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By R. D. Clark, Heinz Althoff

"The advantages and disadvantages of a mobile crushing plant with a belt conveyor system compared to a stationary crushing plant with haulage by heavy trucks are described. The capital expenditure and operational expenses are compared and the individual cost factors are explained.Every open-pit mine or quarry, however, is different and has to be assessed and analyzed individually in order to find the optimum and most economical solution. Due to the high truck haulage costs, many open-pit mines tend to the more economical solution with mobile crushing plants and continuous belt conveyor operation. IntroductionFor the material haulage within an open-pit mine the heavy truck is the most flexible means of transportation. However, due to the high fuel consumption, wear and labour costs, it is also the most expensive one. Therefore, many open-pit mines are looking for alternative solutions for the material transport A mobile belt conveyor system is most likely to meet the requirements because it is efficient and economical with regard to the power consumption. The material , however, has to be crushed at the mining face to lumps which are suitable for conveyor operation. For this purpose a mobile crushing plant is required.In some open-pit mines the crusher is also used to crush the material to lumps of smaller particle sizes in order to save crushing costs in the preparation plant.Although a cost comparison between a system with stationary crusher and truck haulage and a system with mobile crusher and belt conveyors favours in most cases the belt conveyor system, up till now only a few open-pit mines and quarries benefit world-wide from the advantages of the mobile crushing plant with continuous belt conveyor operation."

Jan 1, 1986

By T. F. Bower, M. C. Flemings, H. D. Brody

A series of experiments are reported which show thai reasonable assumptions for analysis of solute redistribution in solidification of castings and ingols illclr(de: negligible undercooling before nuclcation of solid phases, negligible increase of solute in advance of the tips of growing dendrites, complete diffusion within the Liquid oer distances the order, of dentrite arm spacings, platelike dentrite t/o?$holog?.. Qzmli/ali,e and periment is obtained JOY Al-Cu alloys front measurements of amount of nonequilibrium second phase, minimum solute content at the centers dentrites arms (after solidification), and 3) change in minimum solute content in dentrite arms during solidification. In a previous paper' analyses were given describing solute redistribution in dendritic solidification. The analyses were undertaken primarily to permit prediction of "severity of microsegregation" in a casting or ingot after solidification and cooling to room temperature, but can also be used to predict "local fraction solid' as a function of temperature in the solidifying alloy. Numerical results were given for Al-Cu alloys. The aims of this paper are to examine the validity of some of the assumptions of that analysis, and to compare predictions of the theory with experiments. All experiments reported are on Al-Cu alloys, the bulk on Al-4.5 pct Cu alloy. EXPERIMENTS Experiments conducted were on small ingots, 41/2 in. in diam by about 10 in. high. solidified in a controlled thermal environment (i.e., in a resistance-heated, re-circulating air furnace, whose temperature could be controlled to 1°C). In one series of experiments, the ingots were solidified very slowly (solidification time was varied from 8 to 1000 hr) and at constant rate of heat extraction. In a second series of experiments the ingots were solidified ( 'unidirectionally". Fig. 1 is a sketch of the method of making the two different types of ingots. The unidirectionally solidified ingots were fully columnar, and the slowly solidified ingots had equiaxed structures, Fig. 2. In making the slowly solidified ingots, the mold employed was carbon steel, coated with inert refractory. Typically the mold and its environment were heated to 700°C. The molten alloy was then poured into the mold and the temperature of the system equilibrated above the liquidus. The environment of the mold was next cooled with a controller responding to a differential thermocouple that measured the difference between the temperature of the melt and the temperature of the mold environment. The differential thermocouple consisted of one chromel-alumel junction in the melt connected in opposition with a second junction in the furnace. The potential across the differential thermocouple was input to the controller which acted to keep this potential at a set. constant value; hence the temperature difference between the two junctions was maintained nearly constant. Major resistance to heat flow in the above apparatus was between the metal surface and the furnace atmosphere, not in the metal itself. By maintaining a constant temperature difference between the metal and the furnace atmosphere, the controller maintained an essentially constant rate of heat extraction during solidification, regardless of amount of heat released by the metal per unit temperature change. When the metal temperature dropped just below the eutectic temperature. the ingot was removed from the furnace and water-quenched. The unidirectionally solidified ingots were cast in a similar mold to the foregoing except that provision was made to extract heat through the bottom face of the mold by passing air or water through the mold base, Fig. 1. For these experiments, metal and mold temperature were again equilibrated above the liquidus. Temperature of the mold environment was then held

