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By John W. Buch

At a meeting of the Anthracite Investigations Advisory Committee in July 1944 at Washington, D. C., the initial work and objectives of the Anthracite Mechanical Mining Section of the Federal Bureau of Mines were discussed. It was learned that interest existed in an improved performance of mining thin, steeply pitching beds above the gangway by mechanical methods. One company offered the use of one of its mines for experimental work, and this offer was accepted. Following acceptance, an effort was made by Bureau engineers to devise a mechanical method with a performance better than that of a new hand method proposed by company engineers for an actual mining condition. The method proposed by the Bureau was to drive 30-degree slant chutes across a 10-foot bed pitching 85 degrees. These chutes were to be advanced by using a lightweight, compressed air-powered, shearing machine of German origin that had been tried in a flat Pennsylvania anthracite bed in 1934, and which did not then prove successful because of blasting difficulties. An attempt was made to obtain one of these machines from England, where it was known machines of this type had been used. Shortly thereafter John W. Buch was sent to Europe as a member of the Solid Fuels Mission to the European Theater of Operations organized by the Technical Industrial Intelligence Committee of the Joint Chiefs of Staff. Through his efforts the Bureau of Mines obtained two such machines, known as the Eickhoff model DEK, one from England and one from Germany. This report covers the preliminary testing of these machines in order to obtain mechanical operating data for use in determining the power requirements for cutting anthracite with machines of this design.

Jan 1, 1949

By T. J. Williams

Researchers at the U.S. Bureau of Mines monitored rock mass response to mining of an experimental underhand longwall stope in Hecla Mining Co.'s LUCKY Friday Mine, Mullan, ID. This stope design, the Lucky Friday underhand longwall (LFUL), was proposed as a means of controlling rock bursting while also allowing increased mechanization of mining operations. Instruments were specifically manufactured to monitor three geomechanical factors (backfill performance, ramp stability, and rock mass response) that directly affect the success of stope design. Despite considerable difficulties with instrument installation and survival in the harsh mine environment, especially within the stope fill, sufficient information was collected to suggest that the LFUL method was successful in achieving project goals. This conclusion, in addition to favorable economic and operational evidence, led the mine to adopt the method throughout the mine. Indeed, this study and mining experience suggest that underhand longwall mining could be considered a feasible mining method for rock-burst-prone ground throughout the Coeur d'Alene District. Further Reports of Investigations (RI) in this series present in-depth analyses of the effects of backfill on rock bursting and calibration of a stope numerical model.

Jan 1, 2010

By G. G. Xiong

Water-sealed cyclone is a new efficient equipment to use for grading,concentrating of slurry, it was researched and developed by the Kunming Metallurgy Research Institute in the 1980s , and ac-cessed to the 1990 national invention award and the national patent products (ZL02245313.X ). The equipment and technology are innovative, firstly, we design a set of micro-pressure (0.005~0.008MPa) water-sealed device, which eliminates the gas phase of three-phase of sol-id-liquid-gas of the traditional general hydrocyclone grading process, it only has solid and liquid phases. Secondly, the axial velocity envelope surface which is a factor of controlling the accuracy of grading changes to a wedge rotation; and the interface of upward flow to down flow changes from surface to the solid, it can improve grading accuracy. Thirdly, water-sealed box and catheter devices make the stilling which pass through the cyclone cone centrifugal force fields re-enter the pressure grading zone of catheter. The three innovations make water-sealed cyclone can reduce the loss of overflow and has less mud in stilling product. Because the excellent performance, the water-sealed cyclone has application in non-ferrous metals, ferrous metals, non-metallic and aluminum hydroxide industries for grading, desliming and concentrating since it has successfully developed.

Jan 1, 2014

By R. D. Grove

The control of Florida phosphate flotation plants is largely based on the visual observations of experienced operators, both in recognizing good and bad froth conditions and in making judgments about the composition of feed, concentrate and tailing streams. Chemical analyses lag the process by at least two hours and therefore only assist in recalibrating the operators' vision when major deviations occur. An image analysis system has been developed by IMC Fertilizer to improve upon the accuracy and precision of the composition estimation process. It still relies upon the operator taking a representative sample, but provides a measurement in about 15 minutes after that sample is obtained. This paper describes the system and its use and presents comparative results for various process streams.

