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By L. J. Trilli

RECENT studies have indicated that there may be considerable difference in the cleanliness of the central pipe cavity of open top ingots. It has been speculated that some of this difference arises becausk of reactions that occur during reheating in the blooming mill soaking pits. In this respect, it has been recognized that an oxide coating or flow of slaggy material in the cavity could inhibit welding during hot rolling and thus reduce the amount of sound product which may be obtained. This note is intended to summarize some observations on the structure of a split ingot top which may be useful in describing a mechanism whereby the shrinkage cavity could be oxidized during reheating. A 221/2x241/2 in., big-end-down, test ingot was secured from a routine production heat. The chemical analysis representing this ingot is as follows: 0.08 pct C, 1.04 pct Mn, 0.062 pct P, 0.315 pct S, and 0.02 pct Si. The deoxidation practice used on the heat from which the experimental ingot was

Jan 1, 1957

By W. O. Borcherdt

THE 650-ton underground mill of The Empire Zinc Company of Colorado (a subsidiary of The New Jersey Zinc Co.) serves the Eagle mine at Gilman in the Battle Mountain mining district of Colorado, The topography is rough, the mine and mill being under Battle Mountain and the adit entrance in the north wall of the canyon of the Eagle River, which is here a steep gorge. The town of Gilman stands on a shoulder of the mountain at altitude 8975 ft., above sea level, along the Pike's Peak ocean to ocean highway and contains the company office and the usual facilities of a modern mining town. The railway contact is at Belden, altitude 8370 ft., at the bottom of the gorge directly below Gilman, where the westbound track of the Denver and Rio Grande Western Railroad is cut along the north bank and the eastbound track along the south bank of the river. A general description of the mine, mill and service facilities, with an account of the geology, character of the ore and methods of mining and milling having previously been published in a series of articles1, this paper will discuss with as little repetition as practicable the reasons for building the mill underground rather than as a conventional structure, the methods of construction employed and the experiences gained in what is apparently the pioneer plant of this type and extent. There is a large historical background for underground plant construction. From the dim past caves have been used for dwellings, stables, storage and shops. It was natural that where caves were available their many advantages led to their enlargement to suit convenience and where they were not available but the ground conditions, topography and drainage were favorable, to the construction of artificial caves. In the vicinity of Epernay, France, extensive underground chambers protect the manufacture and storage of champagne-an intricate and ingenious application of chemical engineering-and in recent years these plants have been mechanized by the installation of conveyor systems. In the

Jan 1, 1937

By M. E. Austin

Many numerical models for the tubular solid oxide fuel cell (SOFC) exist, with various degrees of sophistication. However, the question of what physics must be modelled, and what can be neglected has not been resolved. A parametric model is proposed and analyzed to quantify the influence of thermal phenomena such as reaction heat generation, cathode channel convective cooling, and thermal radiation with participating media. Currently, a three-dimensional model is used to simulate the conditions of an anode supported cell in the centre of a tube bundle. The interconnect and nickel foam spacers that bridge adjacent cells in series and parallel, respectively, are neglected thus enabling a domain consisting of a 45o azimuthal section of the tube. The conservation equations for mass, energy, and species transport are solved using commercial computational fluid dynamics software, and radiation heat transfer is also modelled using a Discrete Ordinates method. Finally, temperature-dependent material properties and reaction rates are used. The model predicts lower overall temperatures than models that do not include radiative heat exchange and more moderate temperature gradients than those that ignore participating media radiation. Further, different regimes are identified, from convection dominated heat transfer to radiation dominated heat transfer. The results can be used to aid in the selection of appropriate models for a given fuel cell operating regime.

Jan 1, 2005

In presenting this list of permissible explosives and blasting devices, the Bureau of Mines desires to call attention to pages 19 and 24 outlining, the conditions in use under which, and only under which, the explosive or blasting device is permissible. In view of a recent publication recounting trade practice favoring the use of a blasting device without stemming, particular emphasis is placed upon condition 4 on page 19 and condition 2 on page 24. An explosive or a blasting device enumerated in the permissible lists of this Bureau is not permissible in use unless it is properly confined with clay or other incombustible stemming.