Jan 1, 1967

By Marc D. Melker, Geoffrey R. Lane, Michael Jahraus

"The Cresson gold deposit is located in the Cripple Creek district of Colorado. The deposit is composed of volcanic intrusive rocks and volcanogenic sediments where gold mineralization is associated with veins and cavities within a diatremal breccia.The purpose of the current study was to:1. Determine the penetration potential of the Cripple Creek Breccia.2. Assess the relationship between penetration and gold recovery within the Breccia.Two methods were applied using a cyanide solution containing ultraviolet sensitive dye to elucidate pathways of penetration within the rock. The first involved stagnant soaking of twoinch square pieces within the solution. The second method involved column leaching of ~100kg of Breccia crushed to P80 12.7 mm (0.5 inch). Samples were selected at specified intervals to capture the penetration progress of the solution. Photomicrographs and image analysis were utilized to illustrate and quantify the degree of penetration.Results indicated that the maximum penetration for the Cripple Creek Breccia (85 to 97%) due to a porous and permeable groundmass, containing large non-porous lapilli fragments. In addition, ~90% of leachable gold was recovered with only ~50% rock penetration.INTRODUCTIONDye penetration studies were first developed at SGS-Lakefield Research for the purpose of determining the fluid penetration amenability of differing rock types. The method was to place a large rock or drill core in a pan with 1 to 2 inches of a pre-mixed, water based fluorescent dye. The sample was sequentially cut after prescribed time periods to reveal a fresh surface. The cut surface was photographed under long wave ultraviolet light, and the same piece returned to the dye and allowed to sit until the next prescribed cutting."

Jan 1, 2006

By Stanley Dempsey

INTRODUCTION The mining industry has been under- going a significant restructuring for the past several years. Some firms have been liquidated. Some have been merged with other firms. A number of mining divisions of larger firms have been spun off, in some cases in leveraged buyouts by management. Recently, there have been a number of spin outs of gold subsidiaries. There are also a number of start-up situations where new firms are emerging as significant gold producers through mine development or acquisition. This restructuring has created an attractive market for specialized mining finance firms, and a number of firms have recently entered that field in the United States. These firms provide financial advisory and valuation services to mining firms and project developers, and access to venture capital. In some cases, they invest for clients or for their own account. I have been asked to describe how our firm works, and what kind of information we expect to receive when mine developers seek our assistance in financing a project. HOW WE OPERATE Denver Mining Finance works for both mine developers and investors. We also provide expert advice on mining issues to governments and to firms who are in litigation over mining issues. A lot of our work involves valuing mineral interests for various purposes. Our principal expertise is in the technical, commercial, and financial area. We help put mining deals into shape commercially so that they can be financed, and we help investors get into mining deals that are commercially sound. We work on projects, but we also advise mining firms on the structure of their business and how to make the most out of one or several mines that they might own. In addition to our work for clients, we are also prepared to invest in mining projects for our own account, and to recommend gold properties for acquisition by our parent company, Royal Gold, Inc. We are, to some extent, a venture capital firm, in that we sometimes work with projects that are not "bankablerV1 and are prepared to invest in attractive early-stage situations in mining. However, I want to emphasize that our principal focus is on development of commercially attractive mining projects that can be brought into production in the best possible financial terms. If we can make a project qualify for a project loan from a commercial bank, we will not advise the developer to use high cost venture capital. For the purpose of this paper, I will focus primarily upon our functions with respect to development of precious metal projects. In a typical case, a firm with an advanced-staged project approaches us for help with financing the final feasibility study, and construction and operation of a mine. In some cases, the developer has already developed a financing plan and wants only our access to investors. If we agree that the project is sound and that the financing scheme will be attractive to investors, we will help the developer reach investors on a commission basis. In such cases, we are particularly insistent that the financing plan and our participation meet all regulatory requirements, particularly compliance with state and federal securities laws. In most cases, however, the developer asks us for advise both on how to structure the financing, and how it can be implemented. Here we work for fees based on hourly rates for advisory work,

