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By Paul M. Jr. Musgrove, Thomas D. Henderson, Mark B. Nesbitt

Introduction History. Vat leaching operations began in October 1953. Since that time, about 4.2 million tpy of oxide ore averaging 0.72% copper have been treated. Economic conditions, improved operating methods, and increased production facilities have allowed the ore grade to be reduced over the years. The average ore grade was 0.84% Cu during the 1950s, 0.68% in the 1960s, and was 0.50%, or below, during the 1970s. By 1961 some 35 million tons of oxide ore had been removed from the mine, exposing the underlying sulfide ore body. A 5,000¬ tpd concentrator was constructed and put into operation to treat the sulfide ore in November 1961.1 The original crushing plants were used for both oxide and sulfide ores, but by adding a third crushing shift, six days per week, the total ore treated was raised from 12,000¬17,000 tpd. Dwindling oxide ore reserves prompted expansion of the concen¬trator to 13,500 tpd in 19672 so that exhaustion of known reserves of both ores would occur at nearly the same time. Again, no additional crushing facilities were added during the expansion. It was anticipated that the treatment rate of oxide ore would be reduced from 12,000 to about 9,000 tpd. Instead, the addition of two crushing shifts per week and improved operating procedures have increased the treatment rate to 26,500 tpd-13,000 tpd to the leaching plant and 13,500 tpd of sulfide ore treated in the concentrator. Geology. The Yerington ore body is a disseminated porphyry copper deposit where a quartz monzonite porphyry intrusive contains the major part of the ore values, with some copper mineralization found in the adjacent granodiorite and other rocks. The geology of the Yerington Mine was aptly described by J. R. Wilson (Mining Congress Journal, June 1963). The ore body is one of few porphyries where geologic conditions were optimum for the formation of an important concentration of oxidized ore with minimum economic effect of secondary enrichment. The principal mineral in the oxidized zone is chrysocolla, which irreg¬ularly occurs dispersed throughout the rock and as narrow seams along fissures. The grade of the ore in the oxidized zone is generally the same as in the primary sulfide below, although some massive concentrations do occur as vein-like fissures and as filling of breccia zones. Cuprite, tenorite, melaconite, malachite, and azurite occur but are not abundant. Lying between the chrysocolla horizon and the primary sulfide is a transition zone whose limits are irregular but seldom with vertical range exceeding 50 ft. Chalcocite, cuprite, melaconite, native copper, and chrysocolla occur superimposed on the primary sulfides. Chalcocite replacement of chalcopyrite, as well as secondary chalcocite, do occur. The primary sulfide minerals are chalcopyrite and pyrite, with minor amounts of bornite and covellite. There are no appreciable gold or silver values, and molybdenite has rarely been noted.

Jan 1, 1985

By Douglas J. Cork

Biohydrometallurgica1 processes in the areas of coal desulfurization (Hoffman, M.R., et al., 1981) and recovery of metals from metal sulfide ores (Lundgren, D.G. and Malouf E.E., 1983) are well known. In addition other areas of potential interest to the mining industries have been outlined (Spisak, J.F., 1979). For instance, scientists involved in this interdisciplinary field have endeavored to apply microorganisms to the problem of effluent sulfate disposal. This chapter assesses the feasibility of utilizing a particular two stage anaerobic microbial process to convert effluent sulfate to orthorhombic sulfur. Strong emphasis is given to the possibility of substituting this bioprocess for alternative gypsum synthesis and disposal methodologies commonly practiced by companies which emphasize sulfuric acid or ammonium sulfate leaching processes.

