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By D. C. Shumway

Mining in Cushenbury Canyon began in 1860 when John Cushenbury left his Holcomb Valley gold prospect and discovered a silver vein in the limestones of the present day Cushenbury Mine. Cushenbury Canyon is located along the western boundary of the Mojave Desert in Southern California and on the north slope of the San Bernardino Mountains. The Holcomb Valley, immediately southeast of Cushenbury Canyon was the most prolific gold producer in Southern California. Gold mining dominated the area until World War II when interest in the Lucerne Valley Limestone deposits was peaked by the construction of the Kaiser Steel Plant in Fontana, California. The steel making process required large amounts of metallurgical grade limestone and Kaiser geologists located claims in most of the limestone deposits. Kaiser Cement constructed a cement plant and limestone mine in Cushenbury Canyon in 1956. Kaiser Cement modernized the facility in 1982 to a four stage preheater-precalciner system. Mitsubishi Cement Corp. purchased the property in April, 1988, and through further system modifications have created the most efficient and modern cement plant in the United States.

Jan 1, 1992

By H. Ban

Triboelectrostatic separation tests were performed on a bench-scale, analytical separator using coal, coal combustion fly ash and mineral samples. Fractions of separated samples were collected and analyzed to characterize the performance of the separation. Particle size distribution and morphological characteristics for these samples were analyzed by optical microscopy, and scanning electron microscopy (SEM). It was found that smaller size fractions of a fly ash sample had better separation than the full sized parent fly ash sample, indicating interference or particle-particle interactions during the separation process. It was also observed experimentally that small kaolin particles attach to the surface of large silica spheres, causing charge neutralization and subsequent low separation efficiency.

Jan 1, 1997

By C. Patrick

Annually, the U.S. coal industry generates and must dispose of over 200 million tons of coal refuse, expanding a major environmental problem. The purpose of this research was to investigate the feasibility of utilizing coal refuse as an aggregate in concrete. This paper reports on findings of the initial phases of a project to determine the viability of replacing traditional concrete blocks, used for under-ground ventilation stoppings, with refuse blocks. Refuse was collected from three active preparation plants in Eastern Kentucky and used as the primary aggregate for concrete cylinders. The refuse fell within ranges normally observed for size consist, moisture, and ash. Stress-strain analyses yielded compressive strengths of the refuse cylinders from 59% to 77% of the control group strengths. Low variability within groups while significant differences between groups were found. On-going determination of effects of weathering on the refuse cylinders is being conducted and will be completed in early 1995. The research results are favorable for the future direction of the overall project objectives.

Jan 1, 1995

By James L. Vanderbeek

INTRODUCTION Semiautogenous (SAG) grinding technology first started to be seriously considered for comminution circuit design in the mid-1960s as an alternative to the three-stage crushing circuits and rod-ball mill circuits in vogue at the time. Since this time the application of SAG mill technology has become widespread with over 250 mills larger than 20 feet in diameter sold (Jones 2002). The vast majority of concentrators constructed over the past quarter century have utilized SAG mill technology, and a SAG-based comminution-based circuit is generally the first choice considered for new projects. Since the first generation of SAG milling technology was applied, many lessons have been learned, and new supporting technologies have emerged which make the design, installation and operation of SAG &-based grinding circuits more robust and efficient. This paper will discuss the current status of SAG and associated technologies. MILL MECHANICAL DESIGN ISSUES Table 1 summarizes the manner in which SAG design capability has evolved over the past 30 years. Increases in mill diameter came slowly as the industry took a wait-and-see attitude as the technology of larger mill diameters was advanced. The first 32-foot- diameter mill was ordered in 1969. It was not until 1979 that the technology jumped to the 34-foot-diameter mill and 1988 for the first 36-foot-diameter mill. The first 38- and 40-foot-diameter mills were both ordered in 1996. At least five 38-foot-diameter and two 40-foot-diameter mills will be in operation by mid-2004. Advances in structural analytical techniques and increased computational power, coupled with field measurements of large operating mills, have increased the confidence of the mill suppliers in their designs and the operating companies in the viability of larger diameter mills. No failures of a 36-foot-diameter mill or larger have been

Jan 1, 2004

By K. Osseo-Asare

Metal recovery from solution, i.e., the formation of a solid product from a purified aqueous electrolyte solution, is increasingly becoming more than a simple matter of bulk precipitation. In many cases, operating conditions must be carefully manipulated to control the rate and yield of the precipitation/crystallization process, as well as the purity and physical characteristics of the solid product. A review is presented of the underlying physico-chemical processes, with emphasis on solution and interfacial chemistry. Opportunities for the integration of hydrometallurgical metal recovery and materials processing operations are discussed.