Jan 1, 1987

By A. Pagliero

The coulombimetric reduction method has been used as an evaluation technique for Pb-Ca-Sn anode corrosion in the LIX-SX-EW process. The proposed methodology involves three consecutive steps, i.e., cleaning, oxidation and reduction. They are carried out in the same experimental unit, in solutions containing 180 g/L H2S04 with and without the addition of cobalt. The authors have worked in a wide range of oxidation current densities, from 150 to 500 A/m2 for in a period from 0.5 to 120 hours. It can be concluded from the results that there would be a linear relationship between the applied electrical charge I*T (Coulomb) and the anode corrosion expressed as transformed Pb mass (mg), for solutions with or without the addition of cobalt. The exception is that for the last one the corrosion rate is 5 times faster, which was observed for every layer. From the evaluation of different corrosion layers, one can point out that its composition is mainly in the form of sulphate. In the presence of cobalt, about 67% of Pb would be as sulphate (PbSO4), about 27% as a basic oxide (PbO) and only 6% as oxide PbO2 on the surface. From this result, it appears that great transformations occurred in the metal/sulphate interface. Without the addition of cobalt, the proportion of oxide Pb02 on the surface increased up to 11.5%. According to the previous statement and considering that Pb was found mainly as sulphate, the authors recommend its determination be used as quality selection criterion. Its evaluation should be performed in the acid solutions with 200 ppm cobalt addition at an oxidation current density of 300 A/m2 for a period of 24 hours.

Jan 1, 1999

By Wm. Terry Siemer

Abstract. The value of industrial rocks and minerals produced in New Mexico in 1978 exceeded 227 million. 1e state ranked first in the nation in production of potash and perlite and was a leader in pumice production. Other commodities with sizable production in New Mexico include barite, caliche, clay, dimension stone, fluorspar, gypsum, humate, limestone, mica, salt, sand and gravel, scoria, and silica sand Non metallic resources are distributed throw out the state, but development and exploitation is concentrated in western and southeastern New Mexico. The value of industrial rocks and minerals within the state grew about 9 percent to more than $227 million in 1978. Continued growth, however, depends on continued demand, exploration, and development.

Jan 1, 1980

Gravity-driven rock falls account for a large proportion of all fatalities and injuries in underground mines. Potentially unstable blocks of rock are found in the hanging wall of excavations, which may be delineated by natural joints or stress fractures. These blocks, referred to as keyblocks, will fail if their weight exceeds the support capacity or if they are located between support units. This paper describes a computer program which simulates blocks in the hanging wall of an excavation based on measured properties of fractures and joints. Keyblocks are identified and placed at random locations in a predefined excavation. The keyblocks are tested against the support to determine whether they fail the support, drop out between the support or are stable. By repeating the procedure several thousand times, statistics of the probability of failure of the keyblocks are obtained. This approach allows the relative hazard associated with different types of support or excavation orientations to be obtained. It is concluded that a probabilistic approach is suitable for the evaluation of supports effectiveness in situations where large numbers of discontinuities or stress fractures are exposed in excavations.