Jan 1, 1985

By U. K. Custred, E. W. Long, V. R. Degner

In 1973, the United States production of marketable phosphate rock set a record in excess of 42 million tons. This production rate is expected to continue to increase, due to the growing international requirement for fertilizer, at a rate exceeding 5% per year well into 1977. One approach towards increasing plant production capacity to meet the growing demand is through large-capacity flotation cells, provided they achieve metallurgical performance comparable to existing smaller cells. A 9-month evaluation test program recently completed at the Haynsworth mine at Bradley, Fla., demonstrated the feasibility of achieving economically acceptable concentrate grade and recovery levels using large, high-capacity flotation cells. The composition of the feed to the Haynsworth beneficiation plant is a typical Florida pebble phosphate matrix composed of phosphorite pebbles ranging in size from approximately 1-1/2 in. down to 150 mesh and intimately associated with a mixture of clay and sand (essentially silica). The feed contains approximately 22 to 28% phosphate reporting as tricalcium phosphate, Ca3 (PO,) 2, or "bone phosphate of lime" (BPL). The flotation section utilizes the double-float procedure typical of Florida plants. The phosphate is first floated away from the silica in the rougher circuit, using crude fatty acid, ammonia, and fuel oil or kerosene. Rougher conditioning is accomplished at 60 to 70% pulp solids with sufficient ammonia added to raise the pH to 9 to 9.5. Following coarse and fine rougher flotation, the concentrate (overflow) streams are joined and conditioned (sulfuric acid cleaned and washed) prior to entering the cleaner circuit where an amine float (cationic reagent and kerosene added in the feed box; pH 7.3 to 7.8) is employed to float the silica. The feed to the coarse rougher circuit averages 29% +35 mesh while the fine rougher feed averages 10%+35 mesh. Primary attention was directed toward the large rougher cell performance (recovery and grade) on coarse feed during the Haynsworth evaluation program. Flotation Cell Test Program A row of three No. 120 size (300 cu ft) WEMCO flotation cells was installed in parallel with an existing air-cell row. The total installed volume of the large cell circuit was 900 cu ft and required a floor space of 306 sq ft. This compared to the air-cell total volume of 200 cu ft and 152 sq ft floor space. (Both floor areas include feed and tails hoppers but exclude walkways.) Fig. 1 is a schematic cross section of the large flotation cell showing the relative location of key mechanism elements. In operation, the rotor generates a fluid vortex extending up along the walls of the standpipe and creating a sufficient vacuum within its core to ingest air into the standpipe/rotor cavity through the air inlet duct. The ingested air mixes with the pulp, which has been recirculated through the false bottom and draft tube, in the rotor. Further mixing occurs as the air and pulp move radially outward from the rotor, finally passing through the disperser into the flotation cell. Flotation is accomplished outside the disperser, where phosphorite laden air bubbles rise and the remaining pulp recirculates down along the cell wall to the false bottom and draft tube. Large-flotation-cell performance is influenced by the ability of the mechanism to (1) circulate, or suspend, the solids in the pulp; (2) ingest air into the rotor cavity; and (3) mix the air and pulp effectively. The proper balance between pulp circulation and air ingestion is a key consideration in achieving good recovery in a course feed application. Large-flotation-cell pulp circulation and air transfer characteristics are significantly influenced by rotor speed and rotor submergence; therefore, these two operational parameters can be used to "optimize" a particular mill application. Fig. 2 maps the hydraulic performance of the WEMCO No. 120 size flotation cell. This map can be used to relate the cell operational parameters which influence metallurgical performance. At a given rotor speed, power intensity (i.e., pulp circulation) is seen to increase, and airflow decrease, as rotor submergence is increased. The inverse relation between power and airflow is due to the two-phase air-liquid mixture density reduction accompanying the increased airflow rate. For any fixed rotor submergence, the power intensity (i.e., fluid circulation) and airflow both increase as rotor speed is increased. The selection of these two mutually related operating conditions (i.e., rotor speed and submergence) was a key consideration in the Haynsworth evaluation program.

