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By L. A. Stewart

This is one of a series of reports prepared by the Federal Bureau of Mines on mining methods, performance, and costs at individual mines in the United States. The primary purpose of these papers is to promote development and conservation of our mineral resources by disseminating technological information on methods, practices, and results that may be applied to the engineering analysis and solution of mining problems. This report discusses the highly selective mining methods of the Jaquays Mining Corp. at the Regal asbestos mine, 45 road miles northeast of Globe, Ariz. The milling method at the Globe plant is discussed briefly, and the extraction ratios of the various grades of fiber are tabulated. The proposed flowsheet of a new mill under construction is shown.

Jan 1, 1961

By Taghi Sharizadeh, MAURICE SUNKPAL, ELAHEH TALEBI, PEDRAM ROGHANCHI

Excessive heat exposure is among the health and safety hazards that have adverse effects on the mine workers. The very nature of the underground climate makes it a convenient avenue for heat generation and accumulation from several heat sources including autocompression, equipment, and heat from strata. This study investigated the effects of clothing insulation and acclimation on the thermal comfort ofmineworkers who are exposed to heat in underground working environments. Several clothing ensembles have been considered to analyze the effect of the thermal resistance of clothing on the thermal comfort of workers. To investigate the impact of the clothing insulation, the thermal comfort limit of an underground miner was calculated in the form of maximum allowable exposure time for an 8-h shift. Parametric analyses were subsequently performed using an already developed thermal comfort model using MATLAB software. The results demonstrate that clothing insulation has a significant effect on the thermal comfort of mine workers. Increase in the clothing insulation decreases the maximum allowable exposure time of mine workers. This is especially true for the new and non-acclimated workers. Since acclimatized workers can tolerate a higher level of the dehydration and are capable of losing more water by sweat per shift, the maximum allowable exposure time for the acclimated workers are higher than non-acclimated workers. These analyses have shown that acclimated workers can have more continuous work per shift compared with non-acclimated workers. For example, in an environmental condition with clothing that a nonacclimated worker cannot work more than 5 h, an acclimated worker can work a full 8-h shift.

Jun 29, 2020

By J. A. Russell

AS THE LOCAL electric generating plants increased in size and capacity and costs of producing electricity came down, it became apparent that it was no longer good economy to operate individual boiler plants at the Company's various collieries to power steam hoists and compressors. Great economies could be made by burning the Coal at central power plants and using electric power at the collieries. During the period after the end of the war, a steady and continuing programme of replacing first the smaller steam-driven units, and later the steam-driven compressors, was undertaken, and finally the larger hoists, those at No. 12 Colliery in New Waterford and No. 20 Colliery in Glace Bay, were the last units operated by steam.

Jan 1, 1958

By Ram V. Ramachandran

As an industry with major potential impacts (both good and bad) on water resources and due to the rising costs of process water, the mining and metallurgical processing industry has a key responsibility to contribute to the overall goal of preserving both the quantity and quality of water available within the watersheds they operate in. As we plan for new and existing mining and metallurgical projects, we need to consider a hierarchy of options to minimize our impact within the watersheds we operate; keeping clean water clean by diverting around our operations; reduction in the amount of water we use; reuse of water wherever possible, and improvements to water quality for recycling and/or release to environment. Technology is needed to treat water to improve its quality for reuse or release to the original watershed. While lime precipitation followed by sulfide polishing is effective for removal of most of the metals, it does little or nothing to remove ammonium ions and anions such as chloride, nitrate, molybdate, arsenate and selenite/selenite. Regulations on sulfate limits in treated discharge water has drawn close attention to the treatment of acid mine drainage (AMD). Management of brine and salt disposal is also posing a problem due to strict regulations and cost of disposal. This paper will: (a) address issues associated with managing water with the objective of maintaining water sheds—status quo-during and after operation (b) discuss potential technologies for advanced water treatment plants.