Jan 1, 1993

By Park Yun-Seok, Jung Seung-Won, Hwang Nam-Sun, Song Jin-hyuk, Kim Young-ho

This study proposes a new blasting vibration estimation formula that reflects the vibration reduction characteristics of electronic detonators, which are not considered in the current MOLIT blasting vibration estimation formula. Test blasts were conducted across six types of Korean open-pit blasting designs. The results of the test blasts showed that the K value for electronic detonators decreased by approximately 30% compared to electric detonators. Predicted Peak Particle Velocity (PPV) analysis indicated a reduction effect ranging from 30% to 31% across various types. By incorporating this reduction in the K value into the MOLIT blasting vibration estimation formula (K=200), a new electronic detonator blasting vibration estimation formula (K=140) was proposed. Using this new formula, the blasting damage range (R) was calculated under various vibration limits to evaluate constructability improvement. On average, the R value for electronic detonators decreased by about 17%, with significant reductions of around 41% observed for TYPE I under certain vibration limits. In conclusion, the adoption of electronic detonators in open-pit blasting operations improves blasting efficiency, reduces economic costs, and significantly enhances constructability. Electronic detonators have demonstrated superior performance in minimizing damage to safety facilities compared to traditional electric detonators, thereby proving their practicality and effectiveness in blasting operations.

Jan 21, 2025

By Jordan Mayorga, Joseph Kopp, Robert Boehm, Daniel P. Cook

"Solar generated power has long been considered a useful energy source only for low-energy density and non-mission critical use, such as providing electricity for residential applications. The success of Concentrating Solar Power (CSP) facilities, such as the 64 MW Nevada Solar- 1 near Las Vegas, have demonstrated that solar does have application to high energy density industries such as materials and chemical production. This paper will discuss the current operating con_guration of the Solar-1 facility and show how proposed modi_cations, such as addition of thermal storage and use of liquid salts as the working uid, can further improve the e_ciency and energy production of the plant.IntroductionThe United States as a nation is the largest consumer of energy in the world at an estimated 29 trillion kWh per year[1], which is roughly 15% of total global energy consumption[2]. On a per capita basis, the United States also leads most of the world, using an average of 96,000 kWh per person[3]. Roughly 40% of the energy consumed in the United States was produced from petroleum, 23% from coal, 23% from natural gas, 9% from nuclear and the remaining 5% from renewable sources such as solar, geothermal, wind, and hydroelectric[1].Energy demands associated with the industrial-scale production of chemicals and materials is a signicant portion of the United States total energy consumption. Roughly 33% of the energy consumed in the United States is used by the industrial sector, with 22% of this amount used in chemical production, 16% used in petroleum rening, and 14% used in the smelting and rening of metals[1]. Energy costs are also a signicant, if not the most signicant, portion of the total production costs for many materials. For example, roughly 30% of the cost of producing aluminum is the electricity used in the Hall-Heroult process [4], while the cost of energy is roughly 15% of the total cost of producing steel products [5].The primary sources of power for the industrial sector have been coal, petroleum, hydro- electric and nuclear. The use of alternative energy sources, e.g. solar, wind, geothermal, for industrial materials production has been slow in coming because they are intermittent and low-energy density sources, i.e. no solar power at night, no wind power when the air is still, etc. The success of Concentrating Solar Power (CSP) facilities, such as the 64 MW Nevada Solar-1 (NS-1) near Las Vegas, have demonstrated that solar can have application to high energy density industries such as materials and chemical production.This paper will discuss the current conguration of the NS-1 facility, and potential modications to the facility that will increase both its power e_ciency and power density."

Jan 1, 2010

By García B., Wolff E.

The pellet qualities primarily depend upon the thermal profile and gaseous environment during induration. During the induration of the pellet, it undergoes physical as well as chemical changes. Physical changes are mostly due to sintering and chemical changes are due to carbon burning inside the pellet. The effect of carbon on sintering is investigated on a single pellet scale using an optical dilatometer. Sintering is a complex process that involves transportation through various mechanisms. Sintering is quantified in terms of sintering ratio and the kinetic parameter, exponent ‘n’, under isothermal conditions is deduced using the power law equation. The value of n can be used to estimate the sintering mechanism at different isothermal temperatures. Also, the activation energy is calculated using the empirical Arrhenius equation. The activation energy of 0 per cent coke and 1 per cent coke-containing pellets are 531.0 kJ/mole and 552.6 kJ/mole respectively. The order of magnitude of activation energy suggests the same mechanism in both cases. Although the mechanism of sintering is the same, 1 per cent coke-containing pellet sinters more at any isothermal temperature.

