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By Richard M. Winar

In September of 1988, the United States Environmental Protection Agency (US EPA) published final regulations that effect owners and operators of underground storage tanks (USTs). These regulation include stringent technical requirements for owners and operators of USTs containing petroleum and certain hazardous substances. The technical requirements effect both new (those installed after December 1988) and existing USTs and include notification; performance and operating standards; release detection, reporting, investigation; release response and corrective action; and closure requirements for out-of-service USTs. The USEPA also published final regulations regarding financial responsibility requirements for owners or operators of petroleum USTs in October of 1988. These regulations require owners or operators of petroleum USTs to demonstrate financial responsibility for the costs of corrective action and compensating third parties for bodily injury or property damage caused by releases from petroleum USTs. Like thousands of other industries, the mining industry must become thoroughly familiar with these new regulations in order to properly manage their UST systems, ensure compliance, and understand other ramifications imposed by these regulations.

Jan 1, 1990

By William Crowe Kellogg

By now the term AFMAG has been fairly well introduced to the mining industry. Several papers have been presented and published concerning AFMAG (references 1 and 2). However, the term AFMAG remains rather new to the industry, and the exact place which AFMAG occupies in the minerals exploration field is still somewhat uncertain in the minds of many who may have use for this new and interesting method of electromagnetic exploration for conductive minerals. It is the purpose of this paper to again review the basic theory of the AFMAG system, to describe and - illustrate the equipment and data presentation as used in aerial surveying, and to provide some idea of the cost and nature of airborne AFMAG surveys. This paper is primarily of a general descriptive nature and presents AFMAG theory in a schematic rather than mathematical form; emphasis is placed on the equipment, the present method of data presentation and survey operations, since it is believed that such, information will now generally best serve to inform the mining industry of the characteristics of the AFMAG system.

Jan 1, 1960

By A. R. Pasieka, J. C. Wilson

"Compressed air still remains an essential source of power in underground hardrock mining. As a power medium it has a high convenience factor, although, in many cases, it is expensive to generate and maintain. Until new methods of breaking rock and/ or satisfactory alternatives to air-driven rock drills are developed, compressed air is indispensable. It is essential to utilize this source of power to a maximum by keeping all pneumatic tools supplied with air in adequate quantity and at optimum pressure.In search of the optimum operating air pressure, con-trolled drilling tests were undertaken at dynamic pressures ranging from 90 to 140 psi.Test results show the effect of increasing air pressure on: the performance of various jackleg drills at varying pressures, the life cost of drill steels and TC bits, and drilling efficiency at varying pressures. From the fore-going, a criterion is proposed for selecting efficient rock drills, and, also, optimum air pressure at each mine can be determined for its particular conditions."

Jan 1, 1966

By R. E. Perry

To develop new or improved flotation methods for beneficiating low-grade red iron ores, the Bureau made many bench scale and continuous anionic flotation tests on ferruginous sandstones and calcareous iron ores from the north-eastern part of the Birmingham, Ala., iron district. The ores contained about 21 to 36 percent iron, a to 11 percent lime, and 22 to 64 percent silica. Production of iron concentrates containing 48 to 52 percent iron was technically feasible in batch tests, and the iron recoveries ranged from 85 to 92 percent. In continuous tests 49 to 51 percent iron products were obtained with 80 to 90 percent recovery.

Jan 1, 1962

By R. D. Doepker

The dissolution of metals from various metal-mine tailings has been studied through a series of submerged tailings batch tests at U.S. Bureau of Mines laboratories. Tailings selected for this study included both reactive (as determined by column leaching studies) and benign tailings. Metal mobilization in aerated (97 pct saturated with oxygen) samples appeared to be consistently greater than metal dissolution from stagnant (79 pct or 70 pct saturated with oxygen) waters. In the early weeks of the study, sulfate concentrations increased with solid-liquid contact time, while calcium dissolution remained nearly constant. Metal concentrations varied in relation to solid-liquid contact time and aeration. Leachate pH appeared to be nearly independent of solid-liquid contact time for basic tailings. Reactive tailings having both a submerged and a surface component demonstrated increased metal mobility.