Jan 1, 1989

By D. A. Curry

This paper describes a study of the influence of non-normal loading on the indentation behaviour of single bit-teeth. Indentation experiments have been performed on a soft, isotropic limestone using both sharp and blunt 60 degree wedge-shaped indentors constrained to penetrate the rock at 0, 15, and 30 degrees to the rock surface normal. For both sharp and blunt indentors, the normal force required to produce a given normal penetration decreased as the offset of the penetration direction from the rock surface normal increased. The crater volume produced by a given normal penetration with the sharp indentor was largely independent of the penetration direction. In contrast, the crater volume produced by the blunt indentor was increased markedly by non-normal penetration at small penetration depths, because the threshold penetration for chip formation decreased with non-normal loading. The crater volume produced by a given normal force was significantly increased by non-normal penetration, and this increase was particularly pronounced with the blunt indentor. The energy required to form unit crater volume was correspondingly reduced by non-normal indentor penetration. A simple method for predicting the reduction in the rock's resistance to chip formation is outlined, which is supported by the experimental data. The implications of these findings for rotary drilling are considered.

Jan 1, 1986

?Energy-related systems: wind, solar, electric are typically materials intensive. As new technologies are widely deployed, significant quantities of the elements required to manufacture them will be needed. However, many of these unfamiliar elements are not presently mined, refined or traded in large quantities, and, as a result, their availability might be constrained by many complex factors. A shortage of these ?energy-critical elements? could significantly inhibit the adoption of otherwise game-changing energy technologies. This, in turn, would limit the competitiveness of the U.S. industries and the domestic scientific enterprises and, eventually, diminish the quality of life in the US.? -Energy Critical Elements: Securing Materials for Emerging Technologies?- American Physical Society, 2011.

Jan 1, 2012

By Ian J. Duckworth, Pierre F. X. Mousset-Jones

A detailed heat flow project was conducted at the Homestake Mine in Lead, South Dakota. This paper presents results obtained from a stope sensor system installed in the mine. The design and installation of the system are described in detail in a previous paper (Duckworth et al., 1992). The major project objective was the construction of a comprehensive heat flow data base to allow the design and testing of a new climate prediction model applicable to irregularly shaped underground excavations in non-homogeneous surrounding rock types. A mechanized cut-and-fill (MCF) stope contributes a substantial overall heat source in a mine, and the effect of this source on the working environment must be considered when planning future ventilation and refrigeration requirements. The relative magnitudes of the component heat sources encountered with the MCF mining technique are presented, having been computed from actual stope beat flow data collected over a one-month study period. The percentage of the overall heat load that is removed from the scope for each component heat sink is evaluated, and the overall energy balance is analyzed. A one-month study length was considered a sufficient period of time to allow stope monitoring for a complete range of mining conditions.

Jan 1, 1993

By Ken J. Fair, Vicken H. Aprahamian, Gus Van Weert

A consortium of Canadian mining companies, engineering firms and the Ontario Government has supported Hydrochem's efforts in the last year to design and cost a 10 tpd NITROX PROCESS® demonstration plant. A 5 kg per hour mini plant was installed at the Hydrochem facilities in Brampton, Ontario. This paper describes design testwork on Dickenson Mines arsenopyrite concentrate and application of the findings to the demonstration plant flowsheet. The control of nitrates in discard solids and stack gases is discussed in detail. Nitric acid oxidation of arsenopyrite results in significant ele- mental sulphur formation, enough for it to be detrimental to sub- sequent cyanidation. Flowsheet variations aimed at dealing with this process characteristic are discussed and testwork results on one option (hot liming) are described.

Jan 1, 1988

By M. Shen, T. D. Wheelock

The important characteristics of a gas-promoted oil agglomeration process for cleaning coal particles in an aqueous suspension were determined. Numerous agglomeration tests with scale-model mixing systems of different sizes with controlled amounts of air were conducted. The tests were conducted on high- and mediumvolatile bituminous coals using either i-octane or hexadecane as an oily agglomerant. The performance of the system was characterized by employing the minimum time required to produce spherical agglomerates (tE), the mean agglomerate diameter (dp), coal recovery and product ash content. Among these parameters, tE was found the most useful. It is a key parameter in the scale-up of an agglomeration mixing system. This parameter was found to depend on a number of system and operating variables. By controlling these variables and conducting further agglomeration tests with mixing systems that varied in size, an appropriate size scale-up rule was established.

