe haulage in gassy mines appears to be misunderstood, it is the purpose of this paper to clarify this point.Those who originally designed storage-battery locomotives had no thought of safety in mind and it is just as well that they did not, as these early types of storage-battery locomotives with open controller open motor, open rheostats, flimsy headlights, and crude wiring were very dangerous electrically and not at' all suitable for use in a gassy mine. These locomotives, however, were placed in mines in which open lights were permitted and they replaced considerable animal haulage. Later they were gradually introduced in mines where closed lights were used, but in most cases these mines were not extremely gassy."

May 1, 1928

By J. A. Holmes

As part of its investigation of methods of increasing the efficiency of fuel resources, the Bureau of Mines is continuing the study of the general problems involved in the economic use of fuels in gas pro- ducers that were carried on by the technologic branch of the United States Geological Survey. In the course of this study many inci- dental problems of prime importance have demanded consideration. Among them is the determination, under practically constant con- ditions, of the duration of gas-producer tests necessary to reduce the possible error to a minimum. A discussion of this proper length of test period forms the first part of this bulletin. An attempt to dispel some of the uncertainties that arise in an investigation or to account for some of the minor phenomena often necessitates a systematic and careful study of certain chemical, ther- modynamic, or physical laws, or a combination of these, which in turn may result in information or conclusions of great value. Fur- thermore, seemingly unimportant observations often prove to bear directly upon the design, construction, or manipulation of the machine under test. In attempts to improve the gas producer considerable attention has been given by manufacturers to such questions as fixing or isolating the combustible components of the tar, utilizing the waste heat of the producer for the generation of the steam required by the plant, apply- ing the suction principle to bituminous coal and other tarry fuels, and many allied subjects. Little study, apparently, has been given to the physical-chemical conditions most suitable to the formation of hydro- gen and of carbon monoxide, with a view to regulating the percentages of these gases in generators using fuels other than anthracite coal or charcoal. An attempt was made to introduce such investigations when the operations of the government coal-testing plant began at the Louisi- ana Purchase Exposition in St. Louis in 1904, but owing to adverse conditions these problems were abandoned until the plant was trans- ferred to Norfolk in 1907. The delay in beginning this line of inves- tigations, together with the short testing season at Norfolk, limit the second part of this report to preliminary physical-chemical studies, which, however, were sufficiently significant to warrant careful ex- amination pending more complete researches, which are being carried on at the station located in Pittsburg.

Jan 1, 1911

Analyze variably saturated flow in mining and other environmentally sensitive settings through the use of personal computer (PC) based software. The Problem There is an increasing need for mining companies to quantify water contaminants and their migration in the unsaturated zone. New and innovative computer models are therefore necessary to more accurately assess the potential impact of mining activities on groundwater resources and the environment. Applications for unsaturated flow modeling include (1) assessment of the impact of surface or underground extraction mining on flow conditions; (2) assessment of various hydrologic designs for resource recovery, such as in heap leaching or in situ leach mining; (3) assessment of the transport of conservative nonrcactive constituents in response to migration from spoil piles, or acid mine drainage; and (4) assessment of any combination of the above to facilitate expeditious permitting for mining operations. Since the mid 1970's there has been a proliferation of mathematical models and procedures for solving unsaturated now problems; however, few of these approaches have been integrated into software for public use. Commercially available unsaturated flow software models typically are limited to a single application and have restricted approaches. For example, computer codes that solve the unsaturated flow problem based on an analytic expression arc restricted to a simple geometric shape, homogeneous and isotropic materials, and provide only information regarding pressure and flow conditions in specified wells. The variably saturated computer codes that take advantage of the finite difference or finite element approach arc often restricted to either steady state or transient solutions; utilize one specific moisture-retention relationship and moisture-dependent permeability function; do not account for hysteresis; and are used with specific hardware such as a mainframe or work station. This restricts their use because they arc hard to interface, and arc often poorly documented, support dependent, and expensive. The mining community needs a flexible, more broadly based, variably saturated fluid flow simulator for the PC-platform to more accurately predict fluid movement in the mining environment.

