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By John B. Shutack

In an effort to improve the performance of the explosive-anchored rock bolt, the Bureau of Mines studied the effect of the detonation velocity of the explosive and type of charge (linear or nonlinear) on the fracturing tendency of the anchor tube. Also, different types of tubing were compared to determine the one most suitable for use as an anchor tube. The Bruceton up-and-down method was applied in tests in which steel shooting blocks having boreholes ranging from 1-3/8 to 2-1/8 inches in diameter were used. Of the four types of tubing evaluated, a welded A-178 tubing that was drawn over a mandrel and renormalized gave the most promising results. The A-179 seamless tubing that is normally used with the explosive bolt was the most inferior. Both types, however, gave significantly better anchorage than the conventional shell when tested in both laboratory and underground experiments. Within the range of velocities studied (5,000 to 7,000 m/sec), the type of explosive showed no significant effect on anchor-tube performance. In general; linear explosive charges provided better anchorage than nonlinear charges.

Jan 1, 1968

Reduce the number of conveyor-related accidents in the mining industry by providing recommendations for effective operation of conveyor belt cleaning systems. The Problem Troughed belt conveyors provide effective, and efficient materials transport and their use has increased throughout the mining industry. Carryback, the spillage that accumulates under the belt, is a problem even with belt cleaning systems installed. Mining personnel are exposed to safety and health hazard; as they attempt to clean up this spillage. Approach To investigate the influence of metal-segmented-blade belt cleaners on the amount of carryback spillage, the. Bureau identified and tested the following parameters: (1) amount of pressure applied to the cleaner blades, (2) friction between the blades and conveyor belt, and (3) wear rates of different metal blades. The test parameters were quantified by recording the amount of carryback material passing the metal cleaner blade, and the weight loss of the blades. The data collected were used to identify the following critic d operational parameters of a segmented-blade conveyor cleaner system: (1) the existence of an optimal operating pressure for the conveyor and cleaning systems and (2) elimination of recessed patterns on the working side of the conveyor belt.

Jan 1, 1989

By Thomas M. Barczak

Purchasing a new set of longwall shields requires a substantial investment. A poor shield design can lead to economic hardships, safety concerns for the mine workers, and closure of the mine. This paper addresses several key points that should be considered in the procurement process: (1) understanding your goals and the logic in selecting a higher capacity shield, (2) the importance of completing performance testing before production shield fabrication begins, (3) making sure performance testing is properly done, (4) measuring load(stress) development during performance testing, (5) testing a shield to failure, and (6) the value of buying extra shields. In addition, several challenges are proposed for consideration in future shield designs. These include (1) 100,000 life-cycle expectancy, (2) improved hydraulic diagnostic capability, (3) smart load control by optimizing setting pressures, (4) lubricated link joint design concepts, (5) composite material applications to reduce shield weight, (6) incorporating periodic weighting predictive algorithms into the data-processing software, (7) advanced component load measurement on two to three specially instrumented shields to detect loading in the structural components, and (8) constant set leg cylinder design.

Jan 10, 2000

By M. J. Ferrante

The Bureau of Mines determined the enthalpies of Ni3S2 relative to 298.15 K to provide new thermodynamic data for advancement of mineral science and technology. Enthalpies were measured with a copper-block calorimeter from 298.15 to 1,197 K. A solid-solid transition was found at 834±1 K, with an isothermal heat absorption of 13.32 kcal/mol. The temperature of this solid-solid transition was accurately measured by differential scanning calorimetry. Because Ni3S2 does not melt congruently, the heat of fusion of 4.70 kcal/mol was obtained by extrapolation of enthalpy measurements from the solid and liquid phases to approximately 1,064 K. Tabulated values are given for the relative standard enthalpy, heat capacity, entropy, and Gibbs energy function between 298.15 and 1,800 K. Enthalpies are also given in equation form and combined with data from the literature to derive values of standard enthalpies of formation and Gibbs energies of formation. Similar calculations were made for the reaction of crystalline and liquid nickel with gaseous diatomic sulfur for Ni3S-2(c, l).