Jan 1, 1940

In presenting this list of permissible explosives and blasting devices, the Bureau of Mines desires to call attention to pages 19 and 24 outlining, the conditions in use under which, and only under which, the explosive or blasting device is permissible. In view of a recent publication recounting trade practice favoring the use of a blasting device without stemming, particular emphasis is placed upon condition 4 on page 19 and condition 2 on page 24. An explosive or a blasting device enumerated in the permissible lists of this Bureau is not permissible in use unless it is properly confined with clay or other incombustible stemming.

Jan 1, 1940

By C. D. Anderson

A variety of processing technologies exist for recovery from both primary and secondary sources of rhenium. Currently, there are no known primary rhenium deposits; thus, the method in which primary rhenium is produced is dependent on the commodity of which it is a byproduct, e.g., copper, molybdenum, uranium, etc. In addition, focus on the recovery of rhenium from secondary sources, such as alloy scraps and catalysts, is continually growing. This paper presents a review of both primary and secondary processing technologies for the recovery of rhenium.

By Carolina Rouillon

El presente documento señala dos experiencias aplicadas en la mina Yanacocha y Pierina exitosas donde se emplea las geomallas como parte del mejoramiento de suelos blandos con la finalidad de distribuir uniformemente las cargas impartidas por los grandes equipos mineros hacia la subrasante de mina típicamente conformada por suelos con alto contenidos de limos y arcillas húmedas y saturadas. Asimismo, los beneficios obtenidos por usar las mallas, no se limitan al ahorro en costo material de relleno, sino que se logra un mejor desempeño- en tanto optimiza la superficie de circulación de los vehículos mineros, reduciendo los tiempos de operación, así como prolonga los períodos de reparación y mantenimiento.

Sep 21, 2013

By Finn Ouchterlony

"Recently detailed fragmentation data from blasts fired with electronic delay detonators (EDDs) in the Red Dog mine in Alaska were presented. They include a comparison with pyrotechnic detonators fordifferent in-­-row delays in a standard pattern and EDDs in an expanded pattern. Fragmentation was measured on the belt after the primary crusher. The Rosin-­-Rammler function (RR, two parameters) was used to describe the crusher product. This was combined with the JK SimMet crusher model and the CZM blast model (Crushed Zone Model, three to five parameters) to estimate a reasonable crusher feed (muck pile fragmentation or run of mine, ROM) for each fitted product. In this way the effect of delay timing and electronic detonators on the median fragment size x50 and uniformity coefficient n for both crusher product and the blasted rock was determined. "

Jan 1, 2012

By R. Gallinger, B. Kelley

"In 1997, a group of Canadian mining companies met to review the Mine Environment Neutral Drainage (MEND) program and to consider future activities. They concluded that (1) that Canadian industry and Natural Resources Canada (NRCan) should fund a small secretariat in Ottawa to support technology transfer in Canada and, (2) that it was important for industry to continue the momentum of MEND on an international scale.The need for the International Network for Acid Prevention (INAP) was strongly endorsed by consultants, company, university, and government experts from Australia, Canada, and the United States who convened in Vancouver in February 1998. The initial focus was the development of a solid knowledge base covering the collective experience of INAP’s members to promote the dissemination of successful practices, the identification of knowledge gaps, and the formulation of effective research proposals. The sharing of information on past successes and failures alone was viewed assufficient justification for the formation of INAP.INAP was officially launched in October 1998, and is now an effective, industry-based initiative to globally coordinate research and development in the management of sulphide mine waste. Significant improvements in the management of sulphidic mine materials through knowledge sharing, research, and development of technology by INAP will reduce liability associated with acid drainage.This paper will explore the international efforts and activities of INAP."