Jan 1, 1990

By D. M. Hausen

Applications of process mineralogy to the exploitation of molybdenum ores begins with early exploration and continues through the evaluation of mineralized prospects, to amenability testwork and pilot plant studies, to the development of a mineable orebody and to the development of a viable process for recovery. After plant construction, a broad range of products from a wide selection of sampling locations, i.e. grinding circuit, flotation cells, smelters and roasting kilns. etc., are evaluated analytically and microscopically, followed by periodic troubleshooting of problem points. The various crystalline phases of molybdenum that occur naturally and throughout various hydrometallurgical and pyrometal1urgica1 processes are identified and characterized by a variety of instrumental methods, including XRD-XRF, optical and electron microscopy, infrared spectrographic and differential thermal analysis.

Jan 1, 1989

By B Besa

Australia is a world leader in the design and operation of mines accessed by the decline method. Literature indicates that a number of metalliferous underground mines using the decline system are also increasing steadily. This increase has generated a great deal of interest in underground haulage systems in future due to the huge costs involved in the development of declines. Studies have also shown that the current method of accessing orebodies by the conventional decline method has proved to be expensive, especially for narrow vein deposits. This is because decline development costs are influenced by mine planning and design parameters as well as available haulage systems that the mining engineer can choose from. The current access method suggests that more waste material is being extracted because of the size of decline openings adopted (5.5 mW and 5.5 mH). Also the conventional 1:7 decline gradient and turning radius of 20 m makes decline lengths, as well as the duration of development, excessively longer. This paper discusses results of a mine design case study with monorail technology as a method of accessing and developing the decline. The proposed system is economical, has high advance rates and offers a significant reduction in capital expenditure. The study compares costs, development metres and the duration of decline development using the monorail system with the conventional (truck haulage) access method. Results obtained indicate that the monorail technology has the potential to reduce the decline length by over 62.6 per cent and decline costs by 63 per cent for the case study. The monorail system also reduces the total capital development costs by 22 per cent for the same case study. Due to rapid development resulting from increased advance rates and the shorter decline length coupled with smaller decline openings, the time to access the deposit reduced by 71.8 per cent. This case study demonstrates that the economics of narrow vein deposits can be improved significantly by using the monorail system of decline development.

Jan 1, 2008

By Kunzo Amano

Fully-bonded rock bolts have been widely used in construction sites and mines in recent years. The purpose of this investigation is to prove the effectiveness of rock bolt supports in unstable ground through in-situ tests and numerical analysis. 2. MEASUREMENTS AND PROCEDURES OF IN-SITU TESTS The fully-bonded rock bolts as the mine road¬way supports were tested in the Fukasawa mine which is located in the northeast district of Japan (Fig. 1). The Fukasawa mine (one of the Kuroko mines) started the production in 1973, and the present monthly production is slightly less than 20,000 tons. The ore deposit which exsists at about 400 m depth is divided into several small lenticular shaped bodies. They are distributed apporiximately 1500 m in NW-SE direction and 700 m in NE-SW direction. The hanging wall is basaltic lava and the footwall is tuff. The rocks surrounding the ore bodies are soft but relatively intact, so the trackless mining system has been adopted from the beginning of the development.

Jan 1, 1982

By M. Kuchta, S. Iverson

INTRODUCTION Underground blasting with perimeter damage control can reduce the requirement for loose rock scaling and reduce the risk of ground-fall injury. In previous work Johnson and associates (2010) from the National Institute for Occupational Safety and Health (NIOSH) developed a rock blasting perimeter damage and control design method included in the DRIFT software tool (NIOSH, 2018). This blast design software tool utilizes current methods of perimeter blasting with the added design element of a buffer row coordinated with the perimeter row to reduce perimeter damage. The coordination is in applying one of several blast damage models to calculate the damage extent for each production, buffer, and perimeter blast hole in the design. The calculated damage is then applied to the spacing between the blast holes. The basis of the buffer row design is the practical damage limit and is supported in this paper with computer models of signature blasts and full blast rounds using PFC2D (Itasca Consulting Group, Inc., 2018). Practical damage limit The practical damage radius (Rd) is not to be confused with burden. The practical damage extends toward the burden side and inward into the rock mass. For all practical purposes, the practical damage radius equals one half the burden (Hustrulid, 1999). The maximum extent of radial crack damage around a blast hole (Rcrack) and the maximum extent of radial crushing damage (Rcrush) are illustrated in Figure 1. A “practical” radius of damage (Rd) lies between the crushing limit and the cracking limit. The “practical” damage means that if the rock mass lying outside of the practical damage circle were removed, the rock remaining within the circle would easily break apart. Blast pressure and strain Work by Johnson (2010) and associates included signature blast tests in large concrete blocks. A signature blast test in concrete as exampled by strain data in Figure 2 was designed to monitor the dynamic strain at incremental distances from a single fully coupled blast hole, measure crushing and radial cracking extent by wire-sawing to expose the damage, and relate the damage to overbreak that is common in underground blasting. The initial slope or rising strain as the stress wave passes by the strain gages shows very high strain and strain rates close to the single blast hole. Since the derivative of strain versus time is strain rate, a line tangent to each of the strain curves at the maximum strain rate is drawn and the strain rates identified. The high strain rates were then plotted on a semi-log in Figure 3 to show a linear trend. The strain rate is greater than 100 s-1 within 20 cm from the blast hole for the signature blast experiment. The strain rate at the blast hole wall is projected to be 300 s-1. Figure 4 illustrates the damage in the NIOSH block exposed by a wire saw cut.

