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By T. J. Close

The Silver Peak Range North Wilderness Study Area was deleted from the list of US Bureau of Land Management lands in Nevada to be considered for wilderness designation. This decision was based in part on the US Bureau of Mines mineral investigation of the wilderness study area which found indications of precious- and base-metal-resources. The study area lies between the Silver Peak and Dyer mining districts of west-central Nevada. The Bureau examined 13 mineralized sites, most of which were previously unknown and unclaimed. These sites are along two zones that cross the study area from the Silver Peak mining district to the Dyer mining district. The two zones transect a volcanic center of fractured and altered rocks that contain epithermal veins of quartz and calcite with gold, silver, and base-metals. At the surface, these veins are poorly defined. Exposures of the more than forty veins examined were as thick as 17 ft and as long as 1000 ft. Samples from these veins contained significant amounts of precious- and base-metals-as much as 0.02 oz per ton gold, 2.1 oz per ton silver, 1.0% copper, 2.6% lead and 0.4 % zinc. No vein large enough or of high enough grade to be classified a mineral resource was disclosed by the sample data. However, most do have indications of precious- and base-metal-resources at depth. Subsurface exploration of some of the veins is recommended.

Jan 1, 1995

By A. Anttila, S. Persson

The grinding mill is probably the most common unit operation in mineral processing. There is, however, much that can still be done in the way of optimization. This paper discusses the results of a well-controlled pilot-scale test in which the performance of a tumbling mill equipped with a conventional rubber lining and lifter bars was compared to a tumbling mill equipped with a smooth rubber lining of the magnetic type. The results indicate that the smooth lining is about 14% more effective with respect to energy consumption. By optimizing the size of the grinding media, the mill capacity can be increased in many grinding operations. In this specific case, grinding with a -20-mm ball charge required 26% less grinding energy than grinding with a -30-mm ball charge. The energy savings can, in most cases, be converted to an increase in capacity. The wear of the grinding media is difficult to assess over a short test period, but the results suggest that the wear is somewhat less when a smooth lining is used. A determination of the wear of the linings was not included in this study, but it is generally accepted that the use of magnetic linings reduces the lining costs dramatically. Computer simulation using the Millsoft code developed by Mishra et al. confirms and explains the increased effectiveness of comminution when a smooth lining and smaller grinding balls are used.

Jan 1, 2000

By O. M. S. Rodrigues

The kaolinite presence in bauxite ores increases caustic soda consumption during the digestion step in the Bayer process. Depending on the kaolinite content, the utilization of the ore in the Bayer process may be unfeasible. This investigation addressed the variables involved in kaolinite flotation through microflotation experiments in a Hallimond tube and zeta potential determinations. The tests were performed with a pure sample of kaolinite (96%). Five collectors were tested: Duomeen T (DT), cetyl pyridinium chloride (CPC), cetyltrimethylammonium bromide (CTAB), dodecylamine (DDA) and Flotigam EDA (EDA). Four of them performed well as collectors, at a concentration of 1x10-4 mol/L: CPC, CTAB, DDA and EDA. The kaolinite isoelectric point (IEP) was determined at pH = 4.6. The IEP was shifted towards higher pH values in the collectors? presence, indicating specific adsorption onto the mineral surface.

Jan 1, 2011

By LeRoy D. Loy

To facilitate effective land reclamation and comply with existing state regulations, surface coal mine operators have in recent years adopted new mining techniques and equipment utilization for steep-terrain mining. In contrast to direct spoiling with draglines or shovels, these new mining methods require some form of overburden haulage from the excavation face to the point of reclamation. Although current mining methodologies such as Block-Cut, Modified Block-Cut, and Mountaintop Removal with Head-of-Hollow Fills have proven effective in achieving reclamation objectives, some production problems have resulted from pit congestion and excessive equipment delays. Recognizing this problem, the United States Bureau of Mines contracted Skelly and Loy to design innovative mining concepts utilizing conveyor haulage in place of trucks, loaders, dozers and scrapers. Project objectives required definition of current bottlenecks and development of mining systems to: alleviate bottlenecks; increase production potential; facilitate additional resource recovery; enhance reclamation; reduce safety hazards; and improve mining economics. Successful completion of project goals within a four year program involves systematic implementation of the following separate work phases: 1) Develop Concepts for Block-Cut Mining Systems; 2) Determine the Economic and Technical Feasibility of Concepts; 3) Prepare Detailed Design and Specifications of Concepts; and 4) Prepare Specifications for Conducting Field Tests of the Most Promising Concepts. At the time of this publication, Phases 1 and 2 are finalized with Phase 3 efforts well underway.