Jun 17, 2024

By C. E. St. Denis

"Training Program at Sherritt Gordon's Ruttar:. ConcentratorA formal training program for mill operators is being developed and employed at Sherritt Gordon's Ruttan Concentrator. The program was set up to fulfil the needs created by a shortage of skilled reliable labour, a 200% turnover of personnel within the first six months of operation and the rapid advances in mineral processing technology. All new employees are introduced into the Mill Department through an orientation program designed to familiarize them with the operation of the plant. All training and orientation of new employees is the responsibility of a. small training section consisting of two training foremen and one assistant. Using a combination of classroom and supervised on the job training, the training times have been cut by at least one half. High operating standards have been set and are maintained through the use of detailed manuals for each work area and competent instruction at all operating levels. The major portion of training costs arise from the salaries of the training staff but are negligible when compared to the improved recoveries, smaller turnover and less downtime. Government assistance is available to all industrial employers to offset the costs of training employees."

Jan 1, 1974

By Yuriy V. Gulyaev, Galina V. Sedelnikova, Valeriy D. Lunin, Valentin A. Chanturiya, Igor J. Bunin

"Probable mechanisms of disintegration of mineral particles in gold-containing sulphides and benefication products exposed to high-power electromagnetic pulses are discussed. Experimental data are presented to confirm the formation of breakdown channels and selective disintegration of mineral complexes as a result of pulse irradiation, which makes for efficient access of lixiviant solutions to precious metal grains and enhanced precious metal recovery into lixivia during leaching. Experiments with different gold-containing products revealed a synergetic effect of high-power electromagnetic pulses with widely varying parameters and moisture in mineral pores on breaking-up and gold extraction processes for various resistant goldcontaining mineral products.Several series of tests performed in IPKON RAS, IRE RAS and TsNIGRI yielded an absolute gain in precious metal recovery, up to 30-80 % for gold and 20-47 % for silver, with resultant gold recovery as high as 90 %. An essentially new, environmentally safe and energy-saving breaking-up technology has been developed for resistant gold-containing products. This new method would shorten the way from raw minerals to final products and help significantly reduce the cost of these products as it obviates the necessity for using the expensive procedures of energy-intensive and environmentally hazardous oxidative roasting and autoclave breaking-up for ore concentrates."

Jan 1, 2003

By L. M. Pidgeon, P. G. Thornhill

A LTHOUGH a considerable number of experi--ti mental investigations dealing with the roasting of sulfide minerals have been reported in the past,'"" the behavior of the single roasting particle does not appear to have received the attention it perhaps deserves. This is understandable in that the natural sulfide counterpart of the slab or sheet of metal normally used in high temperature oxidation studies is difficult to obtain and prepare, and is, in most cases, virtually impossible to maintain in one piece at roasting temperatures. The employment, on the other hand, of static aggregations of small particles in masses great enough to permit evaluation of the roasting reactions by means of thermobalances or gas analyses introduces other complications, such as localized overheating, sintering, and variable gas-solid contact. In this investigation a compromise between the above extremes was attempted. Closely sized (e.g., 30 to 40 mesh) particles of natural sulfide minerals were isothermally roasted, out of contact with one another, in an air-swept system. The progress of the roasting reactions was followed by examination in polished section of the treated particles and by X-ray and chemical analysis of their residual sulfide kernels. In this way some conclusions respecting the mechanism of roasting reactions were drawn. Roasting Furnace—The vertical tube furnace used for roasting the sulfide particles had a heated zone 10 1/2 in. long in which four retractable light gage stainless steel trays were suspended. Each tray, and an accompanying baffle, was mounted on a 2 in. lengtn of 3/8 in. stainless steel tubing. These assemblies were axially suspended on a 30 in. length of ¼ in. stainless tubing, by means of which they could be lifted from the hot zone to a water-cooled brass chamber at the top of the furnace. Since each tray unit could slide freely on the 30 in. tube, any desired number of the trays could be retained in the cooling chamber by the manipulation of a horizontal plunger during a momentary withdrawal of the four trays from the hot zone. Although three heating coils, each equipped with variable external resistance, were employed, it was found impossible to avoid a drop of 10o to 15oC between the temperature at the top tray and that of the lower three. Consequently, only the bottom three trays were used for roasting. Temperature readings inside the furnace were made by means of a thermocouple inserted down the tray support tube at tray levels, and a separate thermocouple, inserted adjacent to the middle heating coil, was connected to a Micromax temperature control unit. Experimental Method—The sulfide minerals were obtained in a form as free as possible from gangue and contaminating sulfides, and were crushed and screened to the required particle size range. The tray assembly was kept inside the furnace during the heating-up period and, when a steady temperature had been reached, the trays were withdrawn. Each tray was sprinkled with from 200 to 500 mg of the sulfide grains, and the assembly was lowered into the furnace without delay in order to minimize the time required to achieve the operating temperature. The furnace top was then clamped into position and the thermocouple inserted down the center tube to the required level. It was found that, when this procedure was used, the trays achieved their former temperature within 2 min, after which time the flow of dry air was begun at a rate of 1 liter per min, and timing of the run was started. At the required intervals of time the trays were successively withdrawn from the hot zone and stored in the cooling chamber by means of the retention plunger. On completion of the run the treated particles were removed from the trays to be set aside for examination. The roasted particles were mounted in Lucite on a Buehler press Operating at a temperature of 140°C