Jan 1, 1991

By Doris M. Hyde

The petroleum sector continued to dominate Ecuador's economic activity in 1978, accounting for nearly 10% of the gross domestic product (GDP) and 42% of total export earnings. In 1978, a more flexible petroleum export price was adopted which permitted an expansion of foreign market opportunities. During 1979, Ecuador took advantage of disruptions in the world petroleum supply and renegotiated existing export contracts to accomodate the rapidly escalating world market price. Petroleum sales on the spot market were maximized, and prices reached unprecedented levels. By taking royalty payments in-kind, Ecuador increased its volume of petroleum available for term and spot sales. Although the volume remained about the same as in 1978, the total value of 1979 petroleum exports increased to about $920 million, reflecting a 76% increase over the $523 million received for 1978 crude oil exports. Concerned about its steadily decreasing petroleum reserves (recently estimated at about 1.1 billion barrels), rapidly increasing and heavily subsidized domestic consumption, and the prospect of having to import substantial amounts of petroleum by 1983, Ecuador announced plane which it anticipates will double reserves within the next 5 years. The Government plans to spend $1 billion over the 5-year period. Plans include drilling 325 exploration and development wells and establishment of a secondary recovery program for the Sachs and shushufindi fields. Some exploration wells are expected to reach a depth of 17,000 feet. It is thought that substantial reserves exist in the southeast Amazon region. Late in 1979, it was announced that some 10 million hectares will be available for exploration tenders-probably in blocks of about 200,000 hectares each-in the Amazon jungle of the Orients and in the coastal and offshore regions. The Government planned to commission a private, independent analysis of Ecuador's petroleum reserves that would also include recommendations on efficient production levels.

Jan 1, 1981

By Peter F. R. Altounyan, Paul B. Cartwright, John W. Cassie

This paper describes measured data on strata behavior obtained in recent years from sites in the United Kingdom and the implications for pillar design. The data include results from overcoring stress measurements adjacent to coal mine roadways and deformation monitoring related to longwall extraction. The stresses adjacent to mine roadways or entries have been measured at a number of coal mine sites in the United Kingdom. The results are analyzed with regard to the information they provide on pillar behavior and strength estimates. A reduction in stress consistent with yielding of the strata adjacent to the roadways is evident. This is consistent with the confined core model for pillar behavior. The pillar strength is dependent on the rate atwhich vertical stress can increase with distance from the pillar edge and hence the confinement provided to the yielded material. The measured data indicate a wide range in pillar strengths. Two groups of results are identified that show significantly different behavior corresponding to differing effective pillar strengths. Estimates of pillar strengths derived from the measured data for these two groups are compared with established equations used for pillar design. The differing behaviors and strengths are attributed to variations in the amount of yielding and deformation in roof and floor strata and hence in the amount of confinement they provide to the coal seam. Numerical modeling is used to provide a comparison with the measured data and to indicate that this provides a feasible mechanism to account for the measured data. As the depth of mining increases, pillars tend to become increasingly wide and squat. In such cases, it is possible for the surrounding roadways to become badly deformed and damaged while the pillars remain stable. The criteria of comparing pillar strengths and loads to establish pillar stability become less applicable in these circumstances; rather, considerations of roadway stability may be the limiting factor in determining suitable pillar dimensions. This is the case for pillar dimensions typically employed around longwall panels in the United Kingdom. Depending on the properties of the site and what are deemed to be satisfactory roadway conditions, this can lead to wide variations in required pillar dimensions. Measured data for deformations in roadways influenced by adjoining longwall workings are presented. These show that in some circumstances the influence of longwall extraction can be transmitted over large distances and confirm the variability in required pillar sizes depending on site properties.

