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By D. K. Steele

The U.S. Bureau of Mines has investigated the recovery of metals from copper industry wastes by sulfuric acid leaching in a semi-continuous, bench-scale system to solubilize arsenic, copper, and molybdenum. Leach residues contained approximately 40 wt pct PbSO4, suitable for processing at a lead smelter. Arsenic and molybdenum removal from leach solution was examined using solvent extraction with tri-n-butyl-phosphate (TBP), and copper was subsequently recovered by cementation. Bleed electrolyte from a copper refinery was determined to be a viable source of acid for the initial leach stage.

Jan 1, 1991

By Goran Mathisson

In several investigations the correlation has been studied between the fuel consumption and the reducibility of the charged material in the zone of indirect reduction in the blast furnace (1, 2, 3, 4, 5). It has been found that increased reducibility allows greater gas utilization which accordingly lowers the fuel consumption. In direct reduction units this relation has not yet been fully established partly owing to the lack of accurate test methods and experience with different materials as regards reducibility. However, since the production rate in a direct reduction unit is closely related to the time required to reach high degrees of reduction, the reducibility must be regarded as one of the most important properties of the material.

Jan 1, 1979

By D. M. Thayer

Dry magnetic separation of feebly magnetic minerals and wet magnetic separators suitable for highly magnetic minerals have been commercially utilized for a number of years. The beneficiation of magnetic taconite ores by conventional, low intensity magnetic separators in combination with flotation is well established technology (1). However, over the past decade much effort has been expended in developing techniques for the upgrading of feebly magnetic oxidized iron ores (2,5). Such research and development have been stimulated by the limitations of the above-mentioned techniques in application to ores containing fine material; for example, ores requiring fine grinding for liberation or slime-forming ores such as the oxidized taconites or semi-taconites. Wet separation has allowed the reduction of the minim particle size that can be effectively treated from that allowable for high intensity dry magnetic separation. Consequently, the development of high intensity wet magnetic separation methods has received much attention in the search for an effective method to treat the weakly magnetic minerals. Recently Lawver and Hopstock (5) have provided an excellent review of the state of know- ledge of this technology. In addition, a valuable compendium of available laboratory and commercial devices is included. Nonetheless, the feasibility of the use of high intensity magnetic separation in a particular application is largly empirical in nature and requires experimentation to chazacterize the response of a particular ore. The purpose of this paper is to assess the incorporation of high intensity magnetic separation for the beneficiation of a non- sliming specular Michigan hematite. Particular attention is focused on the existence of a lower limit of particle size that can be treated effectively and is inherent in the use of these techniques. Additionally, a rationalization of various models presented in the literature is attempted.

Jan 1, 1979

By R. Hogg

The sedimentation and dewatering characteristics of flocculated slurries are determined by the size and density of the flocs. Studies of the formation and growth of flocs reveal that floc size is limited by breakage of the agglomerates due to shear in the suspension. Floc density measurements indicate that there is an inverse relationship between size and density. This relationship appears to be determined primarily by the nature of the solid particles, especially the size distribution, but it appears that agitation of the suspension can lead to some densification of the flocs, in addition to breakage. It follows that both size and density can be controlled to a significant extent by regulating the physical conditions (agitation, method of flocculant addition etc.) during flocculation. The use of physical floc conditioning to control the sedimentation and dewatering (filtration) behavior of flocculated, fine-particle suspensions is evaluated in terms of its effects on floc size end density.

Jan 1, 1986

By A. Vien

A major thrust in Industrial automation over the past few years has been the exploration and incorporation of higher level (supervisory and optimizing) control strategies aimed at more fully exploiting the existing control system infrastructure. There is little doubt that this is worthwhile pursuit, however, one must not lose sight of the maintenance and development requirements of the basic control system. Statistics on loop performance indicate that this too is an area of considerable potential. In this paper the authors look at some of the less elegant, but arguably equally important issues in mineral processing plant control. The focus is on controller structures, loop tuning, sampling, filtering, and some of the other principal components of the regulatory control foundations. In many cases significant improvements In control performance can be achieved by getting back to the basics.