Jan 1, 1958

By J. H. Wernick, D. Dorsi, K. E. Benson

A LTHOUGH a considerable number of experi--ti mental investigations dealing with the roasting of sulfide minerals have been reported in the past,'"" the behavior of the single roasting particle does not appear to have received the attention it perhaps deserves. This is understandable in that the natural sulfide counterpart of the slab or sheet of metal normally used in high temperature oxidation studies is difficult to obtain and prepare, and is, in most cases, virtually impossible to maintain in one piece at roasting temperatures. The employment, on the other hand, of static aggregations of small particles in masses great enough to permit evaluation of the roasting reactions by means of thermobalances or gas analyses introduces other complications, such as localized overheating, sintering, and variable gas-solid contact. In this investigation a compromise between the above extremes was attempted. Closely sized (e.g., 30 to 40 mesh) particles of natural sulfide minerals were isothermally roasted, out of contact with one another, in an air-swept system. The progress of the roasting reactions was followed by examination in polished section of the treated particles and by X-ray and chemical analysis of their residual sulfide kernels. In this way some conclusions respecting the mechanism of roasting reactions were drawn. Roasting Furnace—The vertical tube furnace used for roasting the sulfide particles had a heated zone 10 1/2 in. long in which four retractable light gage stainless steel trays were suspended. Each tray, and an accompanying baffle, was mounted on a 2 in. iength of 3/8 in. stainless steel tubing. These assemblies were axially suspended on a 30 in. length of in. stainless tubing, by means of which they could be lifted from the hot zone to a water-cooled brass chamber at the top of the furnace. Since each tray unit could slide freely on the 30 in. tube, any desired number of the trays could be retained in the cooling chamber by the manipulation of a horizontal plunger during a momentary withdrawal of the four trays from the hot zone. Although three heating coils, each equipped with variable external resistance, were employed, it was found impossible to avoid a drop of 10O to 15OC between the temperature at the top tray and that of the lower three. Consequently, only the bottom three trays were used for roasting. Temperature readings inside the furnace were made by means of a thermocouple inserted down the tray support tube at tray levels, and a separate thermocouple, inserted adjacent to the middle heating coil, was connected to a Micromax temperature control unit. Experimental Method—The sulfide minerals were obtained in a form as free as possible from gangue and contaminating sulfides, and were crushed and screened to the required particle size range. The tray assembly was kept inside the furnace during the heating-up period and, when a steady temperature had been reached, the trays were withdrawn. Each tray was sprinkled with from 200 to 500 mg of the sulfide grains, and the assembly was lowered into the furnace without delay in order to minimize the time required to achieve the operating temperature. The furnace top was then clamped into position and the thermocouple inserted down the center tube to the required level. It was found that, when this procedure was used, the trays achieved their former temperature within 2 min, after which time the flow of dry air was begun at a rate of 1 liter per min, and timing of the run was started. At the required intervals of time the trays were successively withdrawn from the hot zone and stored in the cooling chamber by means of the retention plunger. On completion of the run the treated particles were removed from the trays to be set aside for examination. The roasted particles were mounted in Lucite on a Buehler press operating at a temperature of 146°C

Jan 1, 1958

By C. J. Hunter

Prior to the commissioning of the Sansu Sulphide Treatment Plant (STP), Ashanti Goldfields' treatment facilities were capable of producing some 700 000 ounces of gold per annum. Ore sources were underground refractory sulphides, surface oxides and old tailings dams. In order to increase the mines production to 1 million ounces per annum and to accommodate the change in ore sources as surface mining moved into fresh sulphides, a part of the strategy adopted was -the construction of a new sulphide treatment plant with a capacity to mill 2.2 million tonnes per annum and produce 400 000 ounces of gold. The STP has been designed, built and commissioned. and has been in operation since February 1994. The Plant boasts the world's largest Bio-oxidation facility which is capable of treating 790 tonnes of sulphide concentrate per day at a grade of 11.4% sulphur and is currently undergoing an expansion to increase its capacity to 1050 tonnes per day. The following sections describe the reasons for selecting the BioX® Process, the plant design, the commissioning experience and the plant performance to date.