Jan 1, 1982

Objective To characterize the roof fall risk associated with underground limestone mining. Background During the first 5 years of this decade, more than one-fourth of the underground mining fatalities occurred as the result of fall of ground. The large openings in underground stone mines present limitations in observing ground conditions because of floor-to-roof heights ranging from 20 to 60 ft or more. Generally, the status of ground conditions in under-ground stone mines is based on observation and experience obtained during mining development. Some mines use monitoring instruments to gain information on roof conditions, but this practice is typically focused and localized to address issues in a particular area or section of the mine. The Roof Fall Risk Index (RFRI) described here provides a method of systematically characterizing the risk of roof falls to enhance the safety of mine workers. Approach As in many industries, a defective product signals the potential for failure. Similarly, defective rock or strata in an underground stone mine presents a potential risk. Assessing ground conditions relative to the concentration of defects observed provides a way to index or rate these conditions and the subsequent risk. Through observations at approximately 50 underground limestone mines in the United States, strata defects were categorized to develop a quantitative method for determining the RFRI. This method is specifically geared to underground lime-stone mines because strata defects are difficult to see and limited by the experience and knowledge of the observer. Overall, 10 defect categories were deter-mined: 3 that relate to geology, 4 that result from mine development, 2 that are indicative of the roof profile, and 1 that addresses moisture or water flow (see RFRI Chart 1). The defect categories are also weighted to reflect the conditions indicative of roof or strata movement or separation (categories 4 through 8). These weights are multiplied by the assessment value to determine the total category value. The benefit of using monitoring techniques to gain information on roof conditions is also factored into the RFRI for overall assessment of an area in the mine. Each defect category is assessed with a value of 1 to 5 based on parameters as defined in each defect category, with 5 indicating the highest degree or level of defect. The assessment value of 3 is also used when information on a parameter is unknown. The parameters for the defect categories are illustrated in RFRI Chart 2 for use in determining the assessed value. How It Works The RFRI provides a means to categorize the risk for roof falls in underground stone mines. The concept of this calculation is aimed at providing an indication of the level or degree of defect relative to the area of the mine examined. The RFRI for the calculation produces a distribution where RFRI values approaching 0 represent a stable condition and those ap-proaching 100 represent an unstable condition.

Oct 1, 2006

Although the technology involved in removing material from below the earth's surface has a long history, communication systems in underground workings are relatively new to the industry. Communication equipment did not begin appearing in underground mines until the early 1900's. Figure 1-1 shows a miner using a Western Electric standard telephone set for underground mines in 1913. These early phones were essentially the same as those used aboveground, except that they were enclosed in cast-iron boxes as protection against moisture, acid fumes, and gases. In the 1950's, the Chesapeake and Potomac Telephone Co. of West Virginia introduced a telephone set for use in explosive atmospheres that was designed around the philosophy of explosion containment. To contain an explosion within the telephone set required a 39-pound casing, which greatly limited portability. [ ] Both of these telephone sets required "pipe" or conduit, not a very practical item for a 100-squaremile coal mine. A common thread prevalent in the first 40 years of design and development is that the primary effort was placed on the telephone set. In the early 19701s, as work began on the design of modern communication systems for use in underground mine environments, the following requirements were established: 1. Must meet intrinsic safety standards .--This applies not only to coal mines, but to other explosive environments. 2. Must be compatible with the environment.--The system must with- stand dust, moisture, and corrosive conditions. 3. Must be rugged in structure.-- The system must withstand maintenance by 10-inch pliers and 4-pound hammers, as well as impact from a falling piece of roof. 4. Must be size-flexible.--Whatever is built must be sized for the small operator as well as the large companies. 5. Must be a total system in design.--The system must be intrinsically safe, including cable, power supply, and station set. 6. Must work with present telephone system.--The system should be compatible with aboveground communications already in place; it should not be the single cause of change in aboveground communications. 7. Value-added pricing.--The system should be reasonably priced so that savings created by its introduction will more than offset the installation cost.