Jan 5, 1999

By Martin Hertzberg

The concept of limit burning velocities is being used to formulate a quantitative theory of flammability limits. Competing and complicating processes dissipate power from a combustion wave and quench propagation at some characteristically low limit velocity. Two earlier reports dealt with process (a), natural convection (RI 8127); and with process (b), conductive-convective wall losses (RI 8469). This report considers process (c), radiative losses, and process (d), selective diffusional demixing. The limit burning velocity for radiative losses is [(SU)c ~ e cp] It is normally not a significant process for premixed gas flames; however, for laminar flame propagation in a gravity-free environment, it is the only loss process present. Accordingly, limit velocities at zero gravity should be much lower than on earth, and premixed gases should be flammable over wider composition ranges. For highly emissive dusts, soot, or ash-forming flames, radiative losses are significant even under normal gravity conditions. Selective diffusional demixing is a complication that manifests itself in the form of cellular flame structures and wide gaps in composition between upward and downward propagation limits. It induces a burning velocity shift (?Su)d, causing the flame to behave as though it were richer or leaner than its initial composition, depending on whether the fuel or the oxidizer mole¬cule has the higher diffusivity. A simple pertubation analysis shows that the process generates stable cellular flames whenever [dsu] > 0, where ci is the concentration of the reactant with the higher diffusivity. The same analysis indicates that at the opposite stoichiometry, where [dcu] < 0, laminar structures are stabilized. A simple formula for the diameter of cellular flames is derived, which predicts that dcell should vary inversely with the product of the diffusivity difference Di -DJ, and the burning velocity slope [dc ]. Those predictions are in good agreement with the available data. Buoyancy forces for upward propagation induce flame front curvature, and hence buoyancy and selective diffusion are mutually supportive in upward propagation. Thus large gaps in composition between upward and downward limits are present in those composition domains where cellular flames are stable.

Jan 1, 1981

By D. H. Baker

This report, one of a series prepared by the Federal Bureau of Mines on the technology of titanium, describes the development of an electrorefining cell from which cathode deposits up to 30 inches long can be produced. Data show individual deposits to be homogeneous, comparable from top to bottom, and not varied in quantity or quality of metal over the entire length of the deposit. Investigations concerned with various diameter cathodes showed that cathode current efficiencies were materially reduced with cathodes less than 2 inches in diameter. Deposits produced on 1/2-inch-diameter cathodes showed cathode current efficiencies of 19 percent and on 1-inch cathodes, just over 22 percent, whereas deposits produced on 2- and 4-inch-diameter cathodes showed about 70-percent current efficiencies. A series of 20 deposits on 4-inch-diameter cathodes was produced under simulated production operation. The average current efficiency, based upon a divalent ion, was 72 percent with 11 1/2 pounds of metal produced per cycle of operation - 98.8 percent of which had a Brinell hardness number (B.h.n.) of 80 or less.

Jan 1, 1959

By John Nagy

Data from experiments and literature on gas and dust explosions in closed vessels are presented to show the effects of the vessel size and shape; the initial pressure, temperature, and turbulence; and the fuel-air ratio and the strength of ignition source on pressure and rate of pressure rise. An attempt is made to correlate some of these parameters with an elementary mathematical model. Analyses of data on 17 vessels--spherical, cubical, cylindrical, and rectilinear in shape--ranging from 0,018 to 905 cubic feet in volume were made. Although the study was initiated to provide information primarily on dust-explosion phenomena, some data were obtained for gas explosions so that the factor of initial turbulence caused by the pneumatic blast used to disperse dust could be evaluated. The mathematical interpretation of the explosion phenomena was in general accord with the experimental data, and correlation was obtained for several factors affecting explosion development. Neither vessel size nor shape affected maximum pressure (when heat loss was neglected). The rate of pres-sure rise decreased as vessel size increased, and the rate was proportional to the ratio of the surface area of the vessel to its volume. Induced initial turbulence markedly increased the rate of pressure rise and slightly increased the maximum pressure. Changes in initial pressure produced proportional changes in maximum pressure and rate of pressure rise. Initial temperature and maximum pressure were inversely related, but initial temperature had little effect on the maximum rate of pressure rise. Comparative experiments showed no appreciable difference in explosion development of gas and dust fuels. The vessels and experimental work are described, and all data are given in tables. The mathematical model will facilitate extrapolation from laboratory to industrial conditions and will allow comparison of results obtained in different laboratories.