Jan 1, 2002

By E. Topuz

The objective of mine ventilation is to provide suitable environmental conditions in working places at an economical cost. Maintaining a stable air flow, i.e., a flow quantity and the direction, is of vital importance. In practice, however, air flow quantity in underground openings varies continuously. Therefore, the stability of air flow means relative stability that is the flow in a branch moves along a certain direction while the variation in air quantity stays within acceptable limits. Many underground mines have low stability of mine air flow because almost all types of ventilation systems involve complex diagonal branches through which air flow quantity and, more importantly, the air flow direction may change easily. This creates extremely hazardous, unstable conditions, especially during emergencies. This paper deals with the directional stability of air flows in underground mines in single or multiple fan systems. It examines the rules of directional changes of air flows in diagonal branches, and proposes a method of calculating the stability conditions. It provides a criterion that may be used in designing new mine ventilation systems to ensure a desired degree of stability or in improving the existing systems.

Jan 1, 1986

By E. Peuraniemi, S. Strömberg, S. Jyrkönen, A. Jokilaakso

Several sulfidic raw materials have been flash oxidized in a laboratory furnace in oxidizing atmospheres at temperatures of 500-1100°C. Chemical analysis was used to determine the oxidation degree, and optical and scanning electron microscopy combined with energy dispersive analysis were used for characterization of the reacted particles. The studied materials included both Cu and Ni concentrates and mattes. In fixed conditions, the temperature dependence of sulphur removal rate was found to be equal for all the copper concentrates studied. Nickel concentrates and Cu- and Ni-mattes were, however, found to be less reactive. Most of the particles were found to spheroidize and to fragment to small droplets under intensively oxidizing conditions. The product particles were frequently found to be iron oxide cenospheres or dense spheres with a composition of mixed Cu-(Ni-)Fe oxides.

Jan 1, 1998

By Young C. Kim

Because of the high cost of obtaining coal washability samples, it has been difficult to obtain washability variograms in the past. As a result. It could not be determined if geostatistical prediction tool known as kriging can be applied to predict in-situ coal washability characteristics.

Jan 1, 1988

By D. W. Park

Longwall mining is considered to be the most efficient underground coal-mining technique in terms of coal extraction ratio, production rate and safety standards. Due to the increasing demand for higher coal production, the recent trends have been to use wider panels and to maintain a faster face movement. For these reasons, longwall mining can result in unexpected roof falls, excessive face convergence, shield damage and numerous maintenance problems, which can cause sudden interruption of coal production as well as serious safety hazards. Recent work at the University of Alabama has been dedicated to the development of a longwall strata-control and maintenance system that can integrate real-time shield monitoring, data presentation, analysis, forecasting and warning modules. This integrated system has been called LoSCoMS. The main objective of the system is to provide minute-by-minute stability and maintenance information such as the following: high-load, low-setting pressure, leg leakage, etc.; the analysis of various mining and geological parameters to determine past and present roof characteristics; and the forecasting of loading patterns for the following mining cycles, the next periodic weighting, etc. By using this unique system, mine operations personnel are well aware of past, present and future roof-loading patterns at the longwall face and, thus, proper safety measures can be taken ahead of time.

Jan 1, 1999

By Terry Ramsey, N. Finkelstein

INTRODUCTION The Cyprus Pima Mine is located 20 miles southwest of Tucson, Arizona and about fifty miles north of the Mexican border. The Cyprus Pima orebody was discovered in 1950. Feasibility studies were begun in 1952 employing underground development. Economic studies showed open pit operations to be feasible and stripping of overburden was begun in late 1955. The first ore was reached in October of 1956. A 3,000 tpd (tons per day) concentrator was completed and began operation January 1, 1957. From these beginnings, Cyprus Pima Mining Company has expanded into a plant processing approximately 54,000 to 56,000 tpd. Four expansions were required to bring the plant up to its present capacity. Each expansion incorporated what had been learned in previous operation of the plant. The first expansion was completed in August of 1963 doubling design capacity to 6,000 tpd. In reality, because of softer than average ore, capacity was closer to 8,000 tpd. Expansion I1 was completed in July 1966 bringing design capacity to 18,000 tpd. Expansion I11 followed immediately. Completed in the summer of 1967, this expansion increased the design capacity to 30,000 tpd. The actual dry tonnage of this portion of the operation in 1974 averaged 38,400 tpd, again, primarily as the result of treating softer than average ore. The latest expansion (Expansion IV), completed in December 1971 and utilizing a semi-autogenous grinding circuit, is located in a separate building (Mill 11) approximately 200 yards southeast of the original mill (Mill I). Expansion IV was designed for 14,000 tpd, but after circuit modifications, averaged 16,500 tpd in 1974.