Jan 1, 1977

By J. M. Duke

Significant resources grading less than about 1 % nickel occur in magmatic sulfide deposits of two types. In the Mt. Keith type, extensive zones of disseminated nickel sulfides occur in the central portions of certain synvolcanic dunite-peridotite sills of komatiitic affinity in Archean greens tone terranes. In the Duluth type, disseminated copper-nickel sulfides occur near the basal margins of intrusions of flood basaltic affinity associated with intracontinental rifting. In both types, the parent magmas became saturated with sulfide following the onset of fractional crystallization and the separation of the immiscible sulfide melt was accompanied by the precipitation of a relatively large proportion of silicates. The abundant silicate crystals impeded the coalescense and accumulation of the sulfide droplets with the resulting formation of large but low grade deposits.

Jan 1, 1982

By Janet Briggs, Michael Vertes, J. F. Shirley, R. E. Cutherbertson, Alexander Sutulov, J. D. Vincent

Molybdenum, the 42nd element in the periodic table, is now a very important metal in industry of the United States and the world. Chemically, it is a very complex element having valences of 0, +2, +3, +4, +5, and +6, and can exist as a mixture of these valences in many compounds. It was discovered and separated by Scheele in 1778. At the present time, molybdenite is the only important commercial source of molybdenum. Molybdenum occurs in some 12-14 different recognized minerals, but practically all of the production is molybde¬nite. As late as 1915, it was considered a rare metal. Germany made extensive use of it during the first world war, which resulted in the establishment of a molybdenum industry in the United States at Cli¬max, Colo., in 1917. It was not until the mid-1920s when its introduc¬tion to steelmaking developed for the US automobile industry that the metal gained important acceptance. The first notable example was the Willis-Saint Claire automobile introduced in 1921 by C. Harold Willis, formerly chief metallurgist for Ford Motor Co. Molybdenum has also been produced in the US from wulfenite (PbMoO4) from Arizona and Nevada, and a small amount of molyb¬denite was produced up to the early 1970s as a coproduct from two molybdenum-bismuth mines in Quebec, Canada. The first byproduct production of molybdenite from a copper con¬centrating operation was achieved at Cananea, Sonora, Mexico, and was followed shortly by the Utah division of Kennecott Copper Co. in the mid-1930s. This section of the Handbook is divided into four main subdivi¬sions: Chapter 2, which deals with the concentration and recovery of molybdenite from primary molybdenum ores; Chapter 3, which deals with the recovery and separation of molybdenite as a byproduct from copper production; Chapter 4, which deals with the conversion practices employed for converting molybdenite to the oxide or metal form; and Chapter 5, which covers the uses of molybdenum in steel¬making, chemicals, pigments, lubrication, catalysts, etc. The Free World molybdenum production, for the past eight years from all sources, is shown in Table 1. 2. Primary Molybdenite Ores General In 1978 there were four primary molybdenite recovery plants in operation: Climax, Endako, Quests, and Henderson. For over half a century the Climax mine has been by far the largest and most important source of molybdenum in the world. During the early 1970s a few primary molybdenite mines in British Columbia were shut down due to adverse economical conditions. The largest of these was British Columbia Molybdenum, formerly owned by Kennecott Copper Corp. The Climax plant and operations are described in detail and the opera¬tions of the other plants are summarized following a general discussion of recovering molybdenite from primarily molybdenite ores. Recovery of molybdenite as a byproduct will be taken up in Chapter 3. Geology Molybdenite production from mines in which it is the sole or preponderantly important mineral is confined to ore bodies in granites or igneous-type rocks or pegmatite dikes. More than one state of mineralization is sometimes present as at Climax, York, Hardy, and Questa. In the case of Climax, important byproduct recovery of tung¬sten, tin, pyrite, and monazite is also attained. The molybdenite is commonly disseminated very finely in the siliceous granitic gangue but does occur in stringers and blebs in silicified mineral solution channels. Treatment of Primary Molybdenum Ores All primary molybdenum ores are low in molybdenite, the content ranging from less than 0.20% MoS2 to a high of about 0.40% MoS2. A high resistance to grinding is a common ore characteristic combined with a habit of very fine dissemination of molybdenite in the siliceous portion of the ore. Because of low MoS2 content, high hardness, and fine dissemination of molybdenite in primary ores, it has been necessary to develop a special mill treatment which yields an accepta¬ble recovery at a reasonably low milling cost. Essentially, the treatment consists of intensive rougher flotation at a relatively coarse grind with production of a low grade rougher froth containing a substantial percentage of the molybdenite in the form of low grade middling particles, followed with multistage regrinding to liberate the locked molybdenite and several stages of flotation cleaning to produce a finished grade concentrate. Rougher flotation on pulp ground to ap¬proximately 35 mesh is common practice, even if complete liberation of molybdenite from siliceous gangue may require regrinding to 200 mesh. The flotation of middling particles with a very low content of molybdenite in the rougher flotation step has been made possible by the application of a combination of reagents first developed at the Climax mill and used with some modification in the treatment of all primary molybdenum ores. The reagents commonly employed and their functions in flotation are:

Jan 1, 1985

By J. F. T. Agapito

The load carrying capacity of oil shale pillars excavated by conventional blasting, presplit blasting, and mechanical mining is evaluated. The study was based on a comparison of in-situ vertical stresses and fractures obtained from overcoring horizontal holes in the Colony Mine, Piceance Creek Basin, Colorado. Results indicate that conventional blasting causes a zone of damage approximately 3 m (10 ft) thick with low stress distributions. Presplit blasting reduces damage significantly, and increases the load carrying capacity in the 3 m (10 ft) thick zone by 5.93 MPa (860 psi). Mechanical mining causes little or no rock damage, and an increase of 9.83 MPa (1425 psi) in strength in the same 3 m (10 ft) thick zone. An example of pillar design is given showing that the use of presplit blasting and mechanical mining techniques can increase the extraction ratio by at least three and five percent, respectively, as compared to conventional blasting. It is speculated that comparable increases in extraction should also occur due to increases in span dimensions.

Jan 1, 1984

By George R. Stephans

The Climax Mine is a panel caving mine which has had a long history of incidents of severe ground control problems. The nature of the ore deposit, which lends itself to the caving meth- of of production, a1 so makes the mine susceptible to large scale stress concentrations which are referred to as ?weight? problems. These problems have increased with increased production and also with expansion to deeper mine levels. During the past twenty years the mine has had to learn to cope with these problems in order to remain in operation. A number of practical guidelines and useful techniques have emerged from these efforts. The purpose of this paper is to share a few of these ideas. The Mine is located near Leadville, Colorado on the continental divide at an elevation of 3450 meters. The ore body is a large low grade deposit of disseminated molybdenum sulfide. It measures about 1200 to 1600 meters in diameter and it had an original depth of about 660 mters. The Mine has been in operation since 1918 and the massive nature of the ore body together with its blocky, highly fractured condition made it natural for the caving system to have been adopted for production at an early date. Open pit mining was started in 1973. The total plant has a capacity of more than 50 thousand tons per day, and more than 400 million tons of reserves at an average grade of 0.310 percent molybdenum disulfide remain to be produced.

Jan 1, 1982

By T. E. Wilson

An exploration bore core sample of coal ought to be reduced from its original cylindrical form to a particle size distribution and particle composition that reflects, insofar as possible, a process stream of? raw? coal it represents. Often, coal cores are reduced with a laboratory crushing machine, the product of which may not match the raw coal size distribution or size fraction composition. This study employs drop-shatter impact breakage followed by dry batch tumbling in a steel cylinder rotated about its transverse axis in order to characterize the core material in terms of first-order and zeroth-order breakage rate constants. These are indices of the propensity of the coal to degrade during excavation and handling. Initial drop-shatter and dry tumbling tests were done with synthetic cores, to reduce material variables and . conserve difficult-to-obtain coal cores.? Cores of three different coal beds --Illinois No.6, Upper Freeport, and Pocahontas No. 5 -were subjected to drop-shatter and dry batch tumbling tests to determine breakage response. Comparison of coal preparation plant feed with a crushed bore core and a bore core prepared by drop-shatter and tumbling (all from the same Illinois No.6 coal mining property) indicated that the size distribution and size fraction composition of the drop-shattered/tumbled core more closely resembled the plant feed than the crushed core.

