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By Edward A. Barrett, Kathleen M. Kowalski-Trakofler

This paper discusses the application of a training intervention that uses degraded images for improving the hazard recognition skills of miners. Method: NIOSH researchers, in an extensive literature review, identified fundamental psychological principles on perception that may be employed to enhance the ability of miners to recognize and respond to hazards in their dangerous work environment. Three studies were conducted to evaluate the effectiveness of the degraded image training intervention. A model of hazard recognition was developed to guide the study. Results: In the first study, miners from Pennsylvania, West Virginia and Alabama, who were taught with the aid of degraded images, scored significantly better on follow—up hazard recognition performance measures than those trained using traditional instructional methodologies. The second and third studies investigated the effectiveness of the training intervention at two mining companies. Data collected over a3—year period showed that lost—time injuries at mines in Alabama and Illinois declined soon after the training intervention was instituted. Impact on Industry: Further exploration of the hazard recognition model and the development of other interventions based on the model could support the validity of the steps in the hazard recognition model.

By Liane J. Terrill

Rockbolt failures at the Waste Isolation Pilot Plant have been recorded since 1990 and are categorized in terms of mode of failure. The failures are evaluated in terms of physical location of installation within the mine, local excavation geometry and stratigraphy, proximity to other excavations or shafts, and excavation age. The database of failures has revealed discrete areas of the mine containing relatively large numbers of failures. The results of metallurgical analyses and standard rockbolt load testing have generally been in agreement with the in situ evaluations.

Jan 1, 1995

By M Logsdon, B Johnson, M Fordham

The underground Neves Corvo copper-tin mine in southern Portugal currently disposes of its tailings in a ætraditionalÆ wet tailings impoundment. With the development of new technologies for preparation and transport of thickened tailings or paste, alternative tailings disposal options have become available that provide a more sustainable water management system and a higher degree of environmental protection. In addition, growing European resistance towards subaqueous disposal may well result in mandatory use of dry disposal methods in the near future. Twenty-four tailings disposal alternatives were evaluated through a cost-effective and easily-implementable bench-scale geochemical screening program conducted at site. Since the tailings are pyrite-rich and highly acid generating, particular emphasis was placed on monitoring likely indicators of sulfide oxidation under the various simulated disposal scenarios. Dependent parameters investigated as part of the program included moisture content, use of alkaline additives (lime, cement), use of a bactericide, and agitation. Response parameters included paste pH, specific conductance, and temperature as indicators of oxidation. The testing was carried out for 30 weeks. Results indicate that a number of relationships between the geochemical stability of the samples (ie sulfide oxidation) and sample characteristics can be readily identified. Based on the results of the screening program, a limited number of tailings mixtures have been selected for further long-term testing in field weathering cells to aid in the identification of the optimal mix design for future tailings disposal at the Neves Corvo Mine. This type of testing provides an inexpensive and reliable methodology for semi-quantitative comparison between tailings disposal alternatives, in particular at locations where more rigorous evaluation through long-term, fully replicated testing is not immediately feasible.

Jan 1, 2003

By W. Greuer

The objective of this paper is to demonstrate how site-specific slope stability analysis can be performed entirely from the desktop with satellite images, Geographic Information Systems (GIS), and slope stability software packages. This type of analysis would effectively eliminate or restrict the need for expensive field testing of geotechnical properties of a hillside, pit wall, berm, or embankment, and would be particularly useful for slope failure mitigation at remote or otherwise inaccessible locations.