Jan 1, 1998

By Richard C. Travis, Edward C. Dowling

A detailed kinetic study of the flotation installation at the Metcalf Concentrator revealed a rougher flotation circuit operating under less than optimal conditions. Based on these results, a project was initiated to approximately double the rougher flotation capacity by increasing the installed volume from 780 cubic meters to 1,560 cubic meters. This project, which was completed in early 1991, replaced the existing cells with larger capacity cells. A comprehensive analysis of results demonstrates significant metallurgical benefits. INTRODUCTION/BACKGROUND The Metcalf Concentrator is located in Morenci, Arizona and is part of Phelps Dodge Morenci, Incorporated. Sumitomo Metal Mining of Arizona is a minority partner in the operation with 15% interest. The Metcalf Concentrator is the newer and smaller concentrator and operates in conjunction with the Morenci Concentrator. Operating capacity at Metcalf is currently 48,000 TPD of ore on a 365 days per year basis. The Metcalf Concentrator was commissioned in 1974 at a 30,000 TPD operating rate for the purpose of treating ores originating in the Metcalf Mine. The mine takes its name from the former town of Metcalf which was named for Robert Metcalf who was credited with discovery of the mineral district. The predominant ore mineral is chalcocite which occurs as a secondary enrichment closely associated with iron pyrite. The host rock is largely granodiorite and quartz monzonite. The original rougher flotation circuit consisted of a total of six rows of Wemco 8.5 cu. meter (300 cu.ft.) flotation machines treating pulp from the two grinding sections (Sections A and B). Each row was configured in a 4-4-4 arrangement. Shortly after startup the throughput rate was increased significantly without sustaining a recovery loss. This increase was achieved by raising the target value for the particle size of the cyclone overflow which is the rougher flotation feed. Shortly thereafter, the capacity of the pump and cyclone began to restrict further increases in throughput. After testing, two 66 cm. (26-in.) cyclones and a 12-in. X 12-in. centrifugal pump replaced the original classification equipment. As a result of these modifications, the grinding section could routinely process in excess of 42,000 TPD. Due to the increased tonnage rates, limitations in the rougher flotation circuit performance were observed. In an effort to improve the performance of rougher flotation with this higher throughput, the speed of the flotation mechanisms was increased from 230 RPM to 250 RPM. This change improved the suspension of the coarser feed and induced additional air into the pulp. It did, however, apply additional strain to the motors, shortening their life. At these elevated processing rates, it was found that the dart valves operated almost entirely at the top end of the control range and could not efficiently handle upsets from the grinding section. This was remedied by lowering the end overflow weir and allowing some of the tailings to overflow the end weir. However, in order to obtain acceptable recovery levels, the flotation pulp was maintained at a high level, adversely influencing the rougher concentrate grade.

Jan 1, 1993

By Charles W. Daugherty, Eugene B. Waggoner

In the construction of recent subsurface projects, particularly mined openings and transit systems, it has been apparent that final costs were often greater than first expected. Because many overruns result from alleged differing site conditions, there was reason to suspect that pre-bid geotechnical investigations were inadequate for predicting subsurface geology. In 1982, nine federal agencies involved in constructing underground facilities funded the National Academy of Sciences' U.S. National Committee on Tunneling Technology (USNC/TT) in an effort to evaluate and clarify the relationship between geological knowledge and construction costs. That committee formed the special Subcommittee on Geological Site Investigations, which spent more than two years studying underground projects by examining contract documents and geologic reports, making site visits, and interviewing owners, designers and contractors. The study ultimately encompassed 84 mined tunnels and chambers and three deep shafts in the United States and eastern Canada. Data collected on geologic exploration, design and construction were analyzed and conclusions reached concerning whether and how site investigations need to be improved. The results of the study are published in two volumes by the National Academy of Sciences. Due to the recent completion date of the main report, it was not possible to prepare a summary for inclusion in the RETC Proceedings. However, the complete two-volume report, "Geotechnical Site Investigations for Underground Projects," can be obtained by writing or calling: U.S. National Committee on Tunneling Technology National Academy of Sciences 2101 Constitution Avenue, N.W. Washington, D.C. 20418 (202) 334-3136

Jan 1, 1985

The criteria used to design, operate and control flotation columns are significantly different from those used for conventional flotation machines. Therefore, a computer-aided design (CAD) package for column flotation has been developed. The simulator has been set up using guidelines developed at Virginia Tech and information published in literature during the past few years. The simulator is useful for determining the proper design and appropriate operating conditions for a column cell to fit a specific set of conditions. In addition, the simulator can be used for training in column operation and for trouble-shooting operational problems. Simulation results have been found to be in good agreement with experimental data obtained using laboratory columns.