Jan 1, 1998

By Charles H. Dowding

INTRODUCTION "On many projects, the tendency is to drop interest and support of the geological service as soon as it seems safe. That . . . is a short-sighted policy . . . It would be of further service perhaps to learn whether predictions were reasonably true to life." --C.P. Berkey, 1938 (3) This paper quantitatively compares predicted and encountered shear zone size and location and rock type for tunnels in the Colorado Mineral Belt. The Colarado Mineral Belt is composed of igneous formations with associated metamorphic rocks and faulting. As such, it represents the most complex of provinces for at-grade predictions (27). Shear zone and rock type data are assumed to be statistically homogeneous because of the common geology of the seven tunnels. It is often felt by geologists and engineers that the amount of information from which a geological model is created is so low that any statistical inferences are unstable. In other words, the addition of new data may drastically alter the means and variances suggested by the model. Therefore, predictions are rarely couched in quantitative statistical terms but modified with "probably," "most likely" or "possibly." Comparison of predicted and encountered data from past geological models will permit the "probably's" to be quantified and distilled from the art of geological prediction. To date, two other studies of the validity of predictions have been made (26 and 16). Together, these two studies indicate that

Jan 1, 1976

By F. Ikumapayi

In order to predict and minimize detrimental production problems due to the recycling of process water in sulfide ore processing, the influence of major species, calcium and sulfate in process water on sphalerite flotation was investigated through Hallimond tube flotation, zeta potential, diffuse reflectance Fourier transform infrared spectroscopy (FTIR) and X-ray photoelectron spectroscopy (XPS) measurements using sphalerite mineral sample. Assessment of process water species in flotation was done using deionized water, process water and simulated water containing calcium and sulfate ions in experiments. Hallimond flotation shows increased sphalerite float-ability in process water compared to deionized water, but no significant effect on the presence of calcium and sulfate ions in deionized water using isobutyl xanthate as a collector. The presence of calcium ions reduced the sphalerite?s negative zeta potential, while at higher concentrations, a charge reversal occurred, at about pH 11. FTIR and XPS studies revealed the presence of surface-oxidized sulfoxy and carbonate species on sphalerite at pH 11.5 in deionized water, process water and water containing calcium and sulfate ions. These surface species do not influence xanthate adsorption.

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 V. A. Kharchenko, V. M. Yeremeev

INITIAL PRECONDITIONS To solve tile problem of using the potential reserves of the different coal ranks efficiently (the coal reserves being many times superior in quantity to those of natural gas and oil) one must realise an integrated estimation of the coalfields' and coal deposits' reserves. The actuality of this problem increased especially in the recent years, when the Russian coal industry is being restructured. A structural reorgaisation of this industry, in order to in- crease its efficiency, covers a large spectrum of social and economic problems, comprising a withdrawal from mining the coal reserves that are unsafe from the standpoint of economy and labour protection. According to the information supplied by the 'Rossoogol' company, the Russian coal industry, accounting for abut a fifth of the world coal reserves, has a rather unsafe resource base, for about 5% of the reserves involved in mining do not correspond to the world practice conditions, regarding the coal grade, thickness, geological conditions, proneness to coal and gas outburst hazards. These concern first of all the Moscow, Kizelovsk and Donetsk coal basins, the Prokopievsk-Kiselevsk district of the Kuzbas basin and several other coalfields. Thus, the problem of efficient potential coal reserves' utilization calls for a paramount solution of two basic problems, namely: a) a comparative coalfields' and sections evaluation, and their ranging, in accordance with their industrial importance; b) choosing the 'active' reserves good to be efficiently mined under the present-day conditions. PRINCIPLES OF EVALUATION The investigations that had been carried out make it possible to formulate the following basic principles of the coalfields' reserves' estimation (or assessment). 1. The necessity, possibility and expediency of coal reserves' industrial exploitation are determined by a set of natural, technical, technological, practical, economic, environmental and social factors, thus calling for an integrated and gradual approach to problem solving. 2. One cannot consider the results of same reserves' estimation as being invariable because the former depend both upon constant (mainly natural) and market demands for different grades of coal (those of different power engineering and technological situations), capital investments, material and labour resources, economic (price-forming), technical, technological, environmental and other factors. 3. The rational size and sequence of the operating mines' and reserve sections' resources' involvement into mining depend upon the final aim of posing the problem, whose variants may be (the corresponding limitations being considered) providing for a maximum run-of-mine output (these essentially featuring the past stages of the industry's evolution); providing for a maximum efficiency, measured possibly by the specific or absolute indices of profits an expenses, labour productivity e. al. ; providing for a maximum extraction (maximum integrate utilization) of the known reserves. In this case an important park goes to the economic evaluation (or assessment) of useful application of the by-products being extracted (methane, mine water, rack) compared to the expenses necessary for their utilization. Regarding the steam coal (posing the problem any way it be possible), and extreme limit to the absolute non-efficiency of the coal reserves' exploitation may be posed, using the proportion of energy obtained from the possible amount of the coal to be extracted to the energy consumption necessary to extract, prepare, haul and consume this amount of coal. 4. The most reliable and efficient solutions providing for opening the reserves can be obtained for group of mines, rather than at the level of an individual mine section or an individual mine because it is only in the former case that the possibilities to re- distribute the expenses necessary to develop the good for mining and worse sections and seams can be realized to the greatest extent. Therefore the prospective approach to be considered is the one providing for an integrated organization and choice of the optimum combination of the variants of the ways to use the mines in operation and the free sections of the region in order to satisfy the growing up requirements for coal rather than the one providing for individual mine sections' resources' evaluation. 5. The future possibility and expediency of coal reserves' industrial exploitation will depend upon the prospects to implant at the mines in operation and reserve sections new engineering and technological elements and coal extraction technologies which may essentially change the actual idea concerning the coal reserves worth of being developed in an efficient way. For example, reserves not available to be exploited using conventional coal extraction technology may be regarded as potential resources for in situ gasification, e. g. the operating mines' reserves (non-commercial and non-mined commercial ones), balance and outside balance reserves of the free sections having proved not available to be mined using conventional coal-extraction technologies, as well as the non-extracted reserves of the abandoned and closed mines.