1. Domestic Production and Use: In 1982, 13 companies operated 30 primary aluminum reduction plants, with 3 firms accounting for 55% of production. Washington, Oregon, and Montana accounted for 34% of the production; Texas, Louisiana, Missouri, and Arkansas 18%; Tennessee, Alabama, South Carolina, and North Carolina 14%; other central and eastern States 34%. Based on published producer prices, output of primary metal in 1982 was valued at over $5.5 billion. Aluminum consumption by 25,000 firms was centered in the East Central United States. Packaging accounted for an estimated 39% of domestic consumption in 1982, transportation 20%, building 14%, electrical 8%, consumer durables 8%, and other uses 11%.

Jan 1, 1983

By Albert L. Toenges

ALMOST all of the coal used for metallurgical purposes west of the Continental Divide is ruined from the Lower Sunnyside bed in Carbon County, Utah. Coke produced from Lower Sunnyside coal and used in western steel plants is weaker structurally than coke used elsewhere in this country. For this reason, western coal operators and steel manufacturers are anxious to develop new areas of coking coal of better quality than the coal produced from the Sunnyside bed. Early in the war, the Bureau of Mines undertook an intensive investigation of western coals to (a) develop additional reserves of coking coal to supplement proved reserves in the Sunnyside bed and (b) develop, if possible, reserves of coal with coking qualities superior to Lower Sunnyside coal. Results of investigations in the Kemmerer district, Wyoming,9 and in the Minnesota Creek area, Gunnison County, Colo.,10 have been published.

Jan 1, 1952

By McElroy. G. E.

"INTRODUCTION The ventilation of development faces in mines for diluting gas or dust concentrations and, in hot mines for providing air-motion cooling effect, by fan pipe units discharging high-velocity streams of air close to the face, is an important phase of mine ventilation. The effectiveness of such units depends largely on the air-flow distribution between the pipe discharge and the face, a subject on which little definite information is available. A study of the air flow from grills into rooms 3/ has provided some basic data, but under conditions no different from mine fan-pipe installations in development ends that the data are of doubtful application to mine installations, owing to differences in air-flow conditions.The purpose of this report is to present and discuss the results of an initial series of tests of a general investigation designed to yield definite laws and design data for the discharge of fan-pipe lines in mines. This series was planned to yield preliminary data on the effects of velocity of discharge, size and shape of pipe, and adjacent walls on the air-flow distribution. Tents were made on 10-inch pine with seven different end conditions and one change of wall condition. Changes in size and shape of discharge wore made by short transformation pieces directly at the discharge end, with the result that effects due to contraction and expansion of the flow, which proved to be of much greater importance than was expected, totally obscured any normal effects due to size and shape. The results of this series of tests, therefore, relate to 10-inch pipe with various discharge designs and are so reported."

Aug 1, 1940

By John A. Gann

WITHIN a very few years magnesium has sprung from oblivion, from classification as a technically unknown, little appreciated, and expensive material to front-page importance in many fields of engineering. Bunsen may be considered the real founder of our present magnesium industry. In 1852; he prepared this metal by the electrolysis of fused magnesium chloride in a porcelain crucible. In the following years the electro¬lytic process became recognized as superior to chemical reduction methods and Germany became the world's chief producer, installing the first electrolytic magnesium plant in 1886. ' Thirty years later, national defense requirements, occasioned by the outbreak of the World War and the cutting off of imports, caused France, England, and the United States to begin production. Most of this magnesium was used for military purposes so the close of the war resulted in a large curtailment in production. A few companies, however, encouraged by the early success of the Germans and by the rapid developments taking place in aeronautics, and confident in the certain future and ultimate value of this metal, undertook extensive re¬searches in its production and utilization. A variety of production methods, both chemical and electrochemical, were investigated and some of them operated on a commercial scale. Reduction of the oxide with carbon at high temperatures and the electrolysis of the oxide in a fluoride

Jan 1, 1932

By Albert L. Toenges

This paper is the fourth of a series5/ that describes some methods and practices of mining coal mechanically where operations are conducted on more than one shift. Many inquiries have been received regarding methods of extracting pillars with mechanized equipment, and this paper deals primarily with this phase of mechanical mining. Some of the questions that have been asked are:

Jan 1, 1942

By C. A. Herbert

During the fiscal year ended September 30, 1932, there were 27 fatalities in the coal mines of Indiana. Four of this number occurred in small mines employing less than 10 men, over which the State Inspection Department has no jurisdiction. In addition to these coal mine fatalities, an explosives accident in a clay mine resulted fatally to one man. Tables 1 and 2, which follow, list the fatalities as to cause and occupation. Table 1. - Fatal accidents by cause, fiscal year ended September 30, 1932 [ ]

Jan 1, 1933

By Jesús E. Gómez

Verification and production testing of soil nails requires debonding the upper section of the soil nail to isolate the portion of the soil nail intended for testing within a specific ground type. In pre-drilled soil nails, debonding can be accomplished relatively easily because the bar can be fitted with a debonding sheath before its installation. However, debonding of hollow core bar soil nails requires special procedures to achieve the intended result. This paper is a summary of the Federal Highway Administration (FHWA) Hollow Bar Soil Nail (HBSN) Test Program, which is described in detail in FHWA-CFL/TD-10-001. The program was developed to establish a suitable debonding procedure to test hollow core bars in soil nailing applications, and to initiate a database on typical bond values of HBSNs. The investigation showed that it is possible to debond hollow core bars for testing. It was also found that hollow core bars installed in granular soils may present ultimate bond values that are significantly higher than those typically used for design of conventional pre-drilled soil nails.

Jan 1, 2011

By Mehdi Asadi, Heinz Palkowski

"In recent years, the increased demand for advanced high-strength steels (AHSS) mainly had been driven by the need of the automotive industry to reduce weight and to improve safety. Beside good ductility and high strength, those steels have a high bake hardening (BH) effect, giving additional contribution to the strength of structural parts, subjected to the paint baking process. In this paper we concentrate results gained for hot rolled dual phase (DP) steels.For the simulation of changing process conditions within the final hot rolling the specimens were hot deformed using different schedules of temperatures and reductions, selected according to the non-recrystallization temperature (TnRX). It was possible to refine the DP steel structure by control-ling the deformation temperature and the amount of strain below TnRX during the thermo-mechanical controlled processing (TMCP). This structure refinement resulted in an improvement of the strength and BH behaviour. A wide spectrum of mechanical properties could be obtained depending on the different hot deformation schedules. The best strength and BH levels were recorded for the deformation below TnRX at the highest amount of strain.IntroductionDP steels are characterized by a good formability, high strength and a good compromise between strength and ductility [1]. Moreover, the DP steels exhibit a continuous yielding behaviour, low yield point and a high strain-hardening coefficient [2]. Furthermore, the DP steels often show a large potential for bake hardening (BH). BH refers to the increase in yield strength as a result of the paint baking treatment of the shaped auto-body parts. The primary mechanism that causes the additional strengthening is the immobilization of dislocations by the segregation of interstitial atoms, known as classical static strain aging [3]. The increase of strength thus achieved allows a further reduction of sheet thickness and improves the crash safety and the dent resistance. The BH of special steel qualities is technically used in DP, where e.g. the increase in strength is realized in the final heat treatment [4]. Previous own investigations [5-6] stated that the BH effect of DP is much stronger than that one for conventional BH steels."

Jan 1, 2012

By D. Brandt

The Dos Pobres Mine is a copper oxide heap-leach operation located near Safford, AZ. Mine production is 163 kt/d (180,000 stpd) with a stripping ratio of 2:1. During the startup phase of the mine, the required fragmentation size was easy to achieve as the ore was highly weathered and fractured near the surface. As the mine depth increased, the proper fragmentation size became harder to achieve due to less weathering and different rock types. Because copper recovery increases as crusher product size decreases, achieving the appropriate product size at the crusher is the operating priority. The engineering department began a project to improve fragmentation through blasting and thus increase throughput in the crushing system. It soon became clear that input from other departments was going to be required to make this a safe and economical project. Therefore, the crush and convey and ore control departments worked together to evaluate all available data on blast patterns and explosive loads and tracked performance and results. Using image analysis, drill energy index and databases, improved fragmentation size has been achieved and results have been verified while controlling costs.

