Search Documents

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 R. S. Dean

[During t!~c past f isc.~l y~:-.r the Xetzllvz-gicd 3ivi oion has made im- portant contribution-, don,- tlie lines rlictnteil by the Bureau policy of con- servation of domestic rnker~l rcsoul-ces throui;h use. During t!~c past f isc.~l y~:-.r the Xetzllvz-gicd 3ivi oion has made im- portant contribution-, don,- tlie lines rlictnteil by the Bureau policy of con- servation of domestic rnker~l rcsoul-ces throui;h use.]

Jan 1, 1940

By R. S. Dean

[During the past fiscal year the Metallurgical Division has made important contributions along the lines dictated by the Bureau policy of conservation of domestic mineral resources through Its investigations include determining of fundamental data of metallurgically important materials improvement of standard ore-treatment method to eliminate waste and improve the grade or recovery of the products; and development o new metallurgical processes.]

Jan 1, 1940

By David C. Oyler, Dennis R. Dolinar, Christopher Mark, Craig S. Compton

Fully-grouted roof bolts comprise more than 80% of the primary roof supports used in U.S. coal mines. However, nearly 1,500 MSHA reportable, non-injury roof falls occur each year, and most of these are attributable to failure of the roof bolt system. Anchorage failure is one failure mechanism for fully-grouted bolts. As roof deformation works its way upward, the bolts can become heavily loaded near their upper ends. If the applied load exceeds the anchorage, the bolts will simply pull out. Research dating back 30 yrs indicates that this type of anchorage failure is most likely when the roof rock is weak, just where roof support is most critical. In soft shale or coal roof, the small amount of data available in the literature indicates that 20-30 in of resin anchorage may be required to achieve the full capacity of the bolt. In other words, the “full resistance zone” of a 60 in bolt may actually be just 30-40 in. Despite its potential importance, there is no widely accepted anchorage test for fully grouted bolts. Standard pull tests have sometimes been employed, but they provide no information on the anchorage near the top of the bolt. An alternative, first described more than 25 yrs ago in the U.S., is the short-encapsulation pull test (SEPT). With this test, the bolt is installed with only a short (1 ft or less) tube of resin. In recent years variations of this test have become international standards. v

By B. Numerov

A numerical proceeding of corrections with respect to topography is worked out for places near the station and a graphical one for the remote parts of the terrain. The description is clear and very detailed. The numerical proceeding is very similar to that of the Eötvös and Schweydar methods which are in use already. In both methods the levelling is accomplished in certain circles and fixed azimuths around the station. While, in using the Eötvös method, the height differences are directly inserted into the final formulas with regard to the foot of the instrument and in case of accurate formulas derived by Schweydar the height differences are calculated with respect to the center of gravity of the attachment, Numerov introduces functions A and B, the heights, of the terrain and of the center of gravity, which are to be inserted into the formulas.

Jan 1, 1929

Objective To develop safety training tools for mine safety and health professionals that fill gaps in currently available materials. Background Over the past 4 years, NIOSH has developed and released a series of training videos targeted at underground mine safety. These videos were developed in close cooperation with the mining industry and safety training professionals and were intended to provide realistic, interesting, effective training to both new hires and experienced miners. The videos cover the mining cycle, starting with the breaking of the rock (Handling Explosives in Modern Mines), making certain the work area is safe after a blast (Rock Falls— Preventing Rock Fall Injuries in Underground Mines), installation of ground supports and recognition of failed supports (Miner Mike Saves the Day), and hauling both ore and materials in the mine (Hazards in Motion). A separate video (Hidden Scars) captures the story of a fatal rock burst through an interview with a miner who nearly died and who lost his partner in the accident. These videos have been extremely popular in the mining industry and are being used across the nation as well as in 18 other countries. Two of them, Rock Falls and Miner Mike Saves the Day, have won Alice Hamilton awards, NIOSH’s highest recognition for Excellence in Occupational Safety and Health, in the Educational Materials category. Approach A recently completed independent evaluation of the SRL mining videos identified five themes that mine safety trainers consider necessary for successful training materials: high quality, credibility, accurate content, effectiveness, and an engaging nature. All of the NIOSH videos were rated very high for each theme. Using these measures of effectiveness, a more difficult challenge has been undertaken. Much has been written in the mining industry about the coming turnover in employment. The average age of miners in the United States is approaching 50, and many of the “experts” will be leaving the industry as they reach retirement age. The mining industry, like many industries involving skilled labor, relies heavily on a master-apprentice relationship between new hires and older miners. New hands learn from experienced miners. Their learning falls into three primary categories—they learn what’s important to know, they learn how to do things correctly and safely, and they learn the culture.

