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By Douglas Batista Mazzinghy

O moinho vertical normalmente é dimensionado através da aplicação de um fator de eficiência sobre o Índice de Trabalho de Bond (BWI) ou sobre o consumo específico de energia (kWh/t) obtido através de teste em moinho de bolas de laboratório. Para confirmar o dimensionamento do moinho vertical são realizados testes em escala piloto com grandes volumes de amostras. O presente trabalho apresenta uma metodologia para simulação de moinho vertical através de testes de moagem em escala de laboratório utilizando um moinho de bolas convencional e amostras bitoladas em diferentes faixas granulométricas. A amostra de minério de ferro estudada foi utilizada em uma campanha de testes em escala piloto com moinho vertical. As simulações realizadas utilizando a metodologia de Austin apresentaram muito boa correlação com os resultados dos testes em escala piloto.

Aug 1, 2018

By Kiyoshi Shibata

A new assessment model has been proposed for metal production and its recycling system. The model consists of three processes, smelting of ore, cascading and upgrading of metal scrap. The present approach is to assess the total energy consumption of the system, by introducing two equations for describing the relationships between critical impurity content and the mass of production or scrap recycling. The total energy consumption in the system can be described as a function of recycling rate and performance of the upgrading process. The present model suggests that the impurity removal efficiency and yield in the upgrading process are essential to reduce the energy consumption in the systems. Especially, such points should be explicitly controlled in recycling of aluminum. The potential capability of this model was also discussed by obtaining several quantitative information required for the process design and the operational control of beneficial recycling of metals.

Jan 1, 1998

By H. M. Harris

The Bureau of Mines investigated chlorination of domestic ilmenites in a fluidized coke bed at temperatures of 950° to 1,150° C. Objectives were to develop an effective process for producing TiC14 suitable for TiO2 pigment and to reduce pollution by collecting and treating waste byproduct chlorides. The reactor for these tests was a 10.2-cm-diam quartz tube heated externally by electrical resistance heaters. It contained coke beds of from 35.6 to 82.5 cm in depth that were fluidized by 13 to 25 l/min of chlorine. Ilmenite concentrate powders were introduced at the bottom of the fluid-bed reactor, using the chlorine gas stream as a carrier. Test results showed that greater than 90 pct of the ilmenite reacted at optimum conditions and recovery of TiC14 ranged from 95 to 99 pct of stoichiometric. A temperature-controlled cyclone was used to remove high-boiling-point chlorides such as MnCl2, MgCl2, and FeCl2, and also collected fine unreacted ilmenite and coke dust. Impure TiC14 and FeC13 were separated by collecting the, iron chloride in a condenser con-trolled at 200° C. Titanium tetrachloride vapor was cleaned by passage-through a salt-column filter followed by condensation in water-cooled and refrigerated condensers. Liquid TiC14 was purified further by well-known industrial methods to meet specifications for pigment-grade chloride. Integrated tests were conducted in which chlorine, produced by dechlorination of FeC13 with oxygen, was recycled and used directly in a fluidized-bed chlorinator.

Jan 1, 1976

By E. K. Kleespies

The Bureau of Mines demonstrated a technique for electrowinning titanium from titanium nitride and titanium dioxide. Titanium-nickel-copper metal solutions containing 26 to 61 weight-percent titanium were prepared by electrolyzing titanium nitride in molten mixtures of potassium fluotitanate) and sodium chloride or barium chloride and depositing titanium on nickel-copper cathodes. A titanium-nickel alloy was prepared by electrolyzing titanium dioxide in an electrolyte of potassium fluotitanate) sodium fluoride) and calcium fluoride and depositing titanium on solid nickel cathodes. This alloy contained 21 weight-percent titanium and had a combined oxygen-carbon content of less than 0.1 weight-percent. A compartmented cell confined the oxide feed near the anode and inhibited oxygen contamination of the metal.

Jan 1, 1966

By Don Runge, John T. Zeni

Cuprex Limited is currently developing a small gold mining operation in Central Australia in the Tennant Creek field. In February 1984, the company undertook a feasibility study on the project and operations commenced in August 1984. Ore reserves are located between 230 and 280 metres (755 ft and 919 ft) below surface and ore extraction required the establishment of a shaft network for access, ore hoisting and ventilation. Zeni Drilling Company from West Virginia, U.S.A. was contracted to blind drill two shafts for the project. Blind drilling was selected for various reasons, including timing which related to cost efficiency. The blind drill shaft technique has proved successful in this application and with enthusiastic and innovative support from the mining industry, will prove an efficient and cost effective means of accessing and operating underground mining projects of all sizes in the future.