Jan 1, 2011

By Randy R. Williams, Roberto R. Olivera, Brian W. Wilson, Marina S. W. Li

"An extensive, deep mixing (DM) ground improvement program using the Cutter Soil Mixing (CSM) technique was designed and constructed in 2013 to support up to 20 m (66 ft) high Mechanically Stabilized Earth (MSE) walls and bulk earthworks for the development of the Kitimat Liquefied Natural Gas (LNG) facility at Bish Cove, British Columbia. The deep mixing program consisted of overlapping rectangular panels of in situ cement treated soils, constructed using the CSM technique, to form barrettes covering an area of approximately 12,900 m2 (15,428 yd2) and providing suitable static and seismic stability of foundation soils for support of the proposed MSE walls, which would in turn support LNG facilities. The deep mixing program comprised of 1645 CSM panels that were constructed through varying thicknesses and discontinuous layers of very soft fine-grained marine deposits and/or loose sandy silt to silty sand deposits, embedding typically 2 m ( 6.6 ft) into dense silty sand to sand and gravel. CSM panel depths ranged from 5 m (16 ft) to 30 m (98 ft) and were constructed by treating approximately 73,900 m3 (96,658 yd3) of soil yielding average unconfined compressive strengths exceeding 2.5MPa (363 psi).This paper outlines the approach adopted for the design of the CSM ground improvement program in the highly variable subsurface conditions present at the site. A phased sitespecific investigation consisting of boreholes, test pits, Cone Penetration Tests, piston tube sampling, advanced laboratory testing and groundwater monitoring was implemented. In addition, soil-cement bench scale testing was completed, followed by installation of 8 CSM panels for field trial purposes. Extensive geotechnical static and seismic analyses including Limit Equilibrium stability analyses, liquefaction assessments, and advanced numerical modeling were undertaken to optimize the design and meet the project-specific design performance criteria. The Design-Build team worked closely to develop and implement Quality Assurance and Quality Control plans and managed significant construction challenges while achieving suitable panel termination depths."

Jan 1, 2017

By Arno C. Fieldner, J. W. Paul, WALTER A. SELVIG

Many mine samples of coal are analyzed each year in the laboratories of the Bureau of Mines. The analyses are made in connection with investigations relating to fuels belonging to or for the use of the United States Government, the causes of accidents in coal mines, the geologic relations of coal beds, and the quality and the value of the coal and lignite in the public lands. The systematic collection and analysis of such samples in connection with fuel investigations was begun in 1904 at the Government fuel-testing plant at the Louisiana Purchase Exposition. and was under the direction of the United States Geological Survey for six years. Descriptions of the coal samples collected between the beginning of this work, July 1, 1904, and the transfer of the fuel-testing and mine-accident investigations to the Bureau of Mines,July 1, 1910, were compiled and published in Bureau of Mines Bulletin 22.' Descriptions and analyses of samples collected during the fiscal years 1911 to 1913 were published in Bureau of Mines Bulletin 85.2 Descriptions and analyses of samples collected during the fiscal years 1913 to 1916 were published in Bureau of Mines Bulletin 123.3 The present bulletin presents analyses and descriptions of samples collected during the fiscal years 1916 to 1919. In order that the material in this bulletin may be used to supplement that presented in Bulletins 22, 85, and ,123, the same plan of geographical classi- fication has been followed, the analyses and descriptions of the samples being grouped in alphabetical order according to the State, county, and town near which the mines or prospects sampled are situated.