Jan 1, 1990

By Błażej Doroszuk, Robert Król

Mill drums are essential equipment in various industries for the comminution of materials. To optimize their design and operation for efficient and cost-effective processing, it is necessary to understand the behavior of the mill drum under various operating conditions. This study investigates the behavior of a mill drum using a combination of experimental and numerical methods. The factors that can influence the behavior of the mill drum, such as the size and shape of the mill, grinding media, rotational speed, filling level, and material properties, are discussed. Experimental methods can be limited in obtaining measurements of internal processes, so numerical methods such as the Discrete Element Method (DEM) are also used to simulate the behavior of the mill drum. In this study, the experimental measurements of the motion of the grinding media are obtained using computer vision methods, and the numerical simulations are calibrated using the experimental results. A designed experiment approach is used to evaluate the performance of the mill drum under various operating conditions. The results of this study can be useful for the design and optimization of mill drums to achieve better performance and energy efficiency.

Jun 25, 2023

By L. Wilfley, G. A. Calboreanu

"Wear resistant 25% Cr hard cast iron is a heterogeneous material containing 1,600 HV hard chromium carbides embedded into martensitic matrix. It is the most used cast alloy in wear operations of mining industry, chemical industry, fracking, etc. During the last decade the hard iron manufacturers increased wear resistance in the following ways: • Reduce / eliminate harmful retained austenite and chromium + carbon depleted areas especially within eutectic carbide cells. In Max5A the improved 25% Cr alloy brittle carbide eutectic cells are removed which increases toughness. • Delay delamination between chromium carbides and matrix particularly during sliding abrasion wear. Delamination commences at the weakest chromium carbide – matrix contact lines exposing chromium carbide to localized bending stresses which gradually chip carbides away. Once the chromium carbide base is removed accelerated wear degradation occurs. • Evenly randomize and avoid discontinuities in the chromium carbide network. When chromium carbides are perpendicularly positioned to the wear surface wear rate is minimized. On the contrary the highest wear rate takes place when 3D oriented chromium carbides are parallel with the wear surface. In this case chromium carbides cover top network discontinuities being just matrix supported which leads to rapid carbide cracking. The new hard iron discussed in this paper fulfils and exceeds the above requirements increasing the casting service life around 250% in various wear operations. Physical and mechanical properties of Max5A are compared with well manufactured Industry Standard 25% Cr hard iron. The paper presents practical accomplishments obtained in the USA, South America and India. INTRODUCTION The chemical composition of Max5A and Industry Standard 25% Cr are identical like specified by ASTM A532 Class III Type A requesting C = 2.0-3.3% and Cr = 23.0-30.0%. Manganese, silicon, nickel, molybdenum and copper contents are limited by maximum accepted quantities.[1] According to published data the eutectic chromium carbides chemical composition of 25% Cr hard iron is Cr = 54% and C = 8% which indicates strong depletion of chromium and carbon along chromium carbide - matrix interfaces. [2] The optimized Max5A microstructure has minimized retained austenite, is devoid of harmful carbide eutectic cells and has the most advantageous chromium carbide and secondary chromium carbide morphology. This favorable microstructure increases wear resistance, mechanical properties and corrosion resistance. The paper compares the Max5A microscopic features and mechanical – physical properties with similar properties of high quality Industry Standard 25% Cr."