Jan 1, 2019

By H S. Smit, P Lotz

Gold mining companies who are signatories of the International Cyanide Management Code (visit www.cyanidecode.org for further information) are required to measure WAD cyanide on-line and destroy WAD cyanide in excess of specified limits. Weak acid dissociable (WAD) cyanide refers to those cyanide species liberated at a moderate pH of 4.5 such as HCN(aq) and CN-, and the majority of Zn, Cd, Cu, Ni and Ag complexes with low dissociation constants. It is potentially toxic to humans and animals. Recognising and in response to the gold mining industryÆs need for an on-line WAD cyanide measurement device, Mintek recently developed the WAD Cynoprobe. The WAD cyanide analysis technique is based on the use of a ligand exchange reagent that liberates WAD cyanide from metal-cyano complexes. The ligand exchange reaction is represented by the following chemical equation: Metal(CN)X + yLigand <-> Metal(Ligand)y + xCN- The plant filtrate is treated with a ligand exchange reagent, to liberate WAD cyanide, before measuring the free cyanide concentration via amperometry. The instrument proved to be accurate and reliable in the laboratory as well as under industrial gold plant conditions during various trials. It has gained wide-spread commercial acceptance both locally and internationally. However, itÆs recent exposure to complex ore types have called for refinements to the analysis technique. It has been further developed to improve its precision and accuracy and to remove reagent and thiosulfate interferences. Its compatibility with the most commonly used destruction oxidants has also recently been investigated. The WAD Cynoprobe may be configured to measure both free and WAD cyanide concentration. Its fast analysis time, approximately 5 min, makes it suitable for use in the control and optimisation of the cyanide destruction and recovery process. Proper oxidant dosage control and monitoring of effluent concentration levels will ultimately result in less cyanide and post-destruction salts discharged into the environment.

Jan 1, 2008

By C. H. Stockwell

THE Eastern Townships of Quebec have been the chief Canadian source of chromite. Production, however, has never been large and, except for a period of steady but small output between 1894 and 1909, mining has been active only in war time. The known deposits are too small or too low grade to compete with overseas sources in normal times. During the last war, shipping difficulties and a greatly increased demand in the United States resulted in high prices and a ready market for low-.grade material. The Quebec chromite industry flourished and played an important part in relieving the wartime shortage. Shipments in 1918 .amounted to 15,600 tons of crude ore and 6,400 tons of concentrates. At the outbreak of the present war it was natural to turn again to the Quebec field as a possible source of this strategic mineral. Some of the old properties have since been opened by private and government money and some new deposits have been found. Total mill capacity is 750 tons per day, as compared with 200 to 300 tons in 1918. During 1940, 1941, and 1942, about 85 chromite deposits were examined by the Geological Survey in a belt 100 miles long and up to 7 miles wide. A part of this belt, about 35 miles long, in which the geological features ?were not well known, was mapped on a scale of 1/2; mile to an inch.? Small ..areas around some of the deposits were mapped in detail, on scales of 20 to 200 feet to an inch. Some geophysical work was done but did not result in the discovery of ore. Prospecting by stripping and trenching was then undertaken and chromite ore was uncovered at several localities. The above work was done by the writer and by J. W. Ambrose, who has kindly contributed his results for incorporation in this paper. Y. O. Fortier actd as assistant in the geological work and, in 1942, was in charge of a sub-party engaged in geological mapping. The author is indebted to H. C. Cooke and J. W. Ambrose for reading the original manuscript and for offering many helpful suggestions for its improvement.