Jan 1, 2005

By Ann G. Kim

The Bureau of Mines investigated factors that influence the formation of methane and other hydrocarbon gases in coal, which is an integral part of the coalification process. Coalification begins with a biogenetic stage in which plant debris decomposes and peat is formed; methane and CO2 are the predominant gases produced during this stage. During the diagenetic stage of coalification hydrogen and the hydrocarbon gases methane, ethane, ethylene, propane, propylene, butene, iso- and normal butane, and pentane are formed by dissociation of the organic material. Methane may also be produced by a secondary reaction of hydrogen with unsaturates or by the reduction of CO2. The gas contained in coalbeds is a mixture of methane, C2 through C5 alkanes, CO2, 02, N2, H2, and H2; its heating value is equivalent to that of natural gas. The gas content of coal depends on many variables, but an empirical relationship between adsorptive capacity, rank, and depth has been observed.

Jan 1, 1978

Talvivaara deposits, Kuusilampi and Kolmisoppi, comprise one of the largest sulfide nickel resources in the world with 642 Mt in Measured and Indicated Resource categories. TalvivaaraÆs preferred extraction technology is bioheapleaching, which is already widely used for other metals, notably copper and gold. The company has demonstrated the viability of bioheapleaching technology for the extraction of nickel using the Talvivaara ore. The leaching process is heat generating and therefore suitable for the subarctic climatic conditions. The planned annual metal production is 33 000 t of nickel, 60 000 t of zinc, 10 000 t of copper and 1200 t of cobalt. Talvivaara deposits are well-suited for open pit mining due to a thin overburden, favourable resource geometry and a low waste to ore ratio. The ore is relatively low-grade, but well suited to bioheapleaching due to its high sulfide content. Bioheapleaching is a process whereby metals are leached from ore as a result of bacterial action. The bacteria used in the Talvivaara bioheapleaching process are endemic to the area and therefore well adjusted to the prevailing environmental conditions. The solution is collected at the bottom of the heaps and either re-circulated through the heap or fed to metal recovery. In metal recovery, nickel, copper, zinc and cobalt are precipitated from the pregnant leach solution and filtered to produce saleable metal products. After the metals are removed, the solution is further purified and returned to irrigate the heaps. Mining, bioheapleaching and metal recovery techniques for the Talvivaara mine have been tested in a 17 000 t on-site demonstration trial during 2005 and 2006. Subsequent construction of the mine started in 2007 and commercial production in the fourth quarter of 2008.

Jan 1, 2009

By L. J. Barnes

The process for bio-remediation of sulphidic slag (BROSS) operates at 45 degrees C in a neutral (pH 7), aerobic aqueous environment in which sulphide in the slag is converted to sulphate. A two-stage version has been developed that involves quantitative chemical pre-oxidation of total sulphide to thiosulphate, followed by complete biooxidation of thiosulphate to sulphate. Ferric hydroxide, also produced, acts as a flocculant in the bioreactor effluent, concentrating heavy metal precipitates in the thickener as a sludge which can be returned to the smelter. The process has a high sulphate production rate and low acid and nutrient consumptions. Microorganism growth is the rate-limiting stage, requiring a minimum liquid residence time in the bioreactor of about 7 h. Although large quantities of oxygen are needed, sufficient gas mass transfer can be attained in an air-lift reactor. Products are an aqueous stream containing sodium sulphate with only trace levels of toxic heavy metals and a benign, toxic metal-containing sludge