Jan 1, 1987

"Cast blasting is the primary means of overburden displacement in a surface coal mine. It serves two purposes; fragmenting the rock and throwing it directly to its spoil pile using explosive energy. Increased throw minimizes costs by reducing the amount of material that needs to be re-handled. Prior studies have shown that by only changing blast hole timing with the same blast design, fragmentation and throw alters. It is necessary in cast blasting to optimize both fragmentation and throw, since larger fragments will require more wasted energy to throw the same distance. In an operating mine in Georgia, an optimum inter-hole delay for fragmentation has been found by studying timings from 0ms to 45ms. Instantaneous timing between holes increased the throw by over 100 ft, but fragmentation was poor. Shock and detonation wave collision is a potential reason for this increased throw. This paper investigates this optimized inter-hole timing while altering top and bottom column primer time to potentially improve throw while maintaining optimum fragmentation. Timings studied are top initiation, bottom initiation, and top and bottom simultaneously. INTRODUCTION Explosives are utilized extensively by the mining industry in order to break rock, as either ore or overburden, for hauling, loading, and crushing. An effective production blast strives to achieve optimum rock fragmentation in order to increase the efficiency and reduce the costs of those downstream operations, while simultaneously meeting environmental regulations. The goal of blasting with respect to fragmentation is, according to Hettinger (2015), “to decrease the average particle size of the fragmented rock without overly increasing fines or, in other words, improve fragmentation.” Additionally, the distance the rock is transported by the explosive energy from the blast, also known as throw, can increase efficiency by reducing the amount of material that is re-handled by mine equipment. The primary situation in which this is desirable is the cast blasting of overburden across the narrow pit of a surface coal mine. A key variable in a blast design pattern that can be manipulated to optimize both fragmentation and throw is delay timing between detonations. The invention and increased utilization of electronic detonators has allowed for precise delay timings to be used between rows, holes, and even multiple primers within the same hole. These recent developments in the blasting community have also provided an opportunity to investigate whether or not the mechanism of shock or detonation wave collisions caused by short timing delays are responsible for any increased fragmentation or throw. This experiment attempts to demonstrate if an increase in throw can be achieved while maintaining optimum fragmentation. This study consisted of 3 full scale bench blasts at a granite quarry in Talbotton, GA. Each blast was characterized by the timing delay between the top and bottom primer in the blast column of each hole; one simultaneous blast with a 0 ms delay, a bottom-initiated blast with a 1 ms delay, and a top-initiated blast with a 1 ms delay. Fragmentation analysis of the resultant muckpiles was performed using a combination of the PortaMetrics handheld point-and-shoot rugged tablet from Motion Metrics and the software WipFrag from WipWare. Throw was measured by placing strategic distance markers on the floor of the bench and then taking overhead photographs of the muckpiles after each blast. High-speed video footage was captured of each blast using an MREL Blaster’s Ranger II camera."

Jan 1, 2017

By W. W. Adams, T. D. Lawrence

"The National Safety Competition covering the calendar year 1934 was the tenth yearly safety contest conducted by the United States Bureau of Mines. Participating in the competition were 334 mines and quarries in 38 States. Five mining or quarrying companies, each the leader in its group for having established the best safety record during the year, were awarded the national safety trophies ""Sentinels of Safety.""Since these annual contests were begun 10 years ago by the United States Bureau of Mines the scope of the competition has been broadened materially, both as to the number of plants participating and as to the number of man-hours of exposure to accident hazards. Moreover, marked progress in the prevention of accidents has been achieved, as is indicated by an almost constant downward course in toe accident-frequency rates for the contest group as a whole during the past 10 years. On the other hand, the accident-severity rates for the group showed an upward trend during the first part of the decade and a downward trend during the latter part of the period except in 1930 and 1934; in each of those years a disastrous explosion with abnormally heavy loss of life occurred at the property of one of the participating companies.Because of the abnormal nature of the major disaster in a bituminous-coal mine in 1930 and 1934 the accident-severity rates for the bituminous group for those years are not strictly comparable with the rates for other years. The rates for bituminous mines for 1934 in the present report have therefore been computed on two bases; first, to exclude the record of the company that suffered the disastrous explosion and, second, to include the record of this company. The first rate represents the normal experience of the bituminous coal companies unaffected by the disaster and is therefore more indicative of the real trend of the accident rates for this class of mines from year to year. The latter rate shows the actual experience of all bituminous mines in 1934. The report of the competition of 1930 likewise showed the rates for the bituminous group as computed both with and without the disaster that occurred that year.The National Safety Competition is conducted by the United States Bureau of Mines in the interest of safety in the mining and quarrying industries. This object is accomplished directly through the efforts of the competing companies and their employees to win first place in the competition and thereupon receive well-merited and national recognition for their achievement. Indirectly the purpose of the contest is accomplished through studies by the Bureau of Mines of the facts contained in the accident reports which the competing companies furnish to the Bureau to enable the Bureau to compute the accident rates that are the basis for determining the winners of the contest."