Jan 1, 1984

By E. D. Gardner

This paper is one of a series on milling methods being published by the United States Bureau of Mines. Atlanta is in Elmore County, Idaho, about 80 miles from the railroad at Mountain Home, Idaho. The concentrator is at an elevation of, 6,000 feet and located in a timbered, well-watered region of relatively heavy snowfall. The snow, however, does not seriously affect local operations. The track road to Mountain Home, which passes over a mountain range, is closed each winter from about November 1 to June 15. During this period concentrates are stored and bullion is shipped by airplane. Power is generated by a Diesel engine and by an auxiliary hydroelectric plant. Gold is the principal metal produced; the or also contains an able quantity of silver.

Jan 1, 1935

By Jeanne A. Zimmer, Charles D. Taylor, Edward D. Thimons

On a continuous miner coal face methane concentrations are affected by the quantity of fresh intake air reaching the face. When mining the box cut on a 12.2-m (40-ft) two-pass extended cut face, the continuous miner is always at the point of deepest penetration, and operation of the machine-mounted dust scrubber and water spray system improve the flow of fresh air to the end of the box cut. However, when the continuous miner leaves the box cut to begin cutting the 40-ft slab, little is known about how much ventilation air reaches the end of the box cut. The Pittsburgh Research Laboratory (PRL) of the National Institute for Occupational Safety and Health conducted a full-scale study to answer this question. Tests were run to determine how much ventilation air reaches the end of the box cut with the continuous miner at three locations in the 40-fttwo-pass extended cut mining sequence; at the end of the 40-ft box cut, at the start of the 40-ft slab cut, and 6.1 m (20 ft) into the slab cut. During these tests, methane gas was uniformly released at the end of the 40-ft box cut. Methane concentrations measured0.3 m (1 ft) from the roof and face of the box cut were used to estimate face airflow quantities. Fresh air flow rates of 1.89and 4.72 m/sec (4,000 and 10.000 cfm), scrubber flow rates of 0, 4,000 and 10,000 cfm, water spray pressures of 0 and3827 kPa (0 and 120 psi), and a blowing curtain setback position of 15.2m (50 ft) were used during these tests. To improve ventilation to the end of the 40-ft box cut when the continuous miner was starting the 40-ft slab cut and 20 ft into the slab cut, the blowing curtain was advanced from its 50-ft setback to locations 40-ft and then 8.5 m (28 ft) from the face. KEYWORDS: Face ventilation, Extended cut, Methane, Box cut, Continuous miner, Scrubber

By C. M. Frost

Crude shale oil produced by gas combustion retorting of Green River oil shale was hydrocracked over a cobalt molybdate catalyst at operating pressures of 500, 1,000, 1,500, and 3,000 psig, an operating temperature of 8900 F, and a liquid hourly space velocity of 1.0. Hydrogen feed rate was maintained at 6,000 scf/bbl. The volume-percent yields of total liquid product were directly proportional to the operating pressure. Weight-percent conversion and volume-percent yields of gasoline were directly proportional to the log of the operating pressure. Catalyst deposit percentages were inversely proportional to the log of the operating pressure. Nitrogen and sulfur removal rates as well as saturation of the liquid products increased as the operating pressure was increased.

Jan 1, 1973

By B. T. Brady

Results of a study are presented of the effect of low radial reinforcement pressure on the postfailure characteristics of two types of granite, a sandstone, and marble. The low radial reinforcement pressure is applied to the rock surface by a gooch rubber sleeve. The gooch rubber sleeve provides a radial pressure of less than 1 psi pressure to the rock surface. This method of support gives an average strength increase of approximately 8 percent to the fractured rock sample.