Jan 1, 1922

By Carleton C. Long, George F. Weaton

Two years ago the general features of the St. Joseph Lead Company's zine-smelting process were described.' At that time the discussion was limited to a description of the production of high-purity electrothermic zinc oxide. The present paper will describe the commercial production of electrothermic zinc metal. Since a fairly complete description of the plant of the Josephtowlz smelter was given in the earlier paper, the brief recapitulation given here will be limited to mention of additional equip-ment recently installed and a bare outline of the preparatory process. The raw material is zinc sulphide concentrate received from the company's mines in St. Lawrence County, New York. An average dry-weight analysis of this material is: Zn, 58.28 per cent; Pb, 0.56; insoluble, 1.6; SiO², 1.12; Fe, 6.4; CaO, 0.47; MgO, 0.27; S, 31.9; Cd, 0.1; Cu, 0.06; Mn, 0.24. Moisture content is usually about 3 per cent. Three 21-ft. 6-in. diameter, 12-hearth Nichols-Herreshoff furnaces roast the coacentrate to a calcine of average conlposition (first 10 months of 1938): Zn, 67.69 per cent; Pb, 0.06; insoluble, 1.75; SiO², 1.26; Fe, 7.24; CaO, 0.47; MgO, 0.36; S, 1.88; Cd, 0.1; Cu, 0.07'; Mn, 0.27. The sulphur dioxide gas, after passing through a waste-heat boiler, proceeds through a recently redesigned low-resistance gas-purification assembly to the contact sulphuric acid plant, of which the capacity is 200 tons of concentrates per day. The calcine is mixed with a portion of returned furnace residues and sintered on 42-in, by 44-ft. Dwight-Lloyd sintering machines, gas-fired. Wind-box gases pass through air coolers before filtering in a 1440-bag Dracco unit, and thence to stack. The Dracco fume is mixed with Cottrell dust, from roaster gas cleaning, and sent to the leach plant, where it is roasted (SO² to acid plant) and leached with sulphuric acid to recover, as end products, lead sulphate, zinc sulphate and cadmium.

Jan 1, 1943

By George F. Weaton, Carleton C. Long

Two years ago the general features of the St. Joseph Lead Company's zine-smelting process were described.' At that time the discussion was limited to a description of the production of high-purity electrothermic zinc oxide. The present paper will describe the commercial production of electrothermic zinc metal. Since a fairly complete description of the plant of the Josephtowlz smelter was given in the earlier paper, the brief recapitulation given here will be limited to mention of additional equip-ment recently installed and a bare outline of the preparatory process. The raw material is zinc sulphide concentrate received from the company's mines in St. Lawrence County, New York. An average dry-weight analysis of this material is: Zn, 58.28 per cent; Pb, 0.56; insoluble, 1.6; SiO², 1.12; Fe, 6.4; CaO, 0.47; MgO, 0.27; S, 31.9; Cd, 0.1; Cu, 0.06; Mn, 0.24. Moisture content is usually about 3 per cent. Three 21-ft. 6-in. diameter, 12-hearth Nichols-Herreshoff furnaces roast the coacentrate to a calcine of average conlposition (first 10 months of 1938): Zn, 67.69 per cent; Pb, 0.06; insoluble, 1.75; SiO², 1.26; Fe, 7.24; CaO, 0.47; MgO, 0.36; S, 1.88; Cd, 0.1; Cu, 0.07'; Mn, 0.27. The sulphur dioxide gas, after passing through a waste-heat boiler, proceeds through a recently redesigned low-resistance gas-purification assembly to the contact sulphuric acid plant, of which the capacity is 200 tons of concentrates per day. The calcine is mixed with a portion of returned furnace residues and sintered on 42-in, by 44-ft. Dwight-Lloyd sintering machines, gas-fired. Wind-box gases pass through air coolers before filtering in a 1440-bag Dracco unit, and thence to stack. The Dracco fume is mixed with Cottrell dust, from roaster gas cleaning, and sent to the leach plant, where it is roasted (SO² to acid plant) and leached with sulphuric acid to recover, as end products, lead sulphate, zinc sulphate and cadmium.

Jan 1, 1943

By V 9. 0 / 300 dpi

ATHOUGH turquoise mining was, so far as we know, the first large, well-organized mining operation,' gem mining, from the Roman con-quest of Egypt until the opening of the South African pipe diamond mines some 50 years ago, was largely carried on' by individuals 01' small companies in a relatively primitive way. Now" however" gem mining does not yield in efficiency or in scale of operations to the mining of other mineral products. In 1926 the diamond mines of the world milled some 9,000,000 cu. yd. and produced 0,43 cu. yd. diamonds, a remarkable example of concentration. The Premier Diamond Mining Co. up to Oct. 31, 1927, had milled 116,455,550 loads (each of 16 cu. ft.) from which 26,-217,144 carats of diamonds" worth £29,897,971 2s, 9d., were recovered. The Burma Ruby Mines, Ltd., operating sever-al alluvial mines in Burma in 1911-12, mined and milled 1,479,845 truckloads of gemmiferous earth for 15c. per load. We of the industry can there-fore, with good grace, disclose some of the myths and also describe some of the primitive mining methods of the, dark ages of our industry. The methods, so far as I know, were not patented, and if they were, the patents have expired, so that engineers in other branches of the art are free to use them if they see fit.