Jul 17, 2000

By D. Harrington

"One of the underlying reasons for the establishment of the United States Bureau of Mines was the desire to promote greater safety in the mining and allied industries. Safety continues to be an important feature in the work of the bureau, and during the fiscal year 1929-30 the Bureau of Mines allotted $387,510 to the safety division out of a total appropriation of $2,249,670 (other than for helium).The safety division personnel consists of 22 engineers, 2 mine safety commissioners, 3 senior foreman miners, 18 foreman miners, 16 first-aid miners, 16 clerks, 10 cooks, and 3 laborers, a total of 90 persons. These safety workers are distributed through the 10 districts into which the United States has been divided for administration purposes. During the year 3 engineers, 1 senior foreman miner, 4 foreman miners, and 4 first-aid miners left the service. In general these men left the service to take up more lucrative positions.The 10 all-steel mine-rescue cars have headquarters as follows: Car 1, Reno, Nev.; Car 2, Raton, N. Mex.; Car 3, Pittsburgh, Pa.; Car 4, Terre Haute, Ind.; Car 5, Pineville, Ky.; Car 6, Pittsburg, Kans.; Car 7, Huntington, W. Va.; Car 8, Duluth, Minn.; Car 9, Butte, Mont.; Car 10, Des Moines, Iowa. There is also a mine-rescue car with headquarters at Juneau, Alaska. In addition, there have been established 11 mine rescue stations as follows: Birmingham, Ala.; Berkeley, Calif.; Seattle, Wash.; Evansville, Ind.; Jellico, Tenn.; McAlester, Okla.; Norton, Va.; Vincennes, Ind.; Wilkes Barre, Pa.; Salt, Lake City, Utah; and Pittsburgh, Pa.Every mine-rescue car and station is fully supplied with equipment and devices for use in case of mine fire or explosion, and at each station is an up-to-date automobile truck. At time of amine disaster the bureau's forces are immediately placed at the disposal of the authorities, State or company, with a view to giving such assistance as may be desired or available.The activities of the safety division at ordinary times are directed largely to securing and disseminating data on safety in mining. Among the numerous studies made, or being made, concerning safety and health in mining are those in connection with the causes and methods of prevention of mine explosions, explosibility of dust and gases, explosives and blasting, haulage, electricity, falls of persons or material, ventilation, dusts, high temperatures, timbering, lighting, machinery, and in fact practically anything which results or may result in acci¬dents of any kind in or around mines."

Nov 1, 1930

By J. J. Beresford

The primary function of a support slurry is to provide stability during excavation. Today foundation structures can be designed around the additional friction afforded by the absence of the filter cake attributable to the use of polymeric fluids as support slurries. These fluids use pseudoplastic fluid viscosity to advantage over the fragile thixotropic gel structures of bentonite. Many of today's polymeric fluids are environmentally acceptable for disposal directly to municipal sewers.

Jan 1, 1989

By A. Davis

Pit lake chemistry depends on the lithology and hydrologic characteristics of the ultimate pit surface, the availability of buffering capacity in the wall rock versus oxidizable sulfide content, and the alkalinity of recharging ground water. In this paper, we assess the geochemistry of 11 existing pit lakes, ranging from the acidic Red Hill and Berkeley pit lakes (pH 2.6 to 2.8) to the alkaline Yerington Pit lake (pH 8.45). Predicted pit lake chemistry is similar to that in existing pits with the long-term pit lake chemistry lying on a geochemical continuum between end members consisting of influent calcium bicarbonate ground water, and naturally occurring sulfate-rich evaporite playa lakes.

Jan 1, 1996

By Ashley P, Neale T

Alluvial gold was discovered on Sudest Island in 1888, and by 1893, an official production of 10 000 oz had been won by expatriate miners from shallow alluvial and colluvial workings. Hardrock sources were located and while one was developed at Mt Adelaide, production is unknown. Sudest Island comprises mostly phyllites and lesser coarse grained rocks which are cut by metamorphic quartz veins and basic (locally shoshonitic) intrusives. Sheared and altered basalts and serpent- inites represent are believed to be elements of an ophiolitic suite inferred to be associated with metamorphic core complex development and thus similar to tectonic environments on Misima and the D'Entrec- asteaux Islands. No epithermal-style quartz suggestive of a target similar to that on Misima Island has been identified on Sudest. A hardrock source for the alluvial and colluvial gold is interpreted to be metamorphic quartz veins similar to those in the Bendigo- Ballarat Zone of central Victoria, possibly remobilising gold derived from the meta- sedimentary and meta-igneous rock pile. The present focus is on development of a small scale alluvial gold mining operation in conjunction with exploration for hardrock targets. This latter programme remains at a reconnaissance stage.