Jan 1, 1998

By Francis Dakubo, Courtney Young, Avimanyu Das

"Hydrometallurgy of copper ores typically involve acid leaching followed by solvent extraction and electrowinning (SX/EW). EW is the final purification stage in which a voltage is applied across two electrodes causing the copper ions to reduce and deposit on the cathode. Applied voltages are near 2 volts with current densities around 300 amps/m2. In this study, research was undertaken to decrease the energy consumed in the EW process. To do this, various cell designs were constructed and tested in which a membrane was used to separate the cathode and anode compartments with concentrated solution from SX serving as catholyte and raffinate from SX serving as anolyte. Therefore, copper in the concentrated solution would be deposited on the cathode while ferrous in the raffinate would be oxidized to ferric at the anode. Because it is assumed that ferric acid leaching is used, resulting ferric would be recycled to the leaching process. A positive side effect of the envisioned process would be the elimination of oxygen evolution at the anode and thus the prevention of acid-misting. After a cell design was selected, it was used to test a number of variables: flowrate, concentration, temperature, ion-exchange membrane, and applied voltage. Following experiments in which variable ranges were determined, statistical modeling was conducted using STATEASE software. Optimal results indicate the voltage can be decreased to 1.5-1.7 volts with current densities of 300-400 amps/m2. Significant reduction in the energy consumption was accomplished by reducing the applied voltage. Furthermore, current efficiencies and copper production rates was increased by application of the membrane.INTRODUCTIONLeaching of copper ores is achieved using mostly sulfuric acid. The copper pregnant solution is contacted with an organic solvent or resin adsorbent which picks up copper ions as in Equation 1:"

Jan 1, 2016

Law of the Sea. - The Ninth Session of the Law of the Sea Conference resumed July 28 at Geneva and will run for 5 weeks. The 160-nation Conference began in 1973. Two of the major factors to be discussed are the transfer of technology to the UN-controlled Enterprise or mining authority, and how the 36-member Council will make decisions. As the session completed its third week, delegates expressed little hope of completing a treaty text this year. There was a feeling, though, that substantive issues might be settled so that a treaty text can be completed next year. The Deep Seabed Hard Mineral Resources Act signed by the President on June 28, provides an interim regulatory procedure for ocean mining activities conducted by U.S. nationals that may be superseded by a law of the sea treaty.

Jan 1, 1980

By Warren E. Morrison

Stimulated by 4 years of continuous economic growth without inflation, demand for energy in the United States reached an alltime high in 1964. Total value of production of goods and services (gross national product--GNP) for the year was $629 billion, an increase of 6.8 percent from 1963. Allowing for price changes based on 1958 dollars, the gain in real terms was 5 percent. In 1964 the Federal Reserve Board Index of Industrial Production stood at 132, a rise of 6.2 percent. With the impetus provided by personal income and corporate tax reductions during the year, aggregate wages, profits, and consumption reached record levels. Even the persistent balance of payments problem improved. Against this background the national consumption of commercial energy increased 4.2 percent. Energy needs continued to be met largely from domestic resources with imports, mainly crude petroleum and refined petroleum products, accounting for but 7 percent of the total. Mineral energy resources, namely gas, coal, and petroleum, contributed 96 percent of the input of energy into the economy during the year. The remaining 4 percent came from hydropower and nuclear power.

Jan 1, 1965

By Budiselic C, Hefter G

The uptake of gold by activated carbon from acid chloride solutions involves the reduction of gold(IIl) to metallic gold. Deposition occurs only on the surface of the carbon at localized active sites at first, and subsequently at less active areas. The rate of uptake depends on the surface area of activated carbon, which confirms results from a recently published study (de Siegel and Soto, 1984). Increased rates of uptake were observed at higher temperatures and lower chloride concentrations due to the decreased stability of AuCI4-, the major gold(IIl) species present. Recoveries of gold approached 100 per cent for a contact time of 60 minutes at 20°C in a well stirred solution initially 5 ppm in gold(III) and containing 0.5 per cent w/v coconut-shell carbon. The presence of 0.25 Mnitric acid or 100 ppm iron(IIl) had no detectable effect on either rate or efficiency of gold uptake.