Jan 1, 1928

The goal of this presentation is to firstly, review mine effluent permits in a number of regulatory jurisdictions; and secondly, to develop a strategy for incorporating the regulatory requirements into the environmental assessment (EA) process for new mine proposals. The focus was on the province of British Columbia (BC) and their recent initiative on the One Project, One Process initiative. This task was approached by reviewing mine effluent permits for various regulatory jurisdictions and comparing them to BC mine effluent permits. This comparison revealed a variety of approaches for designing mine effluent permits, but there is a common scientific approach, which supports the conclusion that the permitting of mine site discharges can be designed based on a scientific approach. This approach is expected to: (a) achieve sound environmental protection, (b) demonstrate the application of Best Achievable Technology, (c) incorporate sound sustainability approaches, (d) incorporate regulatory objectives, and (e) create efficiencies for the regulator and mining industry. There is an essential precursor to developing a sciencebased effluent permit for which the proponent must provide: (1) the necessary technical assessment information to demonstrate the mine effluent permit application will result in an environmentally protective effluent discharge, to the satisfaction of the permitting agency; (2) the necessary level of detail for essential mitigation pollution control works to demonstrate how they will be operated to achieve the necessary treatment goals; (3) the necessary level of detail for non-treatment mitigation (passive treatment, or passive pollution control); and (4) the necessary level of detail to demonstrate how pollutants will be prevented from being mobilized (or being generated) from ore and waste materials, or, if the pollutant originates from an added milling reagent, how the reagent will be reduced. While (active) mitigation pollution control equipment is more definable in an effluent permit, non-treatment mitigation is more nebulous in terms of defining in a mine effluent permit; it is therefore recommended that appropriate management plans be developed for essential mitigation pollution control and non-treatment mitigation (passive treatment) and submitted with the proponent?s EA/effluent permit application packages. Major deficits in adequate active and passive pollution control will inevitably lead to regulatory impediments to developing a permit and allowing a mine to be developed, although EA processes strive to preclude major permit information gaps; the onus is on the proponent to provide complete information in this area. Hence the importance of EA processes ensuring that the major technical information requirements for effluent permit development are included in the EA phase. Recently, this aspect of the mine EA process has taken on a new role under the political efforts of One Project, One Review Process (Canadian) and One Project, One Process (in BC). These initiatives translate, in part, to ?synchronous permitting? (in BC): effluent permitting, for example, may be initiated shortly after the proponent submits the EA application to Ministry of Environment.

Aug 1, 2013

By Woodcock J. T, Zhu T, Sparrow G. J

A lead-zinc-pyrite flotation middling was pressure oxidised to give residues containing iron hydroxy sulphates (a mixture of basic iron sulphate (BIS) and a hydrated iron sulphate (HIS)) and lead sulphate. Direct thiourea leaching of this residue gave over 90 per cent extraction of gold and silver, but thiourea consumption was high (56 kg/t) because of soluble iron(M) in the residue. Attempts were made to decrease the thiourea consumption by removing the water-soluble iron(M) or reducing it to iron(II). Water washing of the pressure oxidation residue removed most of the hydrated iron sulphate but little of the basic iron sulphate. Iron removal was only 12-20 per cent, depending on the number of washes, from one sample which was low in HIS, and 40-45 per cent in another which was high in HIS. Iron extraction was not appreciably increased by using dilute sulphuric acid (with or without additions of sodium sulphite), dilute phosphoric acid, or sodium chloride solution (150 g/L NaCI). Direct thiourea leaching of the water-washed residue gave over 90 per cent extraction of gold but only 49 per cent of silver, with a thiourea consumption of 27 kg/t. However, the application of a water wash followed by reduction of dissolved iron(III) with sodium sulphite in the presence of thiourea, and then thiourea leaching resulted in over 98 per cent extraction of gold and silver for a thiourea consumption of 26 kg/t.

Jan 1, 1991