Jun 1, 1935

By P. D. Kondos, J. R. Goode

Given the global inventory of low grade and refractory gold deposits, processes that could turn these challenged deposits into mines are of considerable interest. Such gold deposits have low value and potentially high operating costs and so they pose special challenges for the mining engineer and metallurgist. Conventionally, refractory gold ores have been processed by whole ore pressure oxidation and roasting, or flotation followed by roasting, pressure oxidation or bio-oxidation of the concentrate. Halide-based processes have recently been developed that might be applied to concentrates. Low grade refractory gold ores have been processed using heap bio-oxidation followed by cyanide heap leaching or milling. Other processes are emerging. Barrick has completed an inventory of emerging technologies (1). A summary of the current status of these developments will be presented with the view that, with additional development, some of these technologies or a combination thereof may unlock the value of these deposits.

Jan 1, 2005

By R. M. Lamya

A considerable part of gold encapsulated in some refractory sulphides can be dissolved during sulphide matrix decomposition in non-complexing aqueous media. This has been concluded from laboratory results of gold extraction from pyritic and arsenopyritic concentrate containing 172.8ppm Au, into 10% hydrochloric, sulphuric and nitric acids. The degree of sulphides decomposition in the non-oxidative acids limits the gold extraction from this specific concentrate to about 10%. In the case of nitric acid, the gold dissolution follows the rapid decomposition of sulphides, reaching 32 to 42% of extraction at almost total decomposition of the sulphidic minerals. Accumulation of gold in this acidic and non-complexing solution to several ppm was not affected by the soluble products of the sulphides decomposition. An attempt to explain this gold behaviour is undertaken and some practical implications are discussed. In the investigated concentrate the "soluble gold" is mostly associated with arsenopyrite. This is consistent with the latest findings of other authors that gold disseminated in arsenopyrite can occur as cation Au3+. In a comparative test the gold extraction into highly concentrated chloride solution, in the same oxidative conditions, amounted to 92.34%.

Jan 1, 1995

By William Gleason

Despite a global slowdown in most mining sectors, Caterpillar, the world?s largest mining equipment maker continues to forge ahead with new, and improved offerings to the global mining industry. On Nov. 10, as part of its Loading and Hauling Mining Focus Session at its Tinaja Hills proving grounds south of Tucson, AZ, Caterpillar unveiled to customers and media its two latest mining products, the 994K wheel loader and 6020B hydraulic shovel. The Caterpillar 994K wheel loader builds on a nearly 25-year history of the 994 model but with one noticeable difference ? the 994K is bigger ? much bigger. Weighing in at 239,142 kg (527,218 lb) the 994 K is up to 20 percent larger and comes with up to 30 percent increase in payload capacity. This behemoth of a machine is currently being tested at Caterpillar?s Tucson Proving Grounds, Anglo Gold Ashanti?s Cripple Creek and Victor gold mine in Colorado and Taseko?s Gibraltor copper and molybdenum mine in British Columbia, Canada.

By Archer E., Wheeler

Producing copper companies were active during 1941 owing to the national defense program the United States and the requirements of the friendly belligerent nation. This activity extended to the American owned properties in South America, the major part of whose product was taken by the Metals Reserve Co., a U.S. Government agency. Anaconda Copper Mining Co. is building a new converter plant at Anaconda, with four units of the Peirce-Smith type. The addition is being built adjacent to the reverberatory plant so as to avoid the necessity of hauling matte and slag. It is expected to be completed this year.