Jan 1, 2017

By R. M. LaDouceur, C. A. Young, W. Nicholas

"Salicyl hydroxamic acid (SHA) adsorption on two rare earth oxides, lanthanum and yttrium, was studied. Adsorption at both surfaces occurred in two stages. Initially, the monovalent anionic collector appears to adsorb significantly and instantaneously in stage one and then becomes sluggish in stage two presumably because the initial stage passivated the surface. By comparison, adsorption during the sluggish second stage with lanthanum oxide did not stop but with yttrium oxide essentially stopped. Surface monolayer formation could not be observed in either of the two cases. Both stages appear to be due to multilayer formation of a surface precipitate as proven by Raman Spectroscopy for lanthanum oxide but inferred for yttrium oxide. Unlike the use of octyl hydroxamate as stated in the literature, increasing temperatures do not increase adsorption such that flotation of rare earths with SHA may be performed at room temperature leading to improved recoveries as well as significant energy savings. INTRODUCTION Rare Earth Elements (REEs) are comprised of the 15 lanthanides along with scandium and yttrium. The REEs with lower atomic numbers are relatively more abundant and are categorized as light REEs. On the contrary, the REEs with high atomic numbers are less abundant and referred to as heavy REEs. They are not really “rare” and, in fact, some of them are quite abundant in the Earth’s crust; it’s just that they “rarely” exist in concentrated deposits but rather are found dilute and practically everywhere. REEs share very similar physico-chemical properties but, because many of these properties are unique, they have gained enormous importance in recent times making several critical and strategic pending their use. For example, REEs have found extensive application in phosphor coatings, magnetics, and automobile cracking catalysts. Early technology was not advanced enough to produce commodity level production of the REEs and can be attributed to REE-bearing ores containing many rare earth minerals (REMs), not just one. In this regard, there are numerous REMs but bastnaesite, xenotime and monazite are considered to be the most common. Bastnaesite is a fluorocarbonate whereas the other two are phosphates. The separation of the individual REE-bearing minerals from the ores is not an easy task since they also exhibit very similar properties. Consequently, when flotation is used, it is done to produce bulk REM concentrates. For REM flotation, it is well established that octyl hydroxamate is more selective than fatty acid collectors (Pradip and Fuerstenau, 1983 and 1991). Liu et al. (2006) indicated that hydroxamates are the most common collectors for REMs. Pavez et al. (1996) studied the adsorption of oleic fatty acid and octyl hydroxamate on REM surfaces. Cui et al. (2012) studied the interaction of octyl hydroxamate with various REMs surfaces using various spectroscopic techniques such as XPS, FT-IR and Raman along with electron microscopy and concluded that such techniques offer ample information to establish adsorption mechanisms. Ren et al. (1997) studied napthalenic hydroxamic acid as a unique collector for flotation of REMs from bastnaesite ore and indicated that more selective collectors were needed. From these studies, it is clear that the quest of more selective collectors is essential but understanding how they adsorb at mineral surfaces is prerequisite to improving flotation performance."

Jan 1, 2016

By Elza Burch

Introduction The feeding to zero concept involves feeding 600 µm x 0 (28 mesh x 0) size raw coal to heavy media (magnetite) cyclones along with the +600 µm (+28 mesh) size coal. Traditional circuits employ desliming or removing the 600 µm x 0 (28 mesh x 0) size fraction and feeding the cyclones +600 µm (+28 mesh) size coal. The feeding to zero concept recirculates 600 µm x 0 (28 mesh x 0) fines in the circuit. At the same time, a portion of the fine material is continuously withdrawn and recovered. This, in turn, prevents a fines buildup. This concept eliminates desliming screens and secondary fines circuitry for recovery of 600 x 150 µm (28 mesh x 100 mesh) coal. The result is a very simple circuit. Feeding to zero at Island Creek Island Creek Corp. was the first involved with the new concept in 1976. The company needed a temporary plant for the 9.5 mm x 0 (0.4 in. x 0) raw coal at its Pond Fork mine, near Madison, WV, while a full-scale plant was being designed and built. At that time, the Childress Corp., of Beckley, WV, became interested in the feeding to zero concept. Island Creek awarded a contract to Childress to build a single cyclone modular plant, incorporating this feeding to zero concept. The plant was erected in three months. The Pond Fork modular plant proved successful in attaining the desired feed rate of about 63.5 t/h (70 stph), while maintaining good separating efficiencies and low magnetite consumption rates. The 9.5 mm x 150 µm (0.4 in. x 100 mesh) clean coal was recovered and the 150 µm x 0 (100 mesh x 0) size was disposed of to waste. The full-scale plant was completed about two years later and the Pond Fork modular plant was moved to Holden, WV. There, it was incorporated into the Holden 29 preparation plant as a separate circuit for cleaning -25 mm (-1 in.) coal. In 1976, a similar plant was installed in Virginia by another company. These two plants are believed to be the first two operational plants in the United States incorporating the feeding to zero concept. Island Creek subsequently contracted with Childress for an identical plant at the Coal Mountain operation in West Virginia. The plant operated for three years before the mine was closed. The unit was then moved to the Spurlock mine near Martin, KY where it continues to operate. The successful operation of the Pond Fork and Coal Mountain plants before and after relocation proved both the performance and moveability of this type of circuit when constructed in a modular fashion. Since the first two plants were built, Island Creek has incorporated the feeding to zero circuit in nine additional plants. A grand total of 33 cyclones have been installed using this concept. One is the Providence mine, near Providence. Providence preparation plant Island Creek contracted with J.O. Lively Corp. of Glen White, WV in July 1978 for the construction of the Providence preparation plant. The plant began operation in February 1979. The construction period was about halved by building the plant with modular design concepts. Prefabricated sections, floors, and sides were brought in as units and then bolted in place. The Providence plant has a good track record of processing coal at a feed rate of 454 kt/h (500 stph). Feed coal to the plant has an average ash content of about 18% and sulfur content of about 4.5%. It contains about 22% refuse. The coal product has an average ash content of about 8% and sulfur content is about 3%. Raw West Kentucky No. 9 seam coal is conveyed from a box cut in the Providence mine to a rotary breaker. The breaker is fitted with 74 mm-diam (3 in.-diam) opening breaker plates. Therefore, it is well suited for removing trash, roof bolts, wood, and pyritic balls that are common in Illinois Basin coal. The -75 mm (-3 in.) coal is conveyed from the rotary breaker to