Jan 1, 1996

By Muhammad I. Kabir

A new minitest of in situ uranium leaching processes at reservoir conditions is proposed. The test which consists of a single observation/ production well and two observation wells is designed to permit the evaluation of both restoration and leaching parameters. It is shown that by using a straight forward history matching procedure, sometimes called a type curve analysis, both the rate of leaching and the ultimate uranium recovery can be evaluated based on the results of the test. Type curves are presented for the case in which the reaction rate is kinetically controlled and there are no other minerals competing for oxidant. Other situations can be approached using the same techniques as those presented here.

Jan 1, 1981

By Xiaoting Li, Richard Adasiak, Benjamin Mirabile

Pumpable J-Crib stopping walls and Jennmar high-strength (HS) stopping panel walls are widely used for resisting high-ventilation air pressures and roof support in underground mines due to their installation efficiency, convergence allowance, and better performance than traditional concrete block stopping walls. Based on the different entry dimensions and required maximum ventilation air pressures of up to 36-in. water gauge, 24-, 27-, 30-, and 36-in.-diameter pumpable J-Crib stopping walls, or 18- to 30-in. thick Jennmar HS stopping panel walls can be designed, and their support capacities and convergence allowances are evaluated for better roof control. Pumpable J-Crib stopping walls consist of 0.187-in.-diameter steel wire reinforced, JCrib material-filled bags, and non-wire-reinforced filler bags in between. Jennmar HS stopping panel walls consist of two, 20-gauge (0.04-in.-thick), galvanized, cold-formed steel panels with J-Seal materials pumped in between. Based on the fundamental structural design principles and specifications of ACI 318-19(22), the compression and flexural strength, and shear resistance of the walls can be checked to be compliant with the codes and specifications following the load and resistance factor design (LRFD) method. Specifically, the 0.187- in.-diameter steel wires in the pumpable J-Crib walls and 20-gauge steel panels in the HS stopping walls are integral parts of the walls and are considered to effectively confine the J-Crib or J-Seal materials. These steel elements play a significant role in reinforcing and improving the strength or support capacity of the stopping walls from a structural design perspective. Because of the high ductility of low-density J-Crib/J-Seal materials and steel elements, these walls are good at accommodating or absorbing the roof convergence (≥5% entry height), which cannot be matched by a traditional concrete block stopping wall (Figure 1). Based on the successful design and application of these stopping walls in different mines for improving roof stability and avoiding the air leakage, the authors have formulated the design methodology and verified independently by the finite element analysis (FEA) simulation technique. This design methodology combines the fundamental structural design principles with specific wall dimensions and material properties, facilitating mine engineers and government agencies to understand the in-depth working mechanism of these stopping walls.