Jan 1, 1971

By J. H. Perry, D. W. Houseknecth, C. A. Kertis, A. T. Iannacchione

Major problems facing the development of coal and gas resources of the Hartshorne Coalbeds include the complex distribution of minable and unminable coal, high methane content and bed pressure, faulting, variations in degree of dip, presence of natural gas fields in associated sandstones, and legal problems, caused by local geologic characteris- tics, in identifying gas origin. This Bureau of Mines study provides pertinent geologic information for long-range planning of subsurface coal and gas production from the Hartshorne Coalbeds. Pittsburg, Coal, Hughes, and Atoka Counties, OK, have approximately 1 billion metric tons (1.1 billion short tons) of Hartshorne coal in place. The methane resource of Hartshorne Coalbeds is estimated at 9.2 billion m3 (325 billion ft3) and varies with overburden and rank, These estimates were compiled from gas well density logs, coal-core data, out- crop measurements, abandoned mine maps, and the literature. The Hartshorne Coalbeds represent a valuable coal and gas resource which at present is contributing nothing to coal and gas production. This report discusses potential problems and serves as a reference for future exploration and development work. Recognition of these potential geologically related problems prior to development of this basin will reduce hazards and allow for an economic recovery of these resources.

Jan 1, 1983

By Bryan F. Ulrich

If resin-grouted roof bolts are not installed correctly, mine roof reinforcement can be affected, and roof falls can result. A bolt installation variable that is readily controlled by the mine operator is annulus thickness. The objective of this U. S. Bureau of Mines investigation was to study the effect of annulus thickness variations on the integrity of 3/4-in-diam resin-grouted bolts. Forged-head test bolts, 2 and 1 ft long, were installed in concrete blocks that had been drilled with 1-, 1-1/8-, 1-1/4-, 1-3/8-, and 1-1/2-in-diam bits. Standard pull tests were performed; then the concrete was broken away from the bolts so that grout mix quality could be inspected. It was found that the optimum annulus thickness is 1/8-in (l-in-diam drill hole), and as annulus thickness increases from the optimum, there is a corresponding decrease in grout mix quality, effective grout ratio, and axial stiffness.

Jan 1, 1989

By R. O. Dannenberg

Exploratory research was conducted to ascertain the feasibility of recovering beryllium from Spor Mountain, Utah, ore by the sodium fluosilicate (Na2SiF6) method used industrially for processing beryl. Na2SiF6 requirements to form water-soluble salts of beryllium were determined for three different samples--two clayey ores containing 0.6 and 1.3 pct BeO, respectively, and a calcareous ore containing 0.6 pct BeO. The ores were sintered with Na2SiF6 at 750° C for 3 hours. About 80 pct of the beryllium was rendered water soluble by sintering the 0.6-pct BeO ores with 30 and 40 lb Na2SiF6 per lb BeO contained in clayey and calcareous ores, respectively. The higher grade 1.3-pct BeO clayey ore required only 20 lb Na2SiF6 per lb BeO to solubilize over 80 pet of the beryllium. As the 0.6-pct BeO clayey ore was possibly the most representative of the deposit as presently delineated, leach liquor prepared from this sample was treated to recover beryllium salts. About 60 pet of the beryllium contained in the ore was recovered as beryllium hydroxide, which then was calcined to yield a 99-pct BeO final product by means of the following operations: (1) Sintering with Na2SiF2, (2) leaching with water, (3) precipitating crude Be(OH)2 by adding NaOH, (4) filtering to separate the hydroxide from the NaF solution, (5) redissolving the crude Be(OH)2 in excess NaOH, (6) filtering to reject caustic-insoluble impurities, and finally (7) boiling the Na2BeO2 filtrate from (8) to precipitate pure Be(OH)2. Reagent consumption was 49 lb Na2SiF6 per lb BeO recovered and 17 lb NaOH per lb BeO.

Jan 1, 1963

By Lisa Steiner, Robin Burgess-Limerick

In 2004, there were 646 underground coal mines in the USA employing 37,445 miners. These mines reported 3556 injuries to MSHA that year - 17% were associated with bolting machines (593 injuries), 8% with continuous miners (283 injuries), and 4% each with Scoop/LHD (1 51 injuries), shuttle cars (134 injuries), and personnel transport (145 injuries). The use of the frequency of reported injuries for the prioritization of risk control strategies has limitations because of the tendency to underestimate the importance of relatively rare but high-consequence events. Injury reports also underestimate the contribution of risk factors such as whole body vibration, which have a long-term cumulative contribution to an elevated risk of injury. However, taking these limitations into consideration, the results of the injury narrative analysis suggest the following hazards as the highest priority for elimination or control: • rock falling from supported roof • inadvertent or incorrect operation of bolting controls • handling continuous miner cable • collisions while driving LHD/Scoop, shuttle cars and personnel transport • rough road while driving or traveling in LHD/Scoop, SC & PT