Mar 1, 2002

By Edward J. Caldwell

In 1930 the Bureau of Mines published a description3/ f the concentrating process of the Magma Copper Co. as it had been developed to that time and was then employed. The prevent paper revises the description to the end of 1943 and includes a description of the smelting processes. Data from the mild and smelter operations for, 1942 are given in tables 1 to 14, Billing costs are included, but it is not permissible to publish smelter costs.

Jan 1, 1944

By John R. Bartels

This Bureau of Mines report describes the development of a preliminary design for a novel, protected, cab-shuttle car for use in working seam heights down to 40 in. Because of the severe restrictions imposed by low-coal operation, Mine Safety and Health Administration (MSHA) regulations only require canopy protection on shuttle cars operating in seam heights of 42 in or greater. MSHA routinely grants variances for canopy use in seams 48 in high or less. The design was generated by giving the operator needs equal priority as related to machine performance parameters. Cab-shuttle car concepts that led to the recommended design are described, along with criteria and testing used to evaluate their potential effectiveness.

Jan 1, 1988

By D. V. Gaffney

This report presents the results of a three-phase investigation of the feasibility of using cemented coal waste in active underground coal mines to prevent mine subsidence. Coal caste was broadly defines to include waste from both coal preparation plants and coal-fired power plants. Underground systems with potential for cemented backfill are determined through identification of current mining methods and available technologies for cemented backfilling and consideration of materials and systems constraints upon mining and backfilling integration technical concerns including operation, health and safety, materials utilization, environmental impact, applicability and techni¬cal status are assessed for six (6) systems which were identified as potentially feasible. Two (2) of these systems are conceptualized in a hypothetical case history and developed to further assess their operational and economic feasibility. Comparisons of costs among both systems and with costs of surface disposal are presented. Also presented are the conclusions reached concerning the present feasibility of using cemented backfill in active coal mines and potential avenues for further research. Appendices contain further consideration for design of cemented mixtures to be used in backfilling, the results of laboratory testing of mixtures rich varying concentrations c: waste products and cementing agents, and an annealed bibliography

Jan 1, 1981

By J. Douglas Causey

In 1984, at the request of Bureau of Land Management, the U.S. Bureau of Mines studied the 13,730-acre Burnt Lodge (MT-065-278) Wilderness Study Area (WSA) in order to evaluate the mineral resources. The WSA is located in Phillips and Valley Counties, MT, about 55 miles south of Malta adjoining the Charles M. Russell National Wildlife Refuge. It is underlain by near horizontal sedimentary beds of the Hell Creek Formation, Fox Hills Sandstone, and Bearpaw Shale. No mines or prospects were found in the WSA. Bentonitic beds were examined, but are generally less than 1 ft thick and, except where exposed, are deeply buried. Kaolin-bearing sandstone in the Colgate Member of the Foxhills Sandstone was sampled and may have potential for use in structural clay products. However, remoteness and inaccessibility of the area make it unlikely that the material could be economically extracted even though some of the material is marginally useful.

Jan 1, 1985

By C. H. Kats, E. G. Meiter, J. J. Bloomfield

"INTRODUCTION.Street manholes for underground power and signal circuits, and for sewers vary in size from small openings that a workman can enter, to large central vaults. As the gas hazards in manholes are similar to those in mines, the mine safety committee of the Bureau of Mines was asked whether the miners' flame safety lamp is a practicable device for indicating inflammable gases in manholes.The coal miner detects firedamp (methane or CH4) with his flame safety lamp by observing the luminous can or the increased length of flame that appears when 1.5 per cent or more of the combustible gas is present. The lamp flame also warns him of lack of oxygen by going out when the oxygen becomes less than 16 to L7 per cent. A tentative rule regarding the proposed use of safety lamps at manholes has been suggested previously as follows: ""Testing for gas: Do not enter manholes until you have .assured yourself that the manholes are free from dangerous gases, by testing with an approved safety lamp, by ventilation, or by other adequate methods."" The Bureau's committee recommended that miners' flame safety lamps are not suitable,testing in manholes because hydrogen gas may be formed therein. The flame safety lamps approved by the Bureau of Mines for use in aseous mines are safe in presence of methane and while not tested against the members of the paraffin hydrocarbon series (CnH2n+2) higher than methane they been have been used therein; the lamps should not be used in hydrogen or acetylene they can initiate explosions in presence of these gases."