Jan 1, 1976

CONTENTS PAGE SOUTHWARD, GLENN B.-Coal Mining by the V System. Discussed by Glenn B. Southward, Howard N. Eavenson, Edward B. Raiguel, Thomas de Venny, J.. J. Rutledge, Graham Bright 1 ARMS, RAY W.-Dry Cleaning of Coal. Discussed by Ray W. Arms, S. A. Taylor, Howard N. Eavenson, W. R. Crane, F. F. Jorgenson, G. R. Delamater. 6 EAVENSON, HOWARD N.-Data About Labor Employed in Various Bituminous Coal Mines. Discussed by Howard N. Eavenson, George H. Ashley, E. A. Holbrook, S. A. Taylor, R. V. Norris 14 Coal Mining by the V System Discussion of the paper of GLENN B. SOUTHWARD presented at the New York Meeting, February, 1924, and issued as Paper No. 1300-C, with MINING AND METALLURGY, February, 1924. MEMBER.-How do you pay your men for loading? GLENN B. SOUTHWARD.-Everything is on time basis. MEMBER.-How many men do you have loading across one of the diagonal faces? GLENN B. SOUTHWARD.-From four to five, depending on the length of the face. Our loaders are so arranged that each will give about twenty-five tons per loader a day. MEMBER.-How much timber do you use? GLENN B. SOUTHWARD.-A little more than will be used in room-and-pillar work, but not a great deal more. We have not attempted to recover much of our timber. MEMBER.-IS the mine union or non-union? GLENN B. SOUTHWARD.-Non-union. HOWARD N. EAVENSON, Pittsburgh, Pa.-I went .through that mine some weeks ago, and think that the author has been very conservative in his statements about the amount of loadings, etc. EDWARD B. RAIGUEL, Huntington, W. Va. (written discussion).¬Page 7 shows 60 men employed in day and night shifts, inside the mine, producing 750 tons per day. Let us reduce these figures to a man-hour basis, for comparison with Mr. Eavenson's figures on West Virginia mines, as follows:

Jan 6, 1924

By JOHN BUZARD

In the following pages the writer has endeavored to give an account of the many methods, advantages, and disadvantages of preparing and burning powdered coal. For much of the information jmparted the writer is indebted to various firms in the United States and Canada whose plants he visited, also to the technical press. Manufacturers and operators of coal-fired furnaces can not afford to disregard the possible advantages of pulverizing their coal before burning it. Hence the purpose of this bulletin will be fulfilled if it leads them, after making careful estimates, either to abandon their present method of burning cGal on grates or stokers, and install the pulverizers, conveying system, powdered-coal feeder, and burner best suited for burning powdered coal in their plant, or to reject, as uneconomical, the replacement of their present system of burning coal by a system for pulverizing and burning it. No research work on the combustion of powdered Canadian fuels has been carried out by the Mines Branch, Department of Mines; but if this research work were carried out scientifically by competent engineers, chemists, and physicists, it would show how powderedcoal furnaces and burners should be designed to gi ve an efficiency higher than they give at present.

Jan 1, 1923

By J. D. Vincent

The Mission By-Products Plant at the Mission Concentrator of the American Smelting and Refining Company, some 20 miles south of Tucson, Arizona in the Twin Buttes mining area, has been in operation slightly over a year. The plant produces a molybdenite concentrate a copper concentrate and, at times, a zinc concentrate. An intensive test program of over three years was conducted at Mission to make the recovery of molybdenum economic. Briefly it was found necessary to make a scavenger concentrate in the copper mill before it was possible to recover economic quantities of molybdenite. By so doing there were resulting additional recoveries of copper, silver and zinc. This program resulted in the design and construction of the existing plant. A paper by the authors discussing the test work and the design was presented at the, October 1965 meeting of the American Mining Congress in Las Vegas, Nevada, by Mr. Wayne Bossard.

Jan 1, 1966

By Samuel S. Wyer

The following abbreviations hare been used in the text: Am. Mfr. and Iron World. American Manufacturer and Iron World, Pittsburg, Pa. Can. Min. Rez;. Canadian Mininy Review, Ottawa, Can. Cuss. Mag. Cassier's Mayazine, New York, N. Y. Col. Eng. Colliery Engineer, Scranton, Pa. Col. Guard. Colliery Guardian, London, Eog. Eng. and Min. Jour. Engineering and Mining Journal, New York, N. Y. Eng. News. Engineering News, New York, N. Y. Engr. Engineering, London, Eng. J. I. and 8. I. Journal of the Iron and Steel Institute, London, Eng. Jour. W. Soc. of Engrs. Journal Western Society of Engineers, Chicago, Ill. Lond. Eng. The Engineer, London, Eng. Mines and Min. Mines and Minerals, Scranton, Pa. North of Eng. I. M. E. North of England Institute of Mining Engineers, Newcastle-on-Tyne, England. Oester. Zeitsch f. Berg u. Hüt. Oesterreichische Zeitschrift für Berg-und Hittenwesen, Vienna, Austria. Pract. Engr. Practical Engineer, London, Eng. Proc. I. C! E. Proceedings Institution of Civil Engineers, London, Eng. Trans. A. I. M. E. Transactions American Institute of Mining Engineers, New York, N. Y. Trans. A. S. M. E. Transactions American Society Mechanical Engineers, New York, N. I!. Duns. Fed. I. M. E, Transuctions Federated Institute Mining Engineers, Newcastle-on-Tgne, England. Trans. L 31. E. Transactions Institute Mining Engineers, Newcastle-on-Tyne, England.