Jan 1, 1991

Approximation of Surface Area of Fines in Blast Induced Fragmentation The surface area of fragmentation, especially in the finer size fractions, is a useful property to characterise the mechanisms of rock breakage. In the past, researchers have made assumptions with regard to the shape of fragments in order to estimate surface area, such as that all fragments are the same shape or all fragments are spheres. Often these assumptions were developed with inappropriate measures and returned considerable variability and error for the calculation of surface area. This paper reports a model to estimate fragmentation surface area from sieve sized raw data and a density measurement, with experimental validation down to a size fraction of -45 + 38 ?m in a range of rock types. After characterisation of particle shape, a range of fragment shapes was found to be in each size fraction. Nevertheless, in the ?2 sieve series the average fragment mass, volume and the size fraction standard deviations could be predicted reasonably accurately. Ellipsoid geometry was used to describe particle shape. Using these parameters and the number of fragments (also predicted) in each sieve size the individual surface area of each fragment was simulated.

Sep 13, 2010

By Rose W. J

Geophysical prospecting was first recognized in Australia as an important aid to mining enterprises in 1927, when, on the suggestion of Sir Herbert Gepp, the "Imperial Geophysical Experimental Survey" was organized. The function of this survey was to carry out investigations on the applicability and efficacy of the most important geophysical methods under Australian conditions. The survey was formed under the joint auspices of the British Empire Marketing Board and Australian Commonwealth Government, and the investigations were carried out in Australia under the leadership of Mr. A. Broughton Edge, B.Sc., A.R.S.M., F.G.S., M.Inst.M.M., during the years 1928 and 1929. A description of the work performed, the apparatus used, the principles of the methods, and the results and conclusions of the survey are included in a report published by the Cambridge University Press in 1931.In 1932 Sir Herbert Gepp became interested in some gold mines in Eastern Victoria, in the Cassilis and Bethanga districts, and realizing that the lodes in these areas would be favorabIe for a geophysical examination, he arranged for a series of tests to be carried out. This work was performed by Mr. E. L. Blazey, B.E.E., a former member of the staff of the Imperial Geophysical Experimental, Survey, under the general direction of, and in consultation with, Mr. W. J. Rose, B.M.E.The work was commenced in October, 1932, at Cassilis, Vie., continued at Talgarno, Vie., .and concluded at...

Jan 1, 1933

A study of the accessory opaque iron-titanium oxide minerals of the Oiles Complex, central Australia, was an Important complement to an investigation of the magnetic properties of the Complex. These minerals constitute no more than 10 per cent of the rocks sampled, and generally less than 1 per cent of the ultramafic rocks ("magnetite" concentrations were not sampled for magnetic investigation).Titanomagnetite is the main opaque mineral present, although ilmenite exsolution has been widely developed. High-temperature oxidation of these produced mainly titanohematite, pseudobrookite and rutile. Both exsolution and high-temperature oxidation developed along specific crystallographic directions (along {111}), and in Irregular patches. Low-temperature oxidation is generally not well-developed, indicating the small degree of weathering in the rocks sampled.Although the Oiles Complex is layered, no apparent overall preferred orientation of the iron-titanium oxide minerals was observed, other than crystallographically controlled alignment of very fine grains within pyroxene crystals. Other textural features of the opaque oxides are restricted to symplectic intergrowths mainly with the ferromagnesian silicates.