Jun 18, 1905

By C. C. Hatch, C. R. Larsen, A. L. Mular

INTRODUCTION As part of a general project to simulate a concentrator, the Brenda Mines Limited secondary crushing plant, located near Peachland, B. C., was sampled during the summer/winter of 1975 and during the summer of 1978 to develop a digital crushing plant simulator specific to the Brenda operation. The purpose of this simulator was to study the influence of manipulable process variables on important responses, to permit study of alternative circuit configurations if desired and to perform off-line optimization. This chapter describes the methodology employed to develop the simulator (1,2) along with typical applications thereof. SECONDARY CRUSHING PLANT FLOWSHEET Figure 1 is a schematic diagram of the Brenda Mines Ltd. secondary crushing plant circuit. The 41,000 tonne coarse ore stockpile(a) is -17.8 cm oversize (3B belt) from two 2.4 m x 6.1 m double deck vibrating screens (not shown). The screens are fed by 2 conveyor belts operated in series with two 198 cm apron feeders that handle discharge from a 152 cm x 226 cm primary gyratory crusher(520 KW). The primary crusher (not shown) treats up to 2725 tonne/hr of run of mine ore. Minus 1.9 cm undersize from the primary screens (#19 belt), referred to as the primary fines, joins secondary screen undersize (#18 belt) for conveyance to fine ore bins (not shown). Feed from the coarse ore stockpile is transferred by means of vibrating feeders and two 120 cm conveyors (#6 and #7) to transfer belts #8 and #9 (Figure 1). The belts feed two secondary 2.1 m standard heavy duty cone crushers (bl and b2). Each crusher is equipped with a modified coarse mantle liner and a standard coarse bowl liner. Mantle throw is 9.2 cm; nominal close side set is 3.2 cm. Crusher .countershafts (443 rpm) are driven by single 260 KW synchronous motors (710 rpm). Discharge from the secondary cone crushers joins discharge from four 2.1 m tertiary heavy duty shorthead cone crushers (fl, f2, f3 and f4). With the exception of the cavity (modified medium mantle liner and standard coarse bowl liner; close side set is nominal .84 cm) and motor (224 KW), these machines are similar to the secondary crushers. The combined discharge is transferred from a 183 cm conveyor to #14 belt which empties into a 1452 tonne secondary screen surge bin (c). Vibrating feeders beneath the bins feed five 2.4 m x 6.1 m double deck vibrating screens (dl, d2, d3, d4 and d5) sloped at 200 from the horizontal. The upper decks are punched plates with 2.7 cm x 7.6 cm slots. Lower decks are woven wire (1.27 cm diameter) sections with

Jan 1, 1982

By Warwick Bullen

A hypothetical economic model has been developed for diamond projects in the Canadian Arctic using data from recent exploration and mining projects. The model reflects economic conditions as it applies to the diamond mining industry in the Canadian Arctic at the present time. The economic model captures pre-production expenditures, production data, and post-production costs. Before- and after-tax cash flows are derived from the data along with the relevant economic criteria. From the corporate perspective, the model provides guidance as to the minimum size and/or in-situ value per tonne requirements of targets being sought in order to ensure economic success and probable exploration and development expenditures. From the government perspective, the model provides an indication of income and mining tax revenues that can be expected from a particular diamond project. The model shows, on average, diamond projects in the Canadian Arctic are expensive to explore for and develop, but are economically robust. They generate great economic return for mining companies and significant amounts of tax revenue for governments. The fact that diamonds can be transported directly from the mine site by air makes them particularly attractive targets for mining companies that operate in the Canadian Arctic.

Apr 1, 2005

By L. R. Verney

About 75% of the market economy countries's supply of cobalt is produced in Central Africa by Zafre and Zambia. Zafre alone accounts for more than 55% and has the resources and capacity to increase output significantly at any time so its dominant position is unlikely to be challenged in the foreseeable future. Other producers from domestic resources are Canada, Australia and South Africa; and Cuba, probably the most important producer outside of the Market Economy countries. Belgium, Norway, France, Japan and Finland produce cobalt from imported materials. While there are "cobalt ores", inMorocco, USA, Canada and the USSR, the mineralisation is complex and cobalt extraction is costly. In ZaPre and Zambia cobalt is a by-product of copper; in Canada, Australia, France, and Cuba it is a by-product of nickel; and in South Africa it is a by-product from the production of Platinum Group Metals. Thus the levels of cobalt production are inevitably related to the production of other primary metals. While some cobalt is produced by predominantly pyrometallurgical methods most of the world's cobalt is produced hydrometallurgically. The existence of large hydro-electric power plants and the mineralogy of the copper and cobalt concentrates encouraged research and development in Zafre of the recovery of copper and cobalt by electrowinning from sulfate solutions; much pioneering work was carried out, which lead to construction of the two largest copperlcobalt hydrometallurgical plants in the world, at Shituru and Luilu. Similar systems were adopted in Zambia. In Canada high pressure hydrometallurgical processes based on the sulfate system were developed and adopted in Australia, the Philippines, Finland and South Africa. Meanwhile developments in chloride hydrometallurgical extraction of nickel and cobalt lead to its use in the extraction of cobalt in Norway, France and Japan.