Jan 1, 1985

By Charles Knilans, Tassos Grammatikopoulos, Nicole Hoffmann

"Twin Metals Minnesota LLC, (TMM) has recently completed a prefeasibility study evaluating the development and operation of an underground copper (Cu), nickel (Ni), platinum (Pt), palladium (Pd), and gold (Au) mining project on the Maturi deposit in northeastern Minnesota.The Maturi deposit is a magmatic Cu-Ni-sulphide deposit occurring along the basal contact of the South Kawishiwi intrusion, comprised of heterogeneous troctolites to gabbros forming a tabular body designated the BMZ (basal mineralized zone). These rocks provide the major gangue mineralogy (plagioclase, olivine, and clinopyroxene) with local, minor rock-forming orthopyroxene, biotite, magnetite, and ilmenite. Altered silicates exist within local fracture and/or alteration zones, which are commonly associated with minor faults in the regional area. The Maturi deposit is comprised of four main mineralized zones within the BMZ, referred to as Stage 1, Stage 2 (S2), Stage 3 (S3), and Upper Heterogeneous. Chalcopyrite, cubanite, pentlandite, and pyrrhotite are mainly hosted within the S2 and S3 units, with talnakhite frequently occurring within the S3 unit.A cross-discipline approach combining geology, geochemistry, metallurgy, QEMSCAN™ mineralogy, and mine planning has been implemented to develop a preliminary geometallurgical model and life of mine metallurgical performance forecast. This paper provides an overview of the methods that were used to develop the model and forecast, and discusses the advantages of such a model in project development and evaluation.INTRODUCTIONThe TMM Project proposes mining a copper-nickel-palladium-platinum-gold-silver (Cu-Ni-PGM) magmatic sulphide deposit situated in the Duluth Complex of northeastern Minnesota, USA. The prefeasibility study proposes a 45,000 million tonnes per day mine and concentrator facility to produce two saleable concentrates. The mine will access two separate deposits, Maturi and Maturi Southwest. Saleable concentrates are the final products of the sequential flotation plant where a high grade Cu concentrate is produced followed by a Ni concentrate."

Jan 1, 2015

By E. G. Noble

The effects of HBr concentration and temperature on the solubility and hydration state of AlBr3,' CoBr2, PbBr2, MnBr2, KBr, and NaBr were investigated by the Bureau of Mines. Saturated aqueous solutions of the AI, Pb, K, and Na bromides were sparged with HBr gas at 20°, 40°, and 60° C, and saturated solutions of Co and Mn bromides were sparged at 20° C. Increased HBr concentration caused decreased salt solubility because of the common ion effect for all the investigated bromides except PbBr2. Aqueous solubilities at 20° C were, in percent, 51.7 for AlBr3, 53.2 for CoBr2, 1.1 for PbBr2, 59.6 for MnBr2' 39.5 for KBr, and 47.4 for NaBr. At 60° C, the solubilities were, in percent, 53.3 for AlBr3, 2.2 for PbBr2' 46.0 for KBr, and 53.8 for NaBr. Solubilities in solutions sparged to HBr saturation at 20° C were, in percent, 0.5 for AlBr3, 42.0 for CoBr2, 32.2 for PbBr2, 21.0 for MnBr2, 0.6 for KBr, and 0.3 for NaBr. At 60° C, the solubilities at HBr saturation were, in percent, 1.8 for AlBr3, 41.4 for PbBr2, 1.3 for KBr, and 0.4 for NaBr. CoBr2 precipitated out to a minimum solubility of 27.2 pct at 28.7-pct-HBr concentration. Further increases in HBr concentration increased CoBr2 solubility because of the formation of bromide complexes. PbBr2 in-creased in solubility as HBr concentration increased.