Jan 1, 1971

By J. Chlopas, D. J. Barratt, M. V. Sharp

"This paper describes the modifications made to the Fox Mine copper concentrate bin in order to obtain a continuous flow condition. The design of these modifications is based on test results obtained on Fox Mine concentrate using the flow factor testing machine in the research facilities of Cominco at Trail, B.C.The original design consisted of a cylindrical steel bin on a concrete slab with two slot-discharge openings. Based on predictions of the maximum mass flow hopper angle for conical and wedge shaped hoppers for the concentrate on various potential bin lining materials, Ultra-High Molecular Weight (UHMW) Polymer was selected as a liner.Using the recommended angles for conical and wedge shaped hoppers, the profile of the bin base was changed to suit the existing slot openings.The initial method of fastening the liner resulted in failure of the connections. Subsequent modification to the zinc concentrate bin fastener system proved satisfactory and it has now been in operation for over one year without any separation of the lining material.Utilization of flow testing procedures on concentrate samples has demonstrated that proper design of storage bins or hoppers to achieve continuous flow can be incorporated into the design of new plants."

Jan 1, 1978

By L. W. Saperstein

Public Law (PL) 95-87 contains explicit demands for land use planning in its reclamation and designation sections. Both sections require recognition of local land use plans; however, they are not well integrated. Accommodation to these planning requirements affords an opportunity to create better links among local land use planners, the state's permitting authority and designation office, and mine designers. Mining companies should take a more positive interest in local land use plans and in resource inventories. States should consider handing down to local government the responsibility for accepting mine permit applications and designation of unsuitability petitions.

Jan 12, 1983

By M. K. Larson

The U.S. Bureau of Mines conducted a study to evaluate support interaction between conventional bolts and the rib at two Western coal mines. The principal goal was to develop guidelines for designing rib support under various conditions. At mine A, seven rib bolts were installed with load cells and monitored over a 4-month period. Fifty-seven pull tests were conducted with two types of anchors at various depths. At mine B, 163 pull tests were conducted. Results showed that during the 4 months of monitoring bolt load, tension in the bolts decreased over time in some situations and that anchorage capacity was reduced in a yielded zone. Load-versus-deformation plots from pull tests in the rib showed fundamental differences in anchorage behavior at the two mines, particularly when comparing results of tests in old pillars at mine A and tests in new ribs at mine B. Numerical modeling experiments indicate that coal bonding strength was the most significant factor in determining postyield anchorage capacity. Correlation of anchor yield point with anchor depth were more pronounced at mine A than at mine B; measuring torque may be a reasonable method of monitoring deteriorating anchorage capacity in a rib undergoing fracturing and yield.

Jan 1, 1996

Supplement to Proceedings Number 120, December, 1940. Drawings accompanying paper by T M Carey 'The Surface Plant at the Zinc Corporation Mine, Broken Hill'