Jan 1, 1988

By Paul Richardson

This paper discusses the drilling techniques and equipment used by Loffland Brothers Company to drill the AEC's UC-3 large diameter shaft in Nevada. Included are pictures and illustrations of the drilling-mud circulating system, the surface and downhole drill-string and the special heavy-duty mast and hoisting equipment required to do the job. The purpose and method of lining the large diameter holes along with the technique and materials used to backfill the lining to hole annulus are discussed briefly. Areas requiring improvement in both equipment design and technology for large diameter shaft drilling are discussed briefly. Shafts as large as 140 inches in diameter to depths of 5,000 feet can be drilled with existing equipment and technology.

Jan 1, 1969

By T. P. Meloy

Economic constraints require that the optimum mineral processing circuit be chosen for a given ore and then the circuit be optimized. Meloy (1983) developed a general methodology for finding the best circuit for a given ore. This analysis, however, did not specify a way to optimize the best circuit once it has been found. To optimize a given circuit, it is necessary to find the optimum set points for each unit operation in the circuit. Developed in this paper is a simple, general equation which, when applied, defines the optimum separation set point for the maximum mineral grade. This general equation, Equation (4), applies not only to simple ores, but also to complex ores containing variable and locked particles. Currently, the optimum set point can be calculated only for heavy media and cyclone circuits, because they are the only unit operations for which the necessary data has been measured.

Jan 1, 1984

By Johnson T, Huynh D

The State Electricity Commission of Victoria is conducting a program of research and development into alternatives to the use of conventional boilers for the generation of electricity from brown coal. The objectives of the program are to reduce the cost of electricity generation and also to improve the thermal efficiency so as to reduce the emission of greenhouse gases and other pollutants. The program presently includes the following four technologies: Hydro-thermal Drying, Direct Coal Fired Gas Turbine, Integrated Gasification Combined Cycle, and Steam Fluidised Bed Drying. The Hydro-thermal Drying process produces a combustible coal-water-slurry fuel which is suitable for use in a coal fired gas turbine. The coal fired gas turbine has the potential to produce electricity at a lower cost than conventional power stations and when combined with a steam cycle, it will also use significantly less coal. The Integrated Gasification Combined Cycle process gasifies the coal, cleans the gas, and then bums it in a gas turbine/combined cycle. This process is being developed in Europe and the USA in order to meet their stringent emission regulations. It offers Victoria cleaner more efficient and cheaper electricity.

Jan 1, 1992

By Marco Olper

The CX-EW technology, developed by B.U.S Engitec, is improved introducing besides the battery paste electrowinning, the direct electrorefining of metallic lead (grids and poles) obtained from CX breakers. This new concept of electrorefming the antimonial lead scrap is aimed to the production of 99.99 lead and of slimes with high antimony content, avoiding gaseous emissions in melting lead scrap, and expensive and pollutant thermic refining operations. In this way, practically all lead content in the battery is available as 99.99 lead, hence with a wider market for the lead recyclers. This paper will describe the tests performed on a pilot plant scale with special reference to the relatively low energy consumption of the process.