By Jeffery M. Listak, Ernest A. Curth

"It took only 9 years from the first appearance of roof shields on the U.S. longwall mining scene to the present predominance of shield faces. An apparent consequence is the welcome downward trend in accidents related to failures in longwall ground control. The report addresses load prediction, the effects of shield geometry on support loads, factors contributing to ground control, and related techniques. The Bureau of Mines took an active part in the evolution of longwall mining, including developing one of the first shield faces in 1975, the first lemniscatetype shields in 1976, the Mine Roof Simulator completed in 1980, the first shields in steeply pitching coalbeds in 1981, and the first multilift working of a thick coalbed in 1982. The information offered in this state-of-the-art report will assist in establishing criteria for roof support selection. INTRODUCTIONIn terms of roof support technology, the introduction of shields in the seventies and eighties was a major step forward. Table 1 indicates the predominance of shield faces on the U.S. longwall mining scene and its apparent consequence, a welcome downward trend of ground-control related accidents.The rapid progress of longwall technology began in 1975 and is highlighted by a development leading from the caliper shield to the modern lemniscate type. Shields have been gaining preference over chocks since their advent to longwall mining in 1975. Safety and productivity factors favoring shields over chocks are--1. A sheltered working space requiring minimum cleanup work.2. Structural stability that allows advancing without delay even with brushing roof contact and, by controlling all lateral loads, removing the requirement for cumbersome leg restraint and restoration devices.3. Wide hydraulic range of mining height.The development of roof shields in the Federal Republic of Germany preceded and paralleled their rapid adoption by U.S. miners. In 1983, 87 pct of the total production in Germany came from shield faces (1).2 In Great Britain, the National Coal Board introduced the first shield support system in 1977 to initiate Advanced Mining Technology (ATM) and Heavy Duty Mechanization (HDM) programs; as of mid-1982, 44 shield faces were operating, or 7.5 pct of a total of 581 faces CJ_)."

Jan 1, 1985

By C. B. Manula, R. L. Sanford, R. A. Rivell

PART I. FLOW CHARTS

Jan 1, 1974

By Joseph C. Zelanko, Anthony T. Iannacchione

One of the most difficult, longstanding engineering problems associated with coal mining is the catastrophic failure of cog mine structures known as bumps. For more than 70 years, researchers and practitioners have assembled a wealth of technical information on coal bumps in an attempt to understand and control them. However, many technical issues raised long ago are still being debated today. This paper examines past experiences and recognizes achievements in the realm of coal bumps. U.S. Bureau of Mines (USBM) researchers collected and analyzed 172 coal bump incident reports and compiled the pertinent statistics into a database. Actual field studied are also discussed. Examination of past experience has shown that there is no one set of defining characteristics that is responsible for coal bumps. In all cases, bumps occur when complex arrangements of geology, stress, and mining conditions interact to interfere with the orderly dissipation of stress. However, it is evident from the database that a tremendous reservoir of knowledge has been established from past experience that has unquestionably limited the severity of coal mine bumps in the United States.

By K. J. Liles

To encourage the development and conservation of the Nation&apos;s ceramic raw materials, the Bureau of Mines, U.S. Department of the Interior, under an agreement with the Georgia Department of Natural Resources, evaluated samples from a clay resource located in Clay County, Ga., for use as a raw material for lightweight aggregate. The physical properties of clay samples, taken from 5-foot increments of drill cores from 11 holes, were determined. Raw materials were plasticized and extruded to produce 1/2-inch by 1-inch pieces and fired in a rotary kiln to produce lightweight aggregate samples. The resulting expanded material had excellent loose pour weights ranging from 30 to 37 lb/cu ft. The expanded test materials were crushed, sized, and used as aggregate to form concrete cylinders. The unit weight and compressive strength of the cylinders averaged 106 lb/cu ft and 3,350 lb/sq in, respectively; these values meet American Society for Testing and Materials specifications for structural concrete. The clay from the drill cores was also used to make test bars that were evaluated for building brick.