Jan 1, 1973

By W. H. Forbes

The prevention of accidents from falls of roof and coal requires a considerable amount of personal judgment , for conditions are different even in adjoining working places in the same mine , and conditions may change from hour to hour even in the same working place as the coal is shot down and loaded . Experience has shown that many men can not or do not exercise enough care to protect themselves from injuries ; moreover , where the foremen have been made directly responsible for the safety of the men under them , accidents from falls of roof and coal have , in many instances , been reduced as much as 50 per cent . There is now no question that careful , constant , and intensive supervision is essential in preventing acci- { dents and more especially accidents of this class . Falls of roof and of coal account for approximately 50 per cent of all coal mine accidents and annually result in the death of approximately 1,000 men and cause lost - time injury to approximately 50,000 men . These and other accidents are recurring with little or no diminution , usually not because of the lack of knowledge of safe and dangerous practices , or of the technique of safety , but largely because of the lukewarm interest of major executives in safety or the failure of mine officials to formulate and to enforce common - sense rules and regulations . When failure is the outcome of a sincere attempt of the managing officials to teach safety to the miner , the fault may usually be attributed to misguided effort . Various solutions for the problem of eliminating accidents from falls of roof or coal have been suggested , some of which are : Standard system of timbering for the specified mining districts ; departure from the room-and - pillar system to a more concentrated system of mining ; more intensive supervision - that is , 1 section foreman for every 25 loaders instead of 1 section foreman for 75 , 80 , or even more loaders the practice in most mining districts . - However , investigations conducted by the writer at numerous mines operated by a number of mining companies in several districts show that employees and officials alike are inclined to work under loose and unsupported roof , relying solely on the " drummy " sound and on their judgment and experience as to whether the roof will stay in place for one hour or for several hours . With but few exceptions , men who had been injured by falls of roof or ccal or were found working under loose and unsupported roof , admitted that they knew of the possibly unsafe condition , but it was their opinion that the roof or coal was not loose enough to fall for some time .

May 1, 1931

By T. J. Ballard, Austin B. Clayton

"INTRODUCTION The Conrad Hill mine was first opened as a gold mine about 1832 and has been operated at various times since then.The Bureau of Mines, as part of the strategic minerals program, mapped and sampled the accessible mine workings.The: work failed to reveal appreciable copper values or evidence of additional ore deposits.ACKNOWLEDGMENTSIn its program of investigation of mineral deposits, the Bureau of Mines has as its primary objective the more effective utilization of our mineral resources to the and that they make the greatest possible contribution to national security and economy. The Mining Division, Lowell B. Moon, chief, conducts preliminary examinations, does the actual investigative work, and prepares the final report. The Metallurgical Division, Oliver C. Ralston, chief, analyzes samples and makes beneficiation tests.The field work embodied in this report was under the direction of Austin B. Clayton. The final report was prepared by T. J. Ballard uncle,- the general supervision of Mitchell H. Kline, chief, Raleigh, N. C., Branch, Mining Division. Analyses of samples and beneficiation tests were made by the Metallurgical Division. ACCESS AND OWNERSHIPThe property is 6 miles southeast of Lexington, N. C., and can be reached by U. S. Highway 64, turning off at Hedrick Grove Church, and is 1 mile north from the highway on the county road. The mine is in Davidson County at approximately 35° 47' latitude and 80° 10' longitude (fig. 1).The property comprises 471 acres and is owned (1943) by the McKeeMining Co., 512 Fifth Avenue, New York City. The property was optioned to Woodrow McKay of Lexington, N. C., in 1943, who in turn subleased the property to the Robert G. Lassiter interests of Raleigh, N. C."

May 1, 1948

By D. E. Kirby

In the preparation of beryllium-copper alloys by fused-salt electrolysis, beryllium was electrowon from beryllium oxide dissolved in fluoride electrolytes and deposited on copper cathodes. At 880° to 1,030° C, deposition was on solid cathodes. Beryllium deposited on the copper, forming a liquid alloy that dripped from the cathode. Dripping alloys containing 3 to 4 weight-percent beryllium were produced with as high as 90 percent cathode current efficiency, by electrolysis at 910°.C. To prepare alloys of higher beryllium content, attempts were made at depositing beryllium on liquid copper cathodes. These efforts were generally unsuccessful, although alloys containing as much as 12 percent beryllium were prepared by this method. At temperatures of 1,100° to 1,300° C, requisite to maintaining beryllium-copper compositions as liquids, the beryllium reacted with cell structures.

Jan 1, 1972

By Roy L. Wilfong

Surface area measurements were conducted on cadmium sulfide powders following thermal treatments at temperatures up to 500° C and treatment times up to 1 week. Thermal treatments at temperatures up to 300° C produced no significant changes in the surface area of the powders even after 1 week of exposure. At 500° C some decomposition of the cadmium sulfide occurred.