Oct 1, 1925

By Rudolf E. Greuer

This report contains a theoretical study on the usefulness of fighting mine fires with inert gases. Four introductory chapters review the role of fires in the mining industry; the conventional fire fighting techniques, the influence of inert gases on the burning of gases and solids, and the well established inert gas fire extinguishing systems for buildings and machinery. A large chapter describes the use which inert gas has found so far for fighting mine fires in major mining countries. More applications have occurred than generally expected and the inert gas method has in the last years become routine in Czechoslovakia and the USSR. This chapter is based on an evaluation of the extensive literature as well as on personal information obtained from leading experts.

Jan 1, 2012

By D. J. Dollhopf

An overburden handling dragline operation was conducted at a strip mine near Colstrip, Montana. The objective of the demonstration was to determine the degree a dragline could mix overburden inhibitory zones during the spoiling process. In order to quantify the mixing process the overburden was: 1) scatter spoiled; 2) dump spoiled; and 3) spoiled in a normal manner. This demonstration was part of an effort to investigate means of reestablishing non-polluted hydrological systems in areas where surface mining directly impacts shallow ground-water resources. When inhibitory material constituted < 5% of the total overburden volume, such material was essentially not detectable in the resultant spoil pile. When inhibitory material exceeded 15% of the overburden volume it was always found in the resultant spoil pile, generally with a smaller extent, indicating partial dilution. Inhibitory overburden zones often reappeared at the base or top of spoil piles in the potential aquifer and root zones suggesting the need for selective placement procedures. Scatter spoiling did not appear to be significantly more effective than normal spoiling, however clump spoiling seemed to produce less mixing of inimical zones. Mining activity has reduced ground-water levels within 1600 m of the active pit. The aquifer which is developing in spoils appears to have similar confinement characteristics as those that existed in the premined aquifer. To date no definite trends have been observed concerning changes in surface or ground-water chemistry associated with mining.

Jan 1, 1978

By Stephen C. Tadolini, Peter (Yunqing) Zhang

A unique circumstance created by monitoring a pre-driven longwall recovery room permitted measuring the stresses of a coal pillar throughout its entire life cycle in less than a week. A fender pillar, created in approximately the middle of a longwall panel at a depth of 650 ft, transformed from a solid barrier pillar - to a yielding pillar - to a residual pillar as 3 ft slices were methodically removed with the longwall shearer. The complete transformation, or life cycle, took place in less than 12 hours. The stresses were quickly transferred from the pillar onto the standing pumpable concrete supports and into the outby pillars. Roof to floor closure measurements, combined with the timing of the pillar behavior, provides a detailed look at the uncontrollable convergence of underground mine openings. Pillars remain the most important form of “primary support” and understanding these life cycles is vital for safe and efficient mine design, in both room and pillar and longwall panel extractions.

By V. R. Spironello

The Bureau of Mines is involved in research directed toward the utilization of municipal solid waste. One of the primary objectives is the recycling of the magnetic fraction of municipal solid waste (raw refuse). This is consistent with one of the Bureau&apos;s goals, which is to minimize the requirements for mineral commodities by maximizing metals recovery from secondary domestic resources, In this investigation, cupola trials were made using combinations of refuse scrap with shredded automobile scrap under basic and acid slag practices. Furnace operating information and the behavior of alloying and tramp elements was obtained. The research showed that it is possible to utilize up to 60 percent refuse scrap in the cupola under basic practice and 30 percent under acid practice, Aluminum in refuse scrap, present in bimetallic cans, increased the recoveries of silicon and manganese charged to the cupola. Increased use of refuse scrap provided iron of lower sulfur. The alumina resulting from oxidation increased the slag volume. The aluminum and tin contents of the iron increased with increasing levels of refuse scrap in the charge. Lead was not a problem with respect to contamination of the iron. In basic practice, operation of the cupola was satisfactory since all slags were adequately fluid. In acid practice, operation became troublesome above the 45-percent level. Under both practices, the cupola iron melting rate decreased. Particulates in scrubber water and stack condensate samples contained lead, zinc, and tin, and the dust load increased. The tensile and transverse strengths of the iron produced under both practices are reported.

Jan 1, 1980

By John J. Henn

An economic evaluation was made of the Bureau of Mines incinerator residue separation process to update the one published in 1971 to a mid-1.974 cost basis. This evaluation includes a modification of the glass-recovery operation used In the original plant design as well as a unit to bale the ferrous scrap product before shipment. Products from the plant are ferrous metal, aluminum and copper-zinc scrap, and a mixed-color glass cullet. The estimated capital and operating costs for four plant sizes are as follows : [ ] The operating costs do not include credits for recovered metals and glass.