Jul 1, 2023

By R. E. Scheffel

In February, 1976, Ranchers Exploration and Development Corporation obtained an option from Foote Minerals, Inc., to evaluate reprocessing of two uranium tailing piles. Foote received ownership of the Naturita and Durango tailing piles in 1967 through the purchase of Vanadium Corporation of America. The option was exercised in November, 1976, after initial testwork proved that a viable extraction technique existed. Preparation for production at Naturita began immediately with engineering and en¬vironmental permitting proceeding simultaneously. Production began in December, 1977, just 22 months after acquisition of the property and-was completed 17 months later.

Jan 1, 1981

By David L. Brostrom, William C. Brice, Perry A. Canton

The business of mineral exploration and development has become much more than simply a problem of locating orebodies and designing mines. Favorable economic analysis of mining and processing feasibility is no longer sufficient to justify the development of a mine. New laws, enacted by virtually every level of government, require comprehensive land use planning, environmental analysis, mineland reclamation, mine safety studies, and many other considerations. This profusion of new laws says in effect, that the mining company is no longer responsible only to its stockholders. Mineral development today can be compared to a partnership. As in all partnerships, the risks and liabilities as well as the benefits are held jointly by all the participants. In addition to the mining operator, this partnership consists of the public, through its many levels of government, interest groups, and individual citizens. Therefore, you'd better know your partner very well - his responsibilities and capabilities. Mining is considered a high risk venture. The risk on the part of the operator and the lending institutions has long been recognized. The risk that the public and its government take in these mining ventures has now been recognized also. The people and their government have learned that they inherit, through abandonment or forfeiture, the mistakes and problems of individuals and companies regarding these natural resources. In looking through the "Mining Engineering Handbook" recently published by the Society of Mining Engineers, I noted that in their discussion of liability for excavations, the authors quoted from the book of Exodus, 20th chap., where Moses said to his people: "When a man leaves a pit open, or when a man digs a pit and does not cover it, and an ox or an ass falls into it, the owner of the pit shall make it good; he shall give money to its owner, and the dead beast

Jan 1, 1974

By D. W. Fuerstenau

One of the more energy-efficient methods for size reduction, namely particle-bed comminution under high compressive loads, was employed to grind eight minerals and ores, using up to 250 MPa pressure. A new approach based on the self-preserving character of the ground product size distributions and a linear energy-reduction ratio relationship is proposed for analyzing this mode of size reduction. The grind ability of the material when comminuted in a particle bed is about one-half that of the single-particle crushing mode, which is in agreement with previous estimates of the relative efficiencies of energy utilization based on direct measurements of surface area produced. The loss of efficiency may be attributed to energy dissipation by particle interaction phenomena in compressed beds.

Jan 1, 1993

By Fred Heivilin

A symposium on the characterization and toxicology of mineral dusts was held Sept. 19 and20 in Washington, D.C. The symposium was held to update the minerals industry, the health community, government agencies and academia on the technical, health and regulatory aspects of minerals to be evaluated by the International Agency for Research on Cancer (IARC).In October 1996, the IARC met in Lyon, France, to discuss and prepare “Volume 68 of the IARC Mono-graphs on the Evaluation of Carcinogenic Risks to Humans on Silica, Silicates, Dusts and Organic Fibers.”

Jan 1, 1996

By Tomás Vergara, Marcelo Vergara, Alonso Pereira, Fernando Concha, María V. Gutiérrez, Fernando Betancourt

Filtration of copper concentrates in the mining industry is carried out in cycles, mainly cake formation, expression, and blowing. Each of these cycles depends on the solid (mineralogy, porosity, particle size distribution), water, and air (density, viscosity, permeability, and relative permeability) properties that define the performance in each cycle. The fundamental phenomenon in cake formation and expression is a one-phase flow (water) through a porous bed of solid particles (the cake), while in blowing it is a two-phase flow (water and air) through the cake. Sometimes, due to a change in the properties of the mineral, a filtration plant is unable to maintain the capacity at the specified humidity. One alternative to solve this problem is to heat the air during the blowing cycle, which in turn heats the water in the cake, this process significantly decreases the water viscosity and only marginally increases the air viscosity. In this work, this strategy was studied at a laboratory scale. The results show a significant decrease in the blowing time as the temperature of the air was increased.

May 24, 2022