Jan 1, 1993

By Thomas P. Forbath

THE sudden realization that known sulphur reserves amenable to mining by the Frasch hot water process are nearing exhaustion focused attention on widely scattered surface deposits throughout the world. These deposits are not necessarily of lower sulphur content than ores located underneath Louisiana or Texas salt domes which usually average about 30 pct sulphur disseminated in limestone matrix. Their near surface occurrence, however, renders exploitation by the Frasch process impossible. As is well known, the Frasch process depends on the presence of 500 to 1000 ft of overburden and cap rock above the sulphur deposits to permit melting underground sulphur in place by diffusing hot water under pressures of 200 to 600 psig in the formation and raising the molten sulphur to surface by air lift. This process renders possible the production of pure sulphur which is 99.5 pct pure without any subsequent treatment. Surface deposits contain sulphur in the same range of concentrations as the salt dome deposits, i.e., from 10 to 50 pct sulphur, associated with various gangue materials such as silica, limestone, and gypsum. The pirincipal distinction, then, does not lie in the percentage of sulphur contained in the ore, but in the geological nature of the deposit. A recent study' of the world sulphur supply situation estimated 1950 sulphur production in the free world countries at 5.6 million long tons, of which the United States produced 5.2 million tons, or 93 pct of the total. While America's domestic needs alone would have been covered by national production, about 1.4 million tons were exported during the same year. Despite all the steps which are being taken to restrict use of elemental sulphur and to force the fullest possible development of alternate sulphur sources here and abroad, the deficit in elemental sulphur production will probably increase with time. As a result of intensive prospecting for oil throughout the Gulf Coast area discovery of significant new salt domes is held unlikely. With the growing scarcity of sulphur and what appears to be an inevitable rise in price, recovery from deposits not amenable to Frasch-process mining assumes greater economic importance. Untapped Reserves The most important deposits in this category are located in Sicily, where elemental sulphur occurs in Miocene limestone and gypsum formation. Sulphur content of these ores ranges from 12 to 50 pct with an estimated average of 26 pct. Although quantitative estimate of these reserves is not available it is held that they exceed 50 million tons of sulphur. Similar deposits occur also on the mainland which contribute about one-third of Italy's total current annual production of 230,000 tons, but these are known to be nearing exhaustion. Significant surface deposits of volcanic origin are located in South America, Japan and western United States, silica being characteristic gangue con-stituent. The largest of these deposits are in South America. More than 100 extend over a zone 3000 miles long, paralleling the west coast of South America. 'Total sulphur content of these deposits has been estimated to be as high as 100 million tons. The main islands of Japan also possess at least 40 known volcanic sulphur deposits with probable reserves of 25 to 50 million tons.' Prospected reserves in western United States might amount to 2 million long tons, principal deposits being located in the northwestern part of Wyoming, southern Utah, and eastern California. Volcanic deposits of lesser importance are found around the Mediterranean, in Turkey and Greece, and in Africa, Egypt, Abyssinia, and Somaliland. Beneficiation Methods Different methods of beneficiation have been used in these various locations. In Italy the Calcarone kiln and Gill regenerative furnaces are used exclusively. Both utilize heat liberated by burning part of the sulphur in the ore to liquify or vaporize the remaining sulphur, which is recovered by solidification or condensation. The Calcarone kiln is of conical shape, generally 35 ft in diam at base and 18 ft high. A kiln of 25,000 cu ft capacity burns for about two months and yields about 200 tons of sulphur. The Gill furnace consists of a series of chambers with domed roofs. Sulphur is burned and melted in one chamber at a time and the hot combustion gases are used to preheat the ore charge in the subsequent cell. These furnaces operate on a cycle of 4 to 8 days. The recovery yield of both systems is about 65 pct. Sulphur losses amount to 25 pct through the combustion to sulphur dioxide; about 10 pct is retained in discarded calcines. Ores containing less than 20 pct are not considered suitable as furnace feed. These methods are not only wasteful because of the low recovery obtained, but represent a serious atmospheric pollution problem. Sulphur produced ranges from 96 to 99 pct purity and thus does not match Texas or Louisiana sulphur. Owing to the present shortage, sulphur in the Middle East sells

Jan 1, 1954