Jan 1, 1975

By K. Thirumalai

The feasibility of breaking hard rock by thermal weakening and subsequent mechanical impacting was tested, Field tests at a basalt quarry in Dresser, Wis., demonstrated that large hard-rock blocks (5 by 5 by 5 ft) could be effectively reduced in size by this method. A series of secondary breakage tests, using air-propane heating, resulted in an average heating cost of $0.36 per ton. This combination energy approach provides an alternate means of size reduction of large hard-rock boulders for quarry and for underground are chute and grizzly operations where mechanical methods, used alone, are ineffective.

Jan 1, 1975

By G. S. Rice, H. O. Howarth, Greenwald. H. P.

"Electricity as a means of transmitting power was introduced in coal mines of the United States about 1891 for both haulage and coal cutting, only a few years after electrical haulage was first installed on street cars. At first, the hazards of electrical ignition ere not appreciated and the general use of open lights and black blasting powder made the possibility of electrical ignition of fire damp seem small by comparison.5 The possibility of igniting coal dust directly by electric sparks or arcs had not been considered.In the nine years prior to 1900, the progress of introducing electricity underground was slow, except that it was used for lighting shaft bottoms. Electricity had to contend with the more general use of compressed air for coal-cutting machines and, in mainline haulage, with the use of mule-teams and endless and tail-rope systems. During this period, great improvements were made in electrical apparatus, in surface installations as well as in mines. By 1900, therefore, the technique of applying electricity in mines had so far developed that it rapidly drove out the use of compressed air in coal mines because of the comparative ease of its installation and the marked economic advantages of its use over that of compressed air in practically all kinds of mining machines except percussion. rock drills.At this time, however, little thought was given to the hazard of indiscriminate use of electricity causing explosions and fires, and the possibility of electricity causing direct ignition of coal dust had not been considered. On December6, 1907, however, the Monongah, West Va. mine explosion occurred. This was the greatest mine disaster this country has over had, 361 men being killed. Although: there was much diversity of opinion, as to the origin of the explosion, many well-informed engineers concluded that the cause was ignition of coal dust do to wreckage of a runaway trip in a rope haulage entrance slope. This slope also served as a ventilating and it extended from the level of the coal bed to the surface. The runaway trip tore down electric cables and stirred up a cloud of dust when it was wrecked at the bottom of the slope. The dust was probably ignited by the electric arcing that occurred when the wires were knocked down. (Another but less probable source of ignition was the open flames of miners lamps.) From the foot of the slops the explosion appeared to have travelled into the mine, spreading as it went. If this ignition of coal dust was the origin of the explosion, it occurred on pure intake air entering directly from the surface."

Oct 1, 1930

8186. Aquilina, C. Determinazioni relative di gravita eseguite nel 1939 (Relative Determinations of Gravity Made in 1939), Ric. Ingegn, Rome, vol. 10, 1942, pp, 6376. In 1939, the author made a survey of relative gravity in the Mediterranean-basin, 41° 35,5&apos; to 41° 56.3&apos; northern latitude and 0° 10,3&apos; to 0° 28&apos; longitude, referred to Mt. Mario. Ten stations were covered; 25 other stations had been occupied in the vicinity of Rome and in Albanian mountains during previous years. The new measurements revealed gravity anomalies ranging within the following limits: (1) With a free-air reduction, between + 29 and + 120 mgl., average + 60 mgl.; and (2) with a Bouguer reduction, between + 22 and + 87 mgl., average + 41.5 mgl. - V. S.

Jan 1, 1946

By Jaffrey Y. Foley

Placer gold and platinum-group metals (PGM) In the Slate Creek area, northern Chistochina Mining District, are derived from silicified and carbonate-altered, Cretaceous-Tertiary (?) dikes and related mafic and ultramafic rocks. Gold In the placer deposits is also derived from Tertiary conglomerate and altered Jurassic-Cretaceous argilllte. The argilllte and Igneous bedrock sources show a unique Ag-Au-Cu-Hg-PGM association that Is attributed to a combination of magmatic, hydrothermal, and supergene processes. The bedrock and placer deposits contain a variety of gold silver-lead-copper and PGM alloys. The observed geochemical, mineralogic, and alteration features are consistent with current geological models for turbidite-hosted gold deposits. No economic gold or PGM lodes were Identified, but several areas warranting more detailed investigation were Identified. Conventional geochemical sampling and analysis of small sample splits weighing between 10 and 30 grams were shown to be unreliable as a consistent means of identifying gold- and PGM-bearing bedrock. Pan concentration of pulverized rock samples was shown to be a more effective, but still Imperfect means of Identifying gold and PGM source rocks. Geochemical data and mineralogic examination indicate that gold and PGM in potential lode sources are very finely and erratically disseminated, are chemically combined with other elements, and may not be entirely free-milling or amenable to simple gravity concentration.

Jan 1, 1990