Jan 1, 1988

By R. V. Lundquist

This report describes results of tests on the desilication of solutions resulting from the leaching of lime-soda sinters. Desilication was accomplished by heating the solutions in an autoclave. A liquor relatively free of silica was obtained after 30 minutes of heating at 200 p.s.i.g. in the presence of lime. The addition of lime improved desilication, but the rate and efficiency of desilication was effected by the quality of lime used. The optimum quantity of lime for effective desilication was about 6 grams per liter. A lesser quantity gave poorer desilication, and a greater quantity increased the alumina loss without improving desilication. The alumina loss was approximately proportional to the lime added. A complex structure, called artificial noselite in this report, was the only silica bearing compound detected in the desilication precipitate. Other compounds present were Ca(OH)2 (calcium hydroxide), CaCO3 (calcium carbonate), and Ca [A1(OH)6]2 (hydrated tricalcium aluminate). Sodium carbonate in the leach liquor retarded the formation of the calcium aluminate; hence it decreased the alumina lost to the precipitate. The extraction of alumina from aluminum silicate materials by the limestone soda-sinter process has been investigated by the Bureau of Mines for many years. The process consists of the following major operations: Blending the aluminum silicate with limestone and soda ash in such proportions that dicalcium silicate and sodium aluminate are formed on sintering; leaching the sinter in a carbonated caustic solution to extract the soluble alumina; purifying the leach liquor by the precipitation of soluble silica (this precipitate is designated desilication precipitate in this paper); precipitating alumina trihydrate from the purified liquors by carbonating with carbon dioxide gas; and calcining the trihydrate to alumina. Commercial grades of

Jan 1, 1962

By Oliver Bowles

The United States has never attained importance as a producer of asbestos, the domestic output being only 1 to 3 percent of the quantity required to supply raw material for its extensive asbestos-product manu- facturing industries. Production is centered in four principal countries-- Canada, Rhodesia, Union of South Africa, and the Soviet Union. Since asbestos first became an important commercial commodity the Province of Quebec, Canada, has led in tonnage produced, but a large proportion of the output of this province is shorter, nonspinning grades. Since 1916 the Union of South Africa and Rhodesia have become increasingly important as sources of the higher-grade spinning fibers. Before the World War Russia ranked next to Canada as a producer of asbestos, but during the revolution- ary period the industry in that country was almost at a standstill. In 1926, however, activity was revived, and each year since then production and exports have increased greatly. Minor quantities of asbestos are produced in Cyprus; Australia, Finland, and several other countries. Chrysotile is the leading product in all the producing countries, but substantial quantities of both blue (crocidolite) and brown (amosite) asbestos also are produced in the Union of South Africa. Minor quantities of anthophyllite and similar amphibole varieties are mined in the United States, Italy, Bulgaria, Finland, and several other countries.

Jun 1, 1934

By C Hoffmann

Given that each year, 30–40 per cent of mining industry deaths are attributable to failures of vehicle interaction controls (EMESRT, 2023), the mining industry is now armed with a growing body of knowledge from EMESRT (Earth Moving Equipment Safety Round Table) and ICMM (International Council on Mining and Metals) global vehicle interaction initiatives over the past years, that enables mining companies, OEM’s and technology providers to work together to develop and implement more effective controls and drive the adoption of capable solutions. Leveraging 20 years of experience as a global technology provider of Proximity Detection System (PDS) and Collision Avoidance System (CAS) solutions for surface mining operations, Wabtec Digital Mine Technology shares key insights for ensuring the effective selection, implementation and integration of these technologies into existing mine site controls and mining operations for effectively reducing vehicle interaction risk enabling mines to drive towards ‘zero-harm’ objectives. Topics covered include: alignment with industry best practice for controlling vehicle interaction risk by EMESRT and ICMM, importance of the maturity of EMESRT levels 1–7 vehicle interaction (VI) control framework, managing expectations that PDS/CAS is a ‘silver bullet’, alignment with industry functional – performance – technical requirements, proximity detection versus collision avoidance, sensor performance and fusion and limitations, rules engine intelligence critical to mitigating nuisance alarms during normal operations, system health monitoring and reporting, human factors engineering and operator acceptance, implementation strategy, installation standards and quality control, change management program, training programs, reporting and data analytics with actionable insights and maintenance strategies. Insights will also be provided on the challenges moving from EMESRT levels 7 (operator awareness) and 8 (advisory controls) to level 9 (intervention controls) as it’s a journey not an out-of-box solution, the growing demand for coexistence of CAS and Autonomous Haulage Solutions (AHS), functional safety approach, value added systems integration and the applicability of new emerging technologies.