Jan 1, 1940

By Andrew Bamber

Between the early 20th Century and the present, the production of mined copper ore has increased from tens of millions of tonnes per annum to several billion tonnes per annum, an unprecedented increase. Contemporaneously, the average grade of mined copper ore has declined from over 2% to approximately 0.5%. Again, in parallel, the price of copper has increased dramatically to reflect the increasing cost of extracting and beneficiating such large amounts of low-grade ore. Within the period the copper price however exhibited several periods of intense volatility, driven by interrelationships between demand and supply, themselves strongly driven by periodic industry developments in the fields of exploration, ore extraction, mineral processing and extractive metallurgy. It is notable that these periods of intense innovation exhibit a strong correlation to preceding periods of price, demand or production volatility. This paper examines key trends in each theme between 1900 and 2018 and draws strong conclusions as to the interrelationship between them. INTRODUCTION The production of copper metal has increased from around half a million tonnes per annum to over fifteen million tonnes per annum over the course of the 20th Century. Contemporaneously, the average grade of copper ore mined has declined from over 2% to approximately 0.5% copper by weight. As a result, the volume of Run-of-Mine (ROM) copper ore required has increased from around ten million tonnes per annum to over a five billion tonnes per annum over the period, an unprecedented increase. The adage is that demand drives supply, that prices rise in times of high demand and/or short supply, and that prices fall when supply catches up with demand. The market sellers in biblical times knew this well; mediaeval market sellers equally. Dutch tulip sellers in the 1600’s experienced both in extremis — much to their dismay — and the great industrialists and early capitalists of the industrial revolution knew this also. Mining recently went through a cycle where investors and producers both believed that the “old” rules did not apply any more and that the industry was in a “supercycle” where prices would continue to rise on the back of increasing demand. However, as the GFC of 2008–2011 showed, this was not to be. Another adage says that we constantly innovate to lower costs of production. Is this true? Do we only innovate to lower cost? Has the true cost of producing copper actually gone down over the century? Or are there other balancing forces at play? Herb Kellogg, in his Centenary address to the American Institute of Mining Engineers (now SME), categorized two types of innovation: ‘Bigger and Better’, and ‘New Process’ (Kellogg, 1970). ‘Bigger and Better’ allows increased production at presumably lower cost; ‘New Process’ supports the same result at presumably the same scale, but at significantly enhanced efficiencies and (therefore) lower cost. It is suggested that both types of innovation are constantly and interchangeably required to achieve the kind of scale-up, at achievable operating and capital cost that the industry has achieved over the period. What can be shown, over this last Century and a bit, is that supply and demand are indeed inextricably linked, price and demand are linked, but so are price and cost of production. And the innovation required to sustain that weaves in between them all, cost-effectively increasing production capacity or lowering the cost of production at the same throughput, and sometimes both.

Jan 1, 2019

[Chairman ZAY JEFFRIES, Vice-chairman S. SKOWRONSKI, Chairman WILLIAM M, CORSE, Secretary Executive Committee J. R. FREEMAN, JR., Local Section G. E. JOHNSON R. S. ARCHER, Non-ferrous Data Sheets P. D. MERICA EDGAR C. BAIN, Journal Committee L. W. SPRING SAM TOUR, Papers Committee F. L. WOLF H. C. JENNISON, Membership Committee Local Section Committee J. R. FREEMAN, JR., Chairman R. J. ANDERSON, Cincinnati R L. DOWDELL, Minneapolis R. S. WILLIAMS, Boston G. H. CLAMER, Philadelphia W. A. SCHEUCH, Chicago R. F, Woon, New York Papers Committee S. SKOWRONSKI, Chairman F. H. DIx, JR.-Aluminum E. R. DARBY-Bronze C. S. WITH ERELL-Copper H. A. BEDWORTH-Brass W. K. FRANK-BRONZE F. W. HARRIS-Copper R. L. SUHL-Nickel H. M. ST. JOHN--Foundry Metallurgy R. S. DEAN-Theoretical Metallurgy]

Jan 1, 1928

Leadership "Organizing a group of people to achieve a common goal" Paradigms ? Rush to produce or cut costs. ? The most important thing after safety is production. ? We?ll take care of ??this?? once in production. ? I know what to do, 30 years of experience in mining. ? I want a 100% reliable operation.