Jan 1, 1970

By G. B. Atkinson

The Bureau of Mines, in cooperation with McDermitt Mine, investigated a method for recovering mercury metal from mercury sulfide flotation concentrate by a hydrometallurgical technique. This preliminary report shows the feasibility of leaching mercury with a cupric chloride solution and electrowinning mercury from the solution. Proposed chemical reactions are shown in equations 1 through 3: 2CuC12 + HgS + 2CuCl + HgC12 + S°; (1) electrolysis HgC12 ?Hg + C12; (2) 2CuC1 + C12 ? 2CuC12. (3) A bench-scale leaching-electrowinning apparatus capable of treating 2.36 kg of sulfide concentrate in 11 days was operated continuously. An acid chloride solution containing, in grams per liter, 70 Ca, 60 Cu, 30 Fe, and 80 Hg was used as both leachant and electrolyte; 99.5 pct of the mercury was solubilized from the sulfide concentrate. Electrolysis of the leaching solution recovered mercury metal at an average current efficiency of 32 pct and an average energy consumption of 1.2 kwhr/lb Hg. Results from a preliminary economic evaluation indicated that a plant producing 31,500 flasks of mercury per year would have a capital cost of $3.7 million and an operating cost of $62 per flask of mercury.

Jan 1, 1983

By Thomas V. Falkie, R. Venkataramani

The method of analysis was to test a large model containing potentially significant variables and on the basis of the results obtained, to reduce the model to one containing only the essential significant variables. This stepwise procedure of "scanning" a large proposed model and subsequently condensing the model consisted of two cycles of the same series of computational procedures. Data Preparation A majority of the variables measured were continuous over a specific range of values; for example, the variable "Width Ratio" varied in practice from 0.25 to 1.00. However, "dummy" values were assigned to certain operational variables in order that these varia¬bles would have numerical expression. The following variables were described in this manner. Variable "Dummy" Value x3: Development/Pillaring 0, if pillar extraction was being done 1, if development mining was being done x4: Boxcut/Open Cut 0, if a box cut was being made 1, if an open cut was being made The variables under consideration are represented symbolically as follows: Y: Loading rate (tons per minuts): TONS/MIN x1: Ratio, width of cut/width of cutter: WDRATIO x2: Ratio, height of cut/diameter of cutter: HTRATIO x3: Development or pillaring: D/P

Jan 1, 1972

"There mines are located in out of the way places, where timber is scarce and its price high, the problem of protecting it from fire risk is of considerable practical importance. Scarcity of timber for mining purposes is commonly met with in the more arid parts of the country, where, so, mine timbers are most likely to become dry, and inflammable, and where, therefore, the risk from fire is accentuated. Any effective and cheap method of securing adequate protection of such timber at the lowest possible cost is worthy of careful investigation. The use of magnesite cement has been suggested in this connection.The necessary qualities for any substance for the use suggested above may be summarized as follows: It should be resistant to abrasion, and to impact and structural stresses. It should be durable when subjected to the action of the elements, and stable to any minor derangements of the base on which it is placed. It must be relatively unaffected by changes in temperature, by the action of water, and should adhere to the material on which it is placed, and it should also be free from shrinkage cracks due to setting up the material, Magnesite cement appears to fulfill the various conditions outlined above.The elasticity of such cement is the outstanding quality which makes it desirable for exterior work. It also possesses the ability to adhere to all materials in a marked degree. Its freedom from expansion and contraction after setting is one of the principal reasons for its adhesive quality. It may be applied to old brick or stone walls, even tho they have been repeatedly covered with paint, and the quality of adhearing to wood makes it possible to secure a weather-proof protection to the wooden casings and trim of exterior openings. The quality of adhesion and elasticity is easily proved by applying a 1/2"" layer to a 1"" board which has been first covered with metal lath. Such a board, about 8 feet long, may be repeatedly deflected about 3"" in the center without breaking the bond to the board, or causing any cracks to appear in the magnesia cement. The ability to resist fire as well as the possession of insulating qualities are also desirable characteristics of magnesia cement stucco.Magnesite cement stucco should be applied in two coats, each 1/4"" thick on various surfaces such as wood lath, metal lath, sheathing board, brick, hollow tile, or stone; care being taken as regards the presence of lime, unless the same has been rendered, or become inert as with very old, walls. The first coat made up to a consistency that works easily under the trowel is applied in the usual manner. No water should be added to the liquid chloride as it would alter its density, and injure the work. The second or finishing coat is then applied. The aggregates are different in each, and no other dry material should be added to them as they come to the trade. The cement may be applied in zero weather without damage. Setting is quite slow under such conditions, but the cement ultimately becomes hard. It should always be applied to a dry surface excepting that under certain conditions it is desirable to apply a solution of magnesium chloride before using the stucco."