Jan 8, 1987

By G. Bonifazi

Certain deposits of incoherent materials which are near the surface and exploitable by open-pit mining may be characterized in real time before they are mined through an analysis of ground-surface imagery. If the slice to be stripped is not overly thick, in fact, surface texture and color properties may be extrapolated to its entire thickness. Image analysis techniques to process ground surface images acquired in situ (or in the laboratory shortly after their acquisition) which yield pattern vectors representative of RGB color component distributions and HSB texture parameters have been developed. The techniques were applied to a mineral sandy ore deposit containing three different lithotypes. Geostatistical analyses showed that the data used to characterize the lithotypes were reliable. The correct recognition of the lithotypes was carried out using a multi-barycenter classification algorithm. Such in situ image-analysis procedures open the way to the proper selection of the ore to be mined, or the proper way for run of mine (r.o.m). processing, or the appropriate blend to be processed before or at the same time as the mining process itself is enacted.

Jan 1, 1999

By William R. Nash

Case histories of water problems in underground mine shafts and tunnels are outlined to illustrate varying degrees of sophistication in grouting techniques currently applied in such construction. The interception of water underground can be one of prior knowledge, i. e., the site investigation provides necessary data to plan a grouting program prior to construction start, or excavations may encounter "changed conditions" or unplanned water. The latter includes catastrophic inflows and emergency situations where the structural integrity of a building or concrete shaft lining are jeopardized. The presentation of this paper outlines some applications of grouting work in underground mine construction throughout the United States in a variety of geological provinces and for a variety of minerals.

Jan 1, 1984

By K. L. Gabby

Since mercury is chemically similar to silver, it tends to dissolve along with the silver and gold during cyanide leaching. Ultimately mercury is recovered along with the silver as mercury/silver/gold amalgam which requires retorting to remove the mercury prior to refining the silver and gold. A possible approach to separating mercury from silver and gold in cyanide leach solutions is to take advantage of the replacement reaction between [Hg(CN)4]2- and Ag2S, which results in precipitation of mercury as HgS while dissolving silver as [Ag(CN)2]-. Results are presented demonstrating that mercury can be removed from cyanide solutions with high effectiveness, while, at the same time, recovering additional silver in solution. Initial experiments with static flasks in the laboratory showed a high degree of mercury removal could be achieved. Results using actual metallurgical plant products in bottle roll tests indicate the process works on a larger scale, but also results in significant gold losses if the solution being treated has high gold levels, such as carbon stripping pregnant eluent solution.