Apr 16, 2024

By Carlos G. Levi

"The principles involved in the rapid solidification of supercooled metal droplets and existing numerical models are reviewed. The limitations of these models for studying alloy solidification are discussed in terms of the effects of alloying elements on the relevant heat transfer, thermodynamic and kinetic parameters. Analysis of the enthalpy-temperature diagram reveals how the extent of rapid solidification is hindered by the addition of solute. A solidification model incorporating thermodynamic and kinetic relationships into a Newtonian heat transfer formulation is used to analyze the thermal history and concurrent solute redistribution during recalescence. IntroductionIt is well acknowledged that microstructural evolution during solidification is intimately associated with the thermal history of the liquid-solid interface. Structural parameters like chemical homogeneity, morphology and size scale of the phases can be related to the interfacial temperature and velocity. These in turn change by the interaction of the interface with the thermal and solutal fields in its neighboring environment.The thermal history of powder solidification can be conveniently described on the enthalpy-temperature diagram for a pure metal shown in Figure 1 and discussed in detail elsewhere [1). A system can conceivably follow any path of decreasing enthalpy in the supercooled region of Figure 1 (~ < 1), starting at the nucleation temperature (oN < 0) on the liquid line, and ending at some point on the solid line. A vertical path represents an isothermal transformation, of which a limiting case is solidification without significant barriers to nucleation and growth, with the interface at the equilibrium temperature TM (path one in Figure 1). A horizontal path represents isenthalpic (or adiabatic) solidification whereby the system could form an amount of solid proportional to the initial supercooling of the liquid without further heat loss to the environment. A liquid that can completely transform to solid under adiabatic conditions is said to be ""hypercooled""."

Jan 1, 1986

By F. Montgomery Lyon

Pyritic sulfur in coal occurs as discrete mineral particles, and its liberation and physical rejection is the primary target for coal desulfurization. An acceptable solution to the pyrite removal problem from fine coal is not found in the standard methods of coal cleaning such as jigging, heavy-medium separation, tabling, and hydro-cycloning. Although fine coal is an ideal size for froth flotation, this method, as currently practiced is also unsuitable as the fine pyrite usually floats with the coal. One possible solution to this problem may be the froth flotation of pyrite from coal. This reverse flotation scheme would have the major advantage of removing both free pyrite and locked pyrite-coal combinations. With this possibility in mind, we have, as a first step, compared the flotation and adsorption. characteristics of an ore-pyrite and coal-pyrite with potassium-ethyl-xanthate as a collector 1 Related Research. The action of ethyl-xanthate on coal has been studied 2 by Horsley. He measured the contact angles of coal in distilled water and in the presence of purified ethyl-xanthate. The results indicated that the xanthate was not adsorbed by the coal. Additional contact angle data were obtained to determine the effects of cupric, ferric, and ferrous salts on the system. Then either the ferric or cupric salts were in, solution with the xanthate, there was an increase in the natural contact angles of the coal. This indicated that these heavy metal ions activated the flotation of the coal with xanthate. However, ferrous iron did not act as an activator. Horsley stated that the most attractive explanation was that the xanthate is oxidized to dixanthogen in the presence of the ferric and cupric ions.