Jan 1, 1976

By D. Belin

Treatment of dilute acid mine drainage and similarly polluted water is becoming increasingly complicated. The pressure of economics is now exacerbated by legislation which classifies as hazardous waste the precipitates from conventional lime treatment plants. Since acid mine drainage invariably contains sulfate ions, one solution is to use anaerobic Desulfovibrio type bacteria to reduce the sulfate into sulphide ions which could then be used to precipitate out the heavy metals. Results are presented for both batch flask cultures and continuous column tests. Compost was used as the bacterial substrate in the column test which simulated heaps. Conditions changed from aerobic to the required anaerobic deep in the columns. Polluted feeds (up to 700 mg/l zinc) gave effluent containing well below I mg/l zinc over extended periods. Direct transfer of sulphide ions through a stream from an isolated container of bacteria is a good alternative.

Jan 1, 1991

By S. J. Jung

In recent years, the use of shotcrete, a practice familiar to the mining industry, has been on the increase. This research has come about because of the recognized need for a safe, easy way to line small diameter bored raises (less then 150cm). Up until now, an operator has gone down the raise and applied the concrete lining manually, using the shotcrete method of application. With an automated system, there would be no man positioned in the bore itself. All control functions could be commanded from a central location. Shotcreting in the confined space of a small diameter bored raise presents problems that can be overcome by the application of shotcrete using a remote system.

Jan 1, 1994

By R. Pakalnis

This paper compiles the research work conducted by the Spokane Research Laboratories in the investigation of shotcrete within underground mines to the applied design of ground support for the operator. It looks at the design methodology in terms of employing shotcrete as a fabric to confine the rock mass between the individual bolts as dictated by empirical charts. The failure mechanism in terms of its implication on design is presented with respect to the design of 6m wide tunnels for temporary and permanent openings within weak rock masses. This is reflected by field observations that relate laboratory tests to the onset of failure for the operator in order to provide guidelines for design and monitoring. This is coupled with guidelines for safe entry in terms of cure times as well as the understanding of the behaviour of the shotcrete in terms of fibres and the overall strengths that result with respect to the design of the mine openings. This enables the operator to better understand his workplace thereby provide a safer work environment.

Jan 1, 2010

By D. A. Dahlstrom

Crushing and grinding consumes about 1.3% of the nations produced electrical energy.( Furthermore, only one in the first ten highest consumers comes from an industry other than mineral processing. This is the grinding of wheat to make flour. The other nine are all in the metallurgical industry and are easily the largest consumers.

Jan 1, 1988

By Malcolm J. McPherson

The highest recorded loss of life due to the collapse of mine workings occurred at the Coalbrook North Colliery in South Africa on 21st January, 1960. A total of 437 persons died when bord and pillar workings collapsed over an area exceeding one square mile The whole mass of strata from the 470 ft (143 m) deep seam to the surface being affected. Although the Coalbrook disaster was unique in its scale the problem of pillar failure, particularly in old workings, is by no means uncommon nor confined to recent history. Indeed, the collapse of large pillared areas was a regular feature of mining in the north of England in the early part of the nineteenth century. When a section of underground workings collapses, the air under the descending roof is compressed very rapidly. The pressurised air is dissipated subsequently at a rate governed by the leakage paths available. If the collapsing area is small, such as the case of a local fall of roof, then there will be little danger of an air blast. If, however, an extensive area collapses suddenly then the rapid compression of air will give rise to a pressure shock wave that will be transmitted through the remaining workings as an air blast. If worked out areas are connected to current workings via unstopped airways an air blast caused by pillar failure in the old workings will be extremely hazardous to personne1 and may cause extensive damage to equipment, perhaps even reaching to surface plant. At Coalbrook, the air blast blew out a wall in the surface fan house. Areas that are considered liable to sudden collapse may be isolated from current workings by the construction of stoppings. However, such stoppings must be capable of withstanding the 'hammerblow' effect of peak air pressures developed during the collapse. The work described in this paper was undertaken in order to determine the maximum air pressures that stoppings must withstand during a major pillar failure. The result has been a method of establishing peak air pressures that is quick and simple to use. Readers who are concerned primarily with the practical application of the method are advised to turn directly to the last main section of the paper.

Jan 1, 1980