Jan 1, 1980

By John Durkin, Roy J. Greenfield

This report discusses the configuration and system deployment for the postdisaster surface seismic system for detecting and locating trapped miners. It analyzes the results of 15 field tests to define a signal model, background noise levels, and subarray performance. A waveform modeling procedure is described and compared with observed waveforms. The resulting similarity indicates that the major factors affecting the signal amplitude, waveform, and spectral character are understood. A model is presented which gives the signal amplitude as a function of source type, source depth, and horizontal off- set between source and receiver. Using this model a curve is presented which gives the range at which a signal will be detected for different signal and noise levels. Finally, and most important relative to the mission of the sys- tem, the ability of the system to detect signals on one or more subarrays is put into a probabilistic framework. For a strong source it is almost certain that a subarray directly over the source will detect the signal. After signal processing, it is highly likely that signals will he detected on sufficient subarrays to locate the trapped miner. Location errors have been found to be less than 100 ft in the majority of cases. Techniques have been used that can reduce the location errors to this level even when soil layer variation between subarrays is severe.

Jan 1, 1981

Roof bolting is a slow, expensive, and absolutely necessary mining activity. Roof-bolting activities also account for the largest percentage of lost-time injuries in underground coal mines. Dull bits result in the application of excessive force, which bends drill steel; and changing starter and finisher steel was the most significant source of injuries associated with roof bolting. The tungsten carbide (WC) roof-bolt drill bits currently used in American, Canadian, and Australian coal mines are capable of drilling only one or two 4-ft holes and then they are resharpened once or twice to drill another one or two holes. Often, in medium-hard rock formations (sandstone), as many as 20 to 30 bits may be required to drill a single 4-ft hole. Objective To compare the performance, longevity, and cost of newly developed polycrystalline diamond compact (PDC) roof-bolt drill bits against WC bits. Research and Background Diamonds for industrial use have been scarce throughout history and continuous attempts have been made to produce synthetic diamonds. The first synthetic diamonds were manufactured in 1953. Economically produced synthetic diamonds are 1 to 10 µm in size, but in this size range, they are totally unsuitable for industrial applications. In 1972, General Electric Co. developed a process to sinter the artificially developed, minuscule diamond particles into usable products. In this process, several layers of diamond-particles-and cobalt mixture are overlaid on WC substrates and sintered at 2,350° F and 1,000,000 psi. The final product is known as PDC. Today, PDC&apos;s are available in many shapes, but their size is limited to 3/4 in, or less, in length or diameter, because pressures greater than 1,000,000 psi are required to manufacture larger PDC&apos;s. At present, diamond and PDC arc the hardest and strongest commercial materials available for wear surfaces. The PDC used in this study had a hardness 3.61 times greater than that of WC.

Jan 1, 1992

Improve surface mine haul-age truck safety and efficiency by providing a means for sensing the weight and distribution of the truck&apos;s load. Approach The safety hazards and the economic implications of overloading mine-haulage trucks were analyzed to determine the net effect of that practice on mine operations. The analysis indicated that: Ten to 25 pct overloading is typical; Higher haulage, tire, and maintenance costs occur due to overloading (a 20 pct overload was estimated to increase operating costs 7 pct per ton); Ten to 15 pct of haulage? truck injuries and fatalities are attributable to overload conditions, at a cost over one million dollars per year to the mining industry. Vehicle-weighing technology that conformed to the performance specification for an off-highway haulage truck overload detection system was selected for in-mine testing. How It Works Normally, volumetric estimates are made during the loading sequence which do not account for the varying ore density. Hence, the actual weight of a load is never precisely known. By providing a precise measurement, an overload detection system will not only indicate vehicle overloading, but will also show under-utilization of the truck. One system tested by the Bureau of Mines was the ?Load Analyzer" manufactured by Pactronics Corporation in California. This system consists of a meter-type analog readout for the cab console and three "Electrosensors" (linear variable differential transformers) attached to the main frame. Using conventional welding and brazing techniques, two sensors are attached between the front members of the engine sub-frame and the top of the left and right trailing arms (one sensor per arm). The same techniques are used to attach the third sensor to the rear centerline of the truck. between the main frame cross member and the A-frame. The readout unit on the cab console indicates by color-coding whether each sensor loading is under (green), at (yellow), or over (red) the pre-set rated load. The three-sensor configuration with in-dependent readouts also detects uneven distribution of the load.

Jan 1, 1984