Jan 1, 1996

By Sepideh Javanshir

The process of liquid?liquid extraction of gold from chloride solution is complicated by the presence of many variables acting simultaneously in an unknown non-linear fashion for the determination of the final extraction efficiency. In practice, it is quite difficult to develop a model for predicting various properties of such process. Therefore, an alternative approach based on Artificial Neural Network (ANN) is considered to assist in the simulation process. Artificial neural networks (ANN) are powerful tools that can be used to model and investigate various highly complex and non-linear phenomena. This paper presents application of Multi-Layer Perceptions (MLP) to construct relationship that will be employed to develop a quantitative method for the estimation of the extraction efficiency as a function of hydrodynamical parameters such as stirrer speed, [AuCl4]- and DBC concentration, organic and aqueous flow rate. The model is shown to be in excellent agreement with the experimental data (R2 value for training and testing data was 0.9992 and 0.9915 respectively). Therefore, it can be used to optimize the parameters of liquid?liquid gold extraction with dibutyl carbitol (DBC).

Sep 1, 2012

By Freeman Bishop

Having obviously cleared the way for fast confirmation by the Senate Interior Committee, the Administration recently named Elburt F. Osborn, vice president of Penn State University, as director of the Bureau of Mines. Interior Committee Chairman Henry M. Jackson (D of Wash.) has moved hearings on the nomination right along, forecasting that they will end in a confirmation. Dr. Osborn not only made a good impression on the Committee, he also made it amply clear that he would be the boss of his job and have right-hand aides to help straighten out conflicts stemming from recently enacted health and safety regulations for miners. Mining engineers in the field hope this clarification can come soon. They are frankly fed up with local, state, regional and national overlapping of restrictions.

Jan 1, 1970

By JAMES J. DOUGHERTY

THIS paper is a general discussion of developments in our reverberatory smelting practice during the past five years. It deals briefly kith changes in furnace types ; changes in furnace feed ; de- crease in the amount of slag produced; relation of slag assays to copper loss in the slag; increase in the grade of matte and furnace capacities. The reverberatory is but one unit in the copper plant ancl 'any discussion of it must be considered with reference to the other units and to the plant as a whole. A change in the grade of concentrates at the concentrator puts an entirely new aspect on reverberatory smelting and, likewise, a change in the grade of matte at the reverberatory affects the rate and cost of converting. The introduction of differential flotation in 1927 had a very marked influence on all smelting operations and its effect will be noted

Jan 1, 1930

By H. W. Hintze

Coal handling facilities, such as utility plants, coal terminals, ports, manufacturing, processing, etc. Protect themselves against interruptions in coal receipts by maintaining a 60 to 90 day supply in storage. Not too long ago, these storage piles were referred in terms of hundreds or thousands of tons. Today, it is not unusual to have hundreds of thousands or millions of tons in storage. Burn, or shipping and receiving, rates have increased dramatically and storage piles have grown in horizontal as well as vertical dimensions. Fixed systems for delivering and retrieving coal from storage have increased in length and operating capacity and have become highly sophisticated in their operation. With few exceptions, mobile material handling equipment is used to handle all coal beyond the reach of these systems. These machines take excess loose coal from the fixed system to the storage area, place it on the storage pile, compact it, maintain the pile, protect against spontaneous combustion, the spread of fires if they occur, and return the coal to the reclaim system as required, plus many of the small chores around the facility.

Jan 1, 1981

By W. J. Schlitt

The development of forced aeration as an approach to enhancing copper sulfide leaching is traced from its earliest development up to present day operations using this technology. Both large-scale field tests and commercial operations are covered. The historical portion is preceded by an introductory discussion of the role of air (oxygen) in sulfide leaching. Sufficient aeration may be achieved via naturally induced convective airflow. However, cases are identified where this is insufficient, leading to retardation of the sulfide leach rate. For these situations, forced aeration may be beneficial. The historical portion of the paper focuses on the major field tests that have been conducted in Montana, Nevada and Utah, along with commercial operations in the United States, Australia, Chile and Peru. The paper concludes with some general comments about the applicability of forced aeration, which are based on the information developed to-date.