Jan 1, 1944

By Rod L. Norfleet

Cemented tungsten carbides are a sintered powder - metallurgical product essential to modern industry. Because of their moderate unit price and our dependence upon foreign sources for the component raw materials, a growing recycling industry has developed in the United States. There are two basic approaches taken to recycling cemented carbides; those which leave constituting materials unaltered or direct-reuse and those which separate the constitutes back to the raw components. To separate the carbides back to their raw components usually requires expensive chemical processing and is thus confined to the lower value materials such as grinding sludges. The direct-reuse methods recycle cemented carbides back to an intermediate level, a powder capable of being recompacted and resintered into a functional end-product. The three most commonly used methods of direct-reuse recycling, the coldstream process, bloaching and the Bernard or zinc process, are reviewed,

Jan 1, 1984

By D. B. Dreisinger, V. van der Pas

Zinc sulphate electrolyte used for zinc electrowinning must be purified for cObalt ions. The cobalt is removed in a cementation stage by the addition of zinc dust. Copper arid antimony are frequently used additives which enhance the cementation of cobalt. This paper aims at a better understanding of how copper and antimony promote the removal of cobalt. Initial experiments were done in a batch cementation reactor. Copper and antimony precipitated in the early stages of cementation indicating that a preferential substrate was formed. The individual role of copper and antimony was further investigated in an electrochemical cell. On a microsca1e, the , growth process, morphology and the composition of the precipitates were examined with SEM and XRD. It was found that cobalt could not be deposited in its pure form but as a cobalt-zinc alloy with zinc as the prime constituent. Copper increases the cathodic surface area of zinc dust by precipitating as numerous dendrites. Antimony acts as a cathodic surface onto which a cobalt-zinc alloy with an increased cobalt content is deposited.

Jan 1, 1997

By S. L. Crouch

The general term "coal mine burp" refers to the sudden and violent failure of in situ coal. Coal bumps occur in most countries where coal is worked by underground methods, and are related to both geological factors and mining tech¬niques. Although no single set of circumstances is solely responsible for causing coal bumps, it is generally agreed that at least one of five conditions is necessary in order for then to occur: 1,2,3,4 (1) the depth of cover exceeds 500-ft., (2) the overlying strata are relatively strong and stiff, (3) the coal is structurally strong, (4) the floor does not heave readily, and (5) the mining method produces high stresses over a large area of the seam. Except for the last, these conditions relate to geological factors that are impractical to control. Consequently, methods of controlling or minimizing the occurrence of coal bumps usually are concerned with mining techniques. Coal bumps have been successfully controlled by both longwall and 'yielding pillar' mining systems 4. The longwall system is particularly effective in controlling bumps, but does not eliminate them completely, especially in strong coal seams being worked at comparatively great depths, say in excess of 2000-ft.5. Under such conditions, auger drilling or water infusion often prove helpful 5,6. These techniques are used to destress or otherwise alter the mechanical properties of the coal in an attempt to shift the peak "abutment pressure" away from the immediate vicinity of the face. Under 'normal' conditions, however, it can be expected that the coal near the face will be crushed under the prevailing high stress conditions and that the peak abutment pressure will automatically be shifted into the solid coal. In this context, coal mine bumps can be characterized as unstable releases of energy associated with successive yielding of portions of the coal seam with progressive mining. Longwall and yielding pillar mining systems are effective in controlling coal bumps because they ordinarily do not allow the zone of peak abutment pressure to become 'surrounded' by destressed coal.

Jan 1, 1973

By M. C. Esposito

Seven pyrite samples purified from coal sources and one ore source pyrite (for comparative purposes) have been characterized by chemical and mineralogical analyses, inherent floatability, apparent specific gravity, surface area, semiconductor type, optical anisotropy, dissolution and oxidation rate. The data clearly demonstrate that pyrites from various coal sources differ not only from ore pyrite but from each other as well.

Jan 1, 1987

Research on alternative lixiviants for gold is ongoing; however few people have looked at complete flowsheets. Alternative lixiviants are typically applied at concentrations 50 to 100 times those required for cyanidation. The higher concentrations make lixiviant recycle a vital part of the process. In addition it is no longer feasible to rely on lixiviant destruction to produce lixiviant free clean water for recycle to the comminution circuit. Ausenco has worked with Barrick to integrate reverse osmosis into the thiosulfate RIL processing flowsheet that forms the basis of the plant upgrade currently being constructed at Barrick’s Goldstrike autoclave facility.