Aug 1, 2013

By P. Scinto, A. Festa

"Orway Mineral Consultants Ltd (OMC) as a consultancy has specialized in comminution circuit design for over 30 years. The OMC Power Model is OMC’s in-house software for specific grinding energy power-based circuit calculations. The model uses ore hardness parameters from testwork, circuit configuration, equipment geometry and operating conditions as inputs. The model combines proprietary and published power models to predict the total specific energy and the specific energy for each of the components of the comminution circuit. This makes the software a valuable tool for the design and validation of JKSimMet modelling or actual plant performance. Proprietary calculations have been developed and reconciled with over 100 plant benchmark projects encompassing a variety of comminution circuit configurations.This paper describes the OMC approach for determining power requirements for the design of grinding circuits. The accuracy of the OMC Power Model software is confirmed through plant survey benchmark performance and comparisons are made with other power modelling techniques. Practical examples are presented, showing applications for single stage SAG (SS-SAG), SAB and SABC circuit configurations.IntroductionConventional low throughput grinding options, consisting of multi-stage crushing followed by rod and ball mills, have almost become obsolete in the move to higher milling rates required to process lower metal grades. Grinding circuits with fully autogenous (AG) or semi-autogenous (SAG) mills, often combined with ball mills are now the standard to meet the high capacity requirement for processing ores with lower metal grades. Simulation software and power-based equation methodologies have been established for designing AG/SAG based grinding circuits and predicting their energy requirements.Presently, power-based equations remain the principal method for designers to size mills and estimate the power requirement for comminution circuits. The technique is simple and it relies on the design to specify a feed and product size plus a hardness index to produce a specific energy value (kWh/t). The desired throughput of the mill is multiplied by the specific energy to give the mill power requirement, which is used in turn to size a mill for dimensions, internal geometry and loading conditions to match the power draw."

Jan 1, 2015

By Cyril Wells, Paul E. Busby, Donald P. Hart

EARLY workers in the field have established that the diffusion of nitrogen follows normal diffusion laws. Concentration-penetration data from layer analyses of reasonably pure iron specimens nitrided in an atmosphere of CO-NH, are given in the literature,' and were utilized to calculate diffusion constants for nitrogen. The diffusion of nitrogen in y-iron may be represented approximately by the equation D - 1.07x10-' e-:".'m'"'r The diflusion coefficients, D, given by this relation represent mean values, since carburization occurred simultaneously with nitriding. The rate of diffusion in a-iron has been determined from calculations based on data obtained by measuring the internal friction of samples at each of several frequencies and temperatures.' The expression for the diffusion of nitrogen in a-iron derived from these experiments is D= 1.4x10-1 e-it tudent Increasing carbon content decreases the diffusion of nitrogen in a-iron.3 Oxygen, either dissolved or combined, has been shown' to exert a marked influence on the diffusion of nitrogen from a nitriding atmosphere. D values increase rapidly with increasing amount of dissolved oxide in the metal, but diffusion is completely inhibited if preliminary oxidation has produced the slightest trace of iron oxide on the surface. Diffusion coefficients have also been calculated from permeability and solubility data,' but these values are of questionable accuracy since the boundary conditions assumed may not actually exist. Experimental Method The method selected for determining the diffusion coefficient of nitrogen in a-iron in the present investigation was the so-called fractional saturation method. This method consists of heating the specimen in an atmosphere of H2-NH8 mixtures of known composition for a predetermined time. Following this treatment, the specimen is analyzed for nitrogen; and the diffusion coefficient is calculated from a known relationship of the fractional saturation, time of diffusion, and dimensions of the specimen. This relation is based on a solution of Fick's second law and assumes that the surface of the specimen immediately reaches a given concentration which remains constant during the diffusion anneal. Tabular solutions are available4-5 and provide the relationship between fractional saturation and the quantity \/Dt /a, where D is the diffusion coefficient in cm2 per sec, t is the time of the diffusion anneal in seconds, and a, in the present experiments, is the half-thickness of the slab specimen in cm. The method is adequately discussed elsewhere.9-10 Material The material for this study consisted of 3/32 in. thick (0.24 cm) carbonyl iron plate. The major impurities in this iron are oxygen and carbon, which are readily removed by long time high temperature treatment in an atmosphere of purified hydrogen. In order to reduce the diffusion anneal time for saturation runs, some of this material was rolled to a thickness of 0.015 in. (0.038 cm) prior to the hydrogen treatment. Hydrogen treatments were carried out in a hydrogen-purified Armco iron tube heated in a global

Jan 1, 1957