Dec 1, 1919

By John C. Machisak

THE INJURY data and related employment experience of the coking industry in the United States for 1955 arc presented in this publication under the following general sections: (A) General Injury experience: Tables 1 to 9: Contain current and comparable data on operating statistics, number and classification of injuries, injury-frequency rates, and accident agencies involved in causing injuries. Data are tabulated for the industry as a whole, for each typo of oven: and for the States in which coke is produced.

Jan 1, 1958

By R. R. Sayers

"According to the dictionary,3 an accident is an event that takes place without one's foresight or expectation, often an undesigned, sudden, and unforeseen occurrence of an afflictive or unfortunate character. However, actual experience with accidents, the second most important cause of death in the United States,4 and a more extensive and scientific inventigation of their causes indicate that from 75 to 90 per cent of accidents are avoidable.5 This is borne out by the results of a three-year study of the accident records of 64,000 motor vehicle operators in New York6 which show that certain drivers are prone to accidents, that accident experience is traceable to certain causes, and that accidents don't just ""happen,"" It is estimated that 97,000 persons were killed by accidents in the United States in 1929 and 15,000,000 were injured. Of these about 23,000 deaths were due to so-called industrial accidents. In a list of accident rates by industries, compiled by the National Safety Council7 from reports of member establishments, mining heads the list, in both frequency (74.43 per 1,000,000 man-hours worked) and severity (9.9 days lost per 1,000 man-hours worked). These figures are very much higher than for any other industry; theindustries showing figures most comparable to mining are meat packing with a frequency rate of 55.94 and a severity rate of 1.47, construction with a frequency rate of 50.41 and a severity rate of 4.62, and refrigeration with a frequency rate of 43.35 and a severity rate of 3.04.While there is no way to determine the actual cost of all accidents from a monetary standpoint, estimates indicate that the cost will run into hundreds of millions of dollars annually.Cost of accidents in metal mines.- The United States Bureau of Mines obtains accurate figures on the number of lost-time nonfatal accidents as well as fatal accidents in the metal-mining industry.8 In 1926 the metal mines employed 127,823 men; 430 men were killed and 30,350 men received lost-time but nonfatal injuries as a result of accidents. Each of these fatal and nonfatal injuries averaged 114 days lost time, computing a fatality as 6,000 days lost time. This is equivalent to 3,523,944 days time lost due to accident, or at $5 per day per man, to a wage loss of $17,619,720. Deducting one-ninth of this figure for compensation paid, the metal miner loses in wage annually about $15,700,000, or about $122 per individual. This average loss per individual includes the cost of fatal accidents reckoned at 6,000 days or $30,000, but even if the loss in time due to fatal accidents be deducted, the average cost to each individual metal miner each year for nonfatal accidents is about $32. Estimating the loss to the metal-mine employer as about the same as that to the employee, the annual cost of metal-mine accidents to the employer is about $15,700,000 and to the employer and miner is $31,400,000."

Oct 1, 1930

By W. M. Weigel

Langley is about nine miles north of Augusta, Georgia., on the Sou them :l;.ilv1ay. According to the South Carolina. Geological Survey, tr.e district is known a.s tr. Horse Creek and Town Creek sub-areas of the Aileen or Savannah River Area f the South Ciil.Z'olina Coastal Plains. The geology and origin of the clays 3l"e Jscribea in a bulletin of the South Carolina Geological Survey (Series 4, Bull, l; "A Preliminary Report on Cl.:.ws of South Carolina.-), According to this report, i.he clays are sedimentary kaolins, in the Mid.dendor! and Hamburg series of the Cretaceous Age.

Aug 1, 1922