Jan 1, 2014

By C. M. McCulloch

The natural vertical fracture system in bituminous coalbeds is called cleat. Cleat orientation commonly controls the direction of mining with major development paralleling the face cleat. Previous researchers have categorized the origin of cleat as endogenetic, relating the origin of cleat to compaction and coalification, and exogenetic, relating the origin of cleat to tectonic forces. In the coalbeds studied for this report it was found that tectonic forces were the controlling factor of cleat formation. Face cleats were formed as extension fractures during structural deformation, and butt cleats, as release fractures during erosion and uplift. The face cleat of the Pittsburgh coalbed rotates from N 80° W in north-western West Virginia to N 57° W in southwestern Pennsylvania; the face cleat maintains a perpendicular orientation to the shifting axial trend of local structures. Cleat orientation in other bituminous coalbeds in Virginia, Utah, Oklahoma, and Central Pennsylvania showed a similar relationship to local structure. Directional permeability of coal is directly related to cleat. Holes drilled perpendicular to the face cleat yield from 2.5 to 10 times the amount of gas released as compared with holes drilled perpendicular to the butt cleat.

Jan 1, 1974

By Oscar M. Wicken

Iran was a lodestone for the attention of industrial countries and companies wanting to share in its industrial development during the last two decades and particularly after 1972. Prior to 1972, most industrial projects in the private sector were on a relatively small scale and could be carried out in reasonably orderly fashion in spite of limitations of a primitive infrastructure and meager indigenous supply of managerial and technical skills. People with these skills could be found but not in abundance, and many of them were total¬ly engaged in various government ministries and projects. Thus it was that nearly all private sector industries, even long-established ones, required one or more foreigners to give guidance. A barite mine and processing facility was brought on line by Dresser Industries in 1960 without serious operational problems. A few expatriate experts were able to select and train a stable group of staff and production people and establish sound management principles for continued growth. The huge and increasing oil revenues after 1973 gave superficial credence to ideas of the Iranian government that the country could easily finance the jump into the industrial 20th century. All necessary skills and equipment not available in Iran would be imported, and intensive training programs would prepare Iranians for operating the new facilities. Large projects for steel production, copper production, oil refining, petrochemical production, cement production, and others were started or expanded. These are basic industries that consume substantial quantities of refractories. Studies made in 1972 indicated that the existing industries plus expanded and new projects scheduled to come into production would lead to a market sufficiently large to justify a refractory producing plant. Concurrent studies of raw material sources showed that clay, the largest tonnage item, would be locally available; the other raw materials would be imported. Iran Refractories Company (IREFCO) was founded in 1973 and started production in 1976.

Jan 1, 1981

By W. G. Mine

In September, 1953, seismographs were installed in the Crowsnest Pass area of Alberta and British Columbia to study the 'bumps' which were occurring in the coal mines at Fernie and Coleman, and to investigate the possibility that they might be related to minor earthquakes in the area. During the initial phase of the work, three seismographs were operated at widely separated stations, two in the region of coal mines at Coal Creek and Coleman and one at Turner Valley, distant from any mining operations. Some earthquakes were recorded at all three stations and, in addition, many small disturbances were recorded by the seismographs located near the mines, especially -by the instrument at Goal Greek. A closer network of radio-linked seismographs was then in-stalled near Fernie, the records of which definitely established that the small disturbances were occurring in the mine workings, both active and inactive. No relationship was established between bumps and local earth-quakes in the area. This work was initiated as a -co-operative effort with the Fuels Di-vision of the Mines Branch, Department of Mines and Technical Surveys, which has undertaken a study into the phenomena of ground stress relief in underground mine workings.

Jan 1, 1958

By Aarne Visnapuu

Dissociation and vaporization of gold chlorides and bromides were investigated by a static pressure method and by Knudsen cell mass spectrometer. Two distinct chlorine and three distinct bromine pressure-temperature relation-ships were observed for the respective salts. The chlorine pressures are attributed to the dissociation of Au2Cl6 and Au2C12, and the bromine pressures to the dissociation of Au2Br6, Au2Br4, and AU2Br2, These pressure-temperature relationships were used to calculate dissociation pressure equations, enthalpies, and entropies of dissociation for each salt. Mass analysis of the vapors over AU2Cls between 83° and 152° C confirmed the existence of gaseous AU2C16. Ionization potential of this molecule is 11.0±0.1 ev and enthalpy of vaporization is 28.l±2.0 kcal/mole. Mass analysis of the vapors over the gold bromides detected only Bra in the temperature range 15°-150° C.

Jan 1, 1971