Feb 23, 2014

By M. S. Celik

Beneficiation of boron minerals by flotation with well defined systems has been very little investigated in the literature. The difficulty arises from the fact that boron minerals cover a spectrum of solubilities. Colemanite, a calcium containing relatively insoluble boron mineral, and borax, a highly soluble boron mineral have been used to exemplify the flotation of boron minerals. The floatability of both minerals using sodium dodecylsulfonate and dodecylamine hydrochloride exhibits considerable differences in their response to flotation. This has been ascribed to variations in the solubilities 0f boron minerals and the presence of the cations in the lattice structure. Flotation tests along with zeta potential measurements are used to elucidate the mechanisms involved in the flotation of boron minerals.

Jan 1, 1992

By H. Shafick Hanna

The major problems are identified the high phosphorus content and the fine dissemination or encapsulation of mineral impurities, in, ,the iron bearing minerals, A process which combined extreme fine grinding techniques with magnetizing roasting or flotation separation methods, is suggested. By these methods concentrates assaying 60-61 percent Fe exceeding the widely accepted upper limit of 56 percent Fe set for these ores, were produced with 75-90 percent iron recovery. Moreover it was shown that the phosphorus content of these concentrates could be reduced from 0.25 to 0.03 percent phosphorus by acid leaching. A brief discussion on production costs based on current and future developments in iron ore processing is given.

Jan 1, 1980

By Garry W. Warren

The electrochemical oxidation of chalcopyrite in the absence of-a chemical oxidant was studied in ammonia/ammonium sulfate solutions by anodically polarizing electrodes prepared from massive chalcopyrite. The effects of pH, cupric and ammonia ion concentration were systematically evaluated. Kinetic parameters were obtained from the anodic charge-transfer regions of the anodic polarization curves. Increasing ammonia and hydroxyl ion concentrations increased the rate of oxidation. A mechanism is proposed which explains the first order dependence of (NH3) and (OR) associated with a single electron transfer process. An empirical relationship which accurately describes the dependence of the rest potential (i.e. zero current potential) on solution composition is also given.

Jan 1, 1978

By Siegmund Babendererde

Introduction Lyon, with 1.1 million inhabitants, is the third largest city in France after Paris and Marseille. It is a lively industrial centre. Lyon lies at the intersection of many transport routes and at the fluence of Rhone and Saone. Like all large cities, Lyon is attempting to alleviate inner city traffic problems by constructing a metro. Metro construction in Lyon began with tunnels built by cut and cover methods along the streets. Deep diaphram walls, which were back anchored and toed into imperveous strata, formed the excavation walls. It is a very economic and simple method. However during the construction period, it has the disavantage of strangling urban life for several years. It is therefore increasingly being replaced by subsurface tunnelling methods. As it was decided to build the metro tunnel between the Guillotiere and St. Jean stations, and thus cross under the Rhone and Saone, the client, Semaly, put the contract out for a competitive tender in which the general parameters were defined. Line, level and construction method were left to the contractor. There was an extensive examination of the tenders, during which reference projects from competitors in Japan were visited, as well as in Germany. Finally the consortium of Dragages, Campenon Bernard, Colas and Hochtief was awarded the contract to drive this tunnel section with an hydro-shield and an extruded steel fibre concrete lining, a Hochtief design.

Jan 1, 1987

A comparison of gold, platinum and sulfide ion selective electrodes was made in this investigation for determination of Eh in solutions containing sulfide ions as a model system. The comparison was made in terms of the accuracy of the measured response, response time, pH dependence, and reproducibility of measurement. In addition, the sensitivity of each electrode towards changes in sulfide ion concentration was measured in terms of the change in measured Eh per decade change in concentration of sulfide ions. On the basis of a detailed investigation of the electrode processes occurring at the surface of gold and platinum electrodes, it was postulated that within the pH range investigated these electrodes respond to changes in the hydrosulfide ion, HS-, concentration in solution. In comparison, the sulfide ion selective electrode is known to respond to changes in the sulfide ion, S-, activity in solution. Although the potentials measured with gold and platinum electrodes were mixed potentials, pretreatment procedures can be used to obtain a signal suitable for process control. For control of sulfide ions in flotation pulps, the use of a gold electrode pretreated at 1.0 V with respect to a standard hydrogen electrode is recommended.

Jan 1, 1990