Jan 1, 2006

By N Berkowitz

Slurry pipelining, as now under-stood and practised, can only command marginal interest in Canada, as long-distance railway freight charges in this country are generally lower than slurry pipeline tariffs. Several alternatives to conventional slurry transmission are, however, now emerging, and two such alternatives are briefly discussed with reference to coal pipelining. In one, coal is transported as a suspension in oil; the other contemplates the movement of paste "slugs." Preliminary economic estimates indicate that the resulant process simplification and adoption of common carrier principles could lead to cost reductions that might make Western coal competitive with imported fuel in the power generating stations of eastern Canada. R ECENT appraisals of slurry pipelining techniques, some-what incongruously 'based upon the performance of successful installations (1,2), are making it increasingly apparent that current solids pipelining methods can only command marginal interest as methods of bulk commodity transportation (3). Although some uncertainties remain in matters of detail, it is now generally conceded that long-distance slurry pipelining cannot compete against rationalized railway transportation unless movement of

Jan 1, 1963

By Claudio Toro, Alex Moyano, Rodrigo Díaz, Hugo Jara, Carlos Caballero, Pedro Morales, Leandro Guzmán

"During 2007, Codelco-Chile carried out industrial tests to validate the Continuous Converting Process in the Teniente Converter of the Codelco Norte smelter.For planning and starting up the tests, a predictive, analytical and dynamic model was developed to support operational decisions. The dynamic model was developed using the METSIM® platform and dynamic data exchange (DDE) to operate smoothly with MS Excel®. The main feature of this real-time model was the on/off-line ability to work and gather data from either the smelter’s PI system® or database, and simulate the operation with or without scheduled events and controls.The main model development aspects, which include thermodynamics, metallurgical and kinetic considerations, slag-type and on-line mass and heat balances of the process, as well as the successful validation of the model, are presented and discussed in the present paper."

Jan 1, 2009

By Kenneth O. Bayard, John D. Sullivan

"This repot continues the series of papers being prepared at the Southwest Experiment Station of the Bureau of Mines in cooperation with the department of mining and metallurgy, University of Arizona, on factors involved in the leaching of ores. A list of previous publications is given in the bibliography at the end of this paper.Four factors are essential to the leaching of any copper-bearing ore: (1) A solution that will attack the copper minerals must get into the body of the ore particles; (2) the copper minerals must be dissolved by the solution; (3) the solution containing the dissolved copper must find its way out of the voids of the rocks into the main solution; and (4) the copper in solution must be recovered by some means of precipitation.This paper considers the third step -- the extraction of the water-soluble copper. From leached ores. The usual commercial practice in leaching copper ores consists of: (1) Crushing the ore to pass through a screen with approximately one-quarter or one-third inch openings; (2) charging the crushed ore into a fat; (3) adding, usually by upward percolation, a solution that will attack the copper minerals; and (4) washing the copper-laden solutions out of the ore with solutions weaker in copper than the pregnant solution, and finally with water. A countercurrent system of leaching and washing is usually employed. In some plants the solutions are circulated during leaching and washing, and in others they are not. In some plants the wash solutions are added by upward percolation, while at Chuquicamata the so-called ""piston-displacement"" method of washing is employed.The following condensed resume of the leaching cycle of the plant of the Inspiration Consolidated Copper Co. is given to present to the reader a picture of the steps involved in a typical leaching cycle. This plant was recently described in detail.4 The ore, crushed to pass openings 11/2 inches in diameter, passes over a screen with ¼-inch openings. The undersize from the screen goes to a Dorr howl classifier that removes the primary slime from the circuit. The overflow from the classifier is treated in a slime-leaching plant. The oversize of the ¼-inch screen passes through rolls, where it is crushed to pass a 3/8-inchscreen. The undersize from the 3/8-inch screen join with the coarse material from the classifier, and the combined products become the feed to the leaching vats. There are 13 concrete lead-lined leaching vats. The complete leaching cycle is 13 days. Eight tanks are under acid-iron content at all times; three are being washed, one tank is being filled, and the one is being excavated. The electrolyzed solution from the tank house, to which amount of new acid has been added, is put on the oldest tank of ore under need This solution travels from tank to tank and finally emerges from the newest ore"

Apr 1, 1931