Jan 1, 1942

By Shenghua Yin, Xun Chen, Rongfu Yan, Zepeng Yan, Wei Chen

The application of cemented paste backfill (CPB) with alkali-activated slag and phosphogypsum (ASP) can contribute to the sustainable development of the mining industry. The effects of PG content, NaOH content, solid content, and waste rock-rod milling sand (W/R) ratio on workability, unconfined compressive strength (UCS), and setting time of ASP-CPB were investigated in this paper. The results showed that the workability of ASP-CPB decreased with PG content and NaOH content. Compared to OPC-CPB, ASP-CPB exhibited better workability and higher long-term (7 days and 28 days) strength at predetermined backfill formulations and curing times. The mid-term strength of ASP-CPB first increased and then decreased with PG content, while the short-term strength (3 days) decreased gradually due to the influence of soluble acidic impurities. Excessive NaOH will harm the strength of ASP-CPB, possibly due to the swelling effect of the reaction products (ettringite and sulfate). Strength and setting time increased and decreased with solid contents, respectively. Moreover, the compactness of the mixture has an important effect on strength and workability. The workability and strength of ASP-CPB reached the minimum and maximum values at a W/R ratio of 5:5, respectively. Pollutant index detection indicated that most of the pollutants in ASP-CPB leachate met the discharge standard, which proved that ASP-CPB is an effective method to recover PG. The results of the research will provide theoretical guidance for the efficient recovery of PG resources and support the achievement of safe, clean, and efficient mining.

Nov 17, 2023

By M. Eraydin

Dewatering fine particles (e.g., flotation concentrates) is inefficient, energy intensive and often requires large capital and operating costs. To address these issues, a series of advanced dewatering technologies have been developed at Virginia Tech. These include novel dewatering aids and a hyperbaric centrifuge. Typically, the moistures obtained using these methods are 30% to 50% lower than those obtained using conventional methods, as demonstrated in laboratory and pilot-scale test programs. With another novel dewatering technology that has been tested in bench-scale tests, the moistures can be reduced to levels once believed to be attainable only through the use of thermal dryers. In this communication, test results obtained with these advanced dewatering technologies are presented and compared.

Jan 1, 2014

By Lloyd A. Morley

A concept is presented to improve underground coal mine electrical system safety and availability. This proposed technique is based upon the ability to predict incipient failures in the mine power system. Prediction is made possible by a new method, developed during this research, which uses a minicomputer to implement pattern recognition algorithms. The theoretical basis of the prediction system is discussed, along with synthesis of the hardware and software sections. The deterioration modes which could be detected are examined, and an evaluation of the system's ability to predict incipient failure is presented. Typical results of the system's performance during laboratory simulation of power system deterioration are given.

Jan 1, 1977

By g. Caproni

The main constituent of the feed to the "Sulphide Section" (31 t/h), in the F. Sartori Plant (6000 tid) of the Miniere di Mon)eponi(Sardinia), is the ore of the Campo Pisano -Funtanaperda Mine (1 ? Difficulties have always been experienced in the treatment of this ore, owing to the extramely fine Pyrite-Sphalerite intergrowth and to the high Pyrite and soluble salts contents. The first mentioned difficulty was overcome with a rather complex flowsheet (see fig.1A and lB)(2); the others with a reagent combination that required a pH of 11 in a pulp loaded with soluble salts. This meant high consumption of alkalizer, and also of the other reagents; furthermore the alkalizer had to be mainly NaOH with only minor amounts of lime, because the latter tended to depress not only the pyrite but also the middlings, which were supposed to float. With dialchylditiophosphate alone as a collector, a lower pH could have been used, but a certain amount of amylxanthate was indispensable to promote flotation of the middlings. On the other hand, a bulk-flotation of sphalerite and pyrite at low pH could not be very advantageous on account of the high pyrite content in the feed(3).

Jan 1, 1971

By Christoph Zschiesche

Managing complex material streams within a lead smelting/reduction process has a long history at Aurubis. After modernization of the secondary lead smelter in 1991, Aurubis operates an electric furnace and a Peirce Smith converter to process complex secondary materials. This process links the copper and the lead metallurgy allowing Aurubis to properly combine primary and secondary process lines to optimize the processing of complex raw materials and smelter intermediates. However, combination of copper and lead metallurgy processing requires a continuous evaluation as raw material quality and composition varies. This paper describes Aurubis lead smelter processing and provides some fundamental considerations to optimize the lead processing, in particular with aspects related to speiss formation