Jan 1, 1971

By D. J. Adams

Cyanide heap leaching is the predominant technology used in processing low-grade gold ores. During closure ora heap leach operation, residual cyanide must be removed from the process and waste solutions, as well as the heap. Biological cyanide oxidation is a proven, economical technology for destroying cyanide in process and waste waters and spent heaps; The use of biological cyanide degradation eliminates the need for toxic or corrosive chemical oxidizers and has- been implemented at full scale at treatment costs usually <$0.50/1,000 gal. The Applied Biosciences biological cyanide destruction (BCND?) process has been demonstrated to effectively treat cyanide, concentrations up to 350 mg/L. Treated process solutions and wastewaters also contained other contaminants such as arsenic, copper, iron, silver. selenium, mercury, nitrate and zinc, much of which was removed in the cyanide degradation process. Under optimal conditions, microorganisms can rapidly oxidize free cyanide in some process solutions from over 250 mg/L down to 0.1 mg/L in 4 to 5 hr. Cyanide is degraded to ammonia and carbon dioxide, with 50%ofthe cyanide carbon liberated as C02. In general, carbon tank based bioreactors degrade cyanide at significantly higher rates than process or wastewater pond configured treatments, which usually degrade cyanide-more rapidly than in heap treatments. Investigation of cell-free enzyme preparations as an alternative to live microbial cyanide degradation shows promise. Advantages of enzymes over live microbes for cyanide degradation include: (I) the ability to tolerate and degrade higher cyanide concentrations (> I ;000 mg/L), (2) nutrients to support live microbial cells are not required, (3) the effects of other toxic contaminants, such as metals, found in mining waters-are eliminated and (4) the potential for greatly increased kinetics.

Jan 1, 1998

By W. A. Ver Wiebe

Under the impetus of new demands caused by the war, the oil and gas industries of Kansas established new records during the year 1942. In all, 1513 test wells were drilled, which is somewhat of a drop from the figure of 2113 for 1941. The 1513 tests yielded 800 new oil wells, 85 new gas wells, and 628 dry holes. Thus there was an increase in two categories, the gas wells and the dry holes. Considering that last year only 592 dry holes were completed out of a much larger total, the figure of 628 dry holes causes some concern among local operators. The increase in unsuccessful tests illustrates once more the fact that each year it is becoming harder to find the remaining oil pools. When the figures are broken down, it appears that western Kansas accounted for 1177 tests, of which 689 were oil wells, 33 gas wells, and 455 dry holes. The total potential daily initial production uncovered by the successful wells amounted to 753,178 bbl. of oil and 619,734,000 cu. it. of gas. The figures for new initial production indicate a substantial drop from last year&apos;s figure of 1,691,000 bbl., and the figures on gas are slightly higher. In 1941 the new gas as based on a daily potential was 508,000,-000 cu. ft. In eastern Kansas, 336 wells were drilled. Among these in were oil wells, accounting for a daily initial potential production of 16,199 bbl. The new gas wells numbered 52 and added 253,725,000 cu. ft. per day to the gas potential production of the state. The remaining tests, 173 in number, were dry. It is interesting and perhaps significant that about half (630) of the wells drilled in western Kansas were drilled by the independent operators. Among the large companies, the Cities Service Oil Co. drilled 163 wells; the Stanolind Oil and Gas Co., 52; the Skelly Oil Co., 39; the Continental Oil Co., 48; and the Shell Petroleum Co., 21 wells. In eastern Kansas the independent operators drilled 297 out of the total of 336 wells. Among the major oil companies the Cities Service Oil Co. drilled 19; the Phillips Petroleum Co. 7; and the Sinclair-Prairie Oil Co., 6 wells. The figures based on relative percentages of dry holes to successful oil or gas wells are revealing. In western Kansas almost one half of all wells drilled by the independent operators were dry holes. The major oil companies drilled approximately 547 wells in western Kansas, of which 379 were successful in finding either oil or gas. It thus appears that the major companies fared better than the independents. In eastern Kansas the major companies drilled 39 wells, of which 29 were successful in finding commercial quantities of oil or of gas. No less than 329 wildcat test holes were put down in western Kansas and 105 in eastern Kansas, or 434 in all, during 1942. Considerably more than half of these were drilled by independent operators; 223 in western and 99 in eastern Kansas. An exploratory hole 56 mile or more from production is regarded as a wildcat well in Kansas. On a percentage basis, about 80 per cent of wildcat wells were failures. In western Kansas only 20 per cent found either oil or gas. The failures were about equally distributed among the major

Jan 1, 1943