By Yoichi Takeda

Copper solubility in SiO2-CaO-FeOx slag equilibrated with copper matte in 10% SO2 atmosphere at 1573K was estimated. Essential parameters are activity coefficients of copper oxide and sulfide, oxygen and sulfur potentials and activity of copper in the slag-matte system. A large number of experiments provide proper parameters to calculate copper solubility in slag, A classical thcrmodynamic model postulating CuO0.5 and CuS0.5 moleculcs assists in estimating copper solubility in slag. Additions of Si02 and CaO reduce copper solubility almost in any condition. Copper loss, that is function copper solubility and slag volume, in the silica saturated SiO2-FeO. binary slag is minimally equilibrated with less than 66% copper matte while copper loss in the 50% FeO.-SiO2-CaO ternary slag saturated with 2CaOSiO2 is minimally equilibrated with over 66% copper matte. This slag composition has potential for continuous copper converting, direct copper production or high-grade matte production.

Jan 1, 2003

By C. H. Elder

Coal degasification boreholes drilled from the surface have been success-fully hydraulically stimulated to increase the flow of gas from the coal. This Bureau of Mines report describes the hydraulic stimulation procedure for a coalbed, the equipment required, and design criteria for the treatment. Three examples of stimulation treatment show from fivefold to twentyfold increases in gas production. The tests were conducted in the Pocahontas No. 3 coalbed, Buchanan County, Va., the Pittsburgh coalbed, Washington County, Pa., and the Mary Lee coalbed, Jefferson County, Ala. Two sites where the coal was hydraulically stimulated and exposed by subsequent mining show no adverse effect on mining operations through the treated zones. The induced vertical partings were confined to the coalbed and did not in any way affect either roof or floor rock.

Jan 1, 1975

By Robert Griesbach

One of the important decisions in the development of new mining concentrator projects is the electric power supply for the project. Key decision factors often include the production of power on site versus purchase of power from the grid, ownership of the power supply infrastructure and the choice of generation technology. In this sense, the paper introduce a methodology that will allow the evaluation of the various alternatives to supply electric power to remote mining and mineral processing sites in a consistent and equitable manner taking into account economic and technical considerations and providing the same degree of availability and reliability. The most employed evaluation method consists of carrying out trade-off studies between the various alternatives but care should be exercised so that the alternatives analyzed offer comparable operational results.

Sep 14, 2009

By Charles T. Baroch, Charles J. Baroch

In 20 years of commercial development, nuclear reactors have demonstrated that they are a safe, dependable, and economical source of power. Operating experience with commercial power reactors has firmly established nuclear power as a practicable source of energy. For both boiling-water (BWR) and pressurized-water (PWR) reactors, performances to date have generally met design predictions. Furthermore, it has become possible to analyze and understand the nuclear, chemical, and physical behavior of this class of reactors. The United Kingdom has had similar success with gas-cooled reactors. The Experimental Boiling Water Reactor (EBWR), the first United States experimental power reactor, began operation in 1956. It proved to be capable of supplying 4500 kw (kilowatts) of power. This reactor was followed in 1957 by the Pressurized Water Reactor at Shippingport, Pa. It was the nation's first full-scale central-station nuclear power plant and the first in the world designed exclusively for the production of civilian electric power. The Shippingport reactor was originally designed to produce 60 megawatts (MW) of electricity, and it was increased to 100 MW in 1964. In 12 years of operation, this reactor generated over 4.5-million MW-hr of electricity in spite of many shutdowns for experimental alterations. By 1969, a total of 15 central-station nuclear power reactors were operable in the United States with an aggregate electrical capacity of over 4200 MW. Early in 1964, Jersey Central Power & Light Co. announced plans for constructing a 515-MW nuclear plant at Oyster Creek, N.J., to be built without government financial assistance. The company stated that a nuclear plant had been selected competitively with fossil fuel. Their economic analysis stated that the plant would generate electricity at less than 4 mills per kw-hr, competitive with coal at 200 per million Btu (British thermal units). Coal cost in 1964 for that region averaged 26.00 per million Btu. The Jersey Central report created wide interest in public-utility circles; nevertheless, no other nuclear plants were announced in 1964. However, in 1965, six nuclear plants were announced by utilities, followed by 26 in 1966, 29 in 1967, and 10 in 1968. By Sept. 30, 1969, 97 reactors were operable, being built, or planned. The cost of these, all committed by electric utilities, was estimated at about $12 billion. When constructed, they will have an electrical capacity of 72,000 MW. The unprecedented surge in power reactors that started in 1965 tapered in 1968, and only seven reactors were announced in

Jan 1, 1976