Jan 1, 1907

By G. Akdogan, C. Dorfling, L. Bumhira

Hydrometallurgical recovery of rare earth elements from phosphors obtained from end-of-life fluorescent lamps involve leaching followed by extraction of dissolved metals from solution. The similar properties of rare earth elements make selective recovery of individual elements challenging. The aim of the paper was to investigate the extraction of primarily Ce and Tb from a hydrochloric acid leach solution also containing other REEs (Y and Eu) as well as impurity metals (Ca and Al). Batch solvent extraction tests were performed with di-(2-ethylhexyl) phosphoric acid (DEHPA) at various extractant concentrations, pH conditions and organic-to-aqueous ratios. Extraction isotherms and separation factors for the range of conditions tested in the project are presented in the paper. Separation of individual REEs was not possible at the conditions investigated, but a mixed REE product can be obtained in a single extraction stage operated at a pH of 0.5 using 1 M extractant at an O/A ratio of 1; these conditions resulted in complete Y and Eu, 95 % Ce and 99 % Tb extraction with Ca and Al coextraction below 7 %. A better quality mixed REE product can be produced by operating at lower pH and extractant concentration conditions, but this would require a larger number of extraction stages. Keywords: Cerium, fluorescent lamps, phosphors, rare earth elements, solvent extraction, terbium

Jan 1, 2020

By N. Brown, R. Kossowsky

Plain carbon steels with 0.48 and 0.95 pct C were quenched and tempered at 705°C to produce carbide dispersions with spacings on the order of 1 p. The morphology of the structure consisted of a carbide-dislocation network. The strengthening due to the dispersion was found to vary linearly with M-½ where M is the mean free-ferrite path determined by the entire network. No preyield microstrain preceded the upper yield point. After prestraining, the lowest stress at which dislocation movement could be detected was 104 psi; this frictional stress was independent of the dispersion and prestrains up to 6.5 pct. The Cottrell-Petch equation for grain-size strengthening was used to discuss the dispersion strengthening in this investigation. The results support an impurity mechanism for the upper yield point rather than one based on the Johns ton- Gilman theory. IT was pointed out by Gensamerl that the mean free path in the matrix is the important variable which controls the degree of dispersion strengthening. Gensamer's data on steels showed a linear relationship between the yield point and the logarithm of the mean free-ferrite path. The first theory was by Orowan,2 who suggested the mechanism of dislocations bowing between particles, with the resulting relationship that where a is the yield point, P is the distance between particles, and G is the shear modulus. The Orowan equation applies to that range of dispersion where P is large compared to the particle size and the particles are not coherent, so that the matrix is essentially free of internal stresses. As was pointed out by Orowan, there are different degrees of dispersion which, in turn, will influence the strength-controlling mechanism. However, in this investigation, we wish to confine ourselves to the region of coarse dispersion, where the Orowan mechanism should, intuitively, be applicable. It was soon evident that the data, which existed at about the time that the Orowan theory was proposed, did not agree with the Orowan equation in that the actual strengthening was always greater than the theoretical predictions. Thus, Fisher, Hart, and pry3 modified the Orowan theory by suggesting that the Orowan process took place in the microstrain region preceding macroscopic yielding and the subsequent, rapid work hardening in the form of residual dislocation loops around the particles largely determined the observed macroscopic yield point. The F-H-P theory stated that the increment of strengthening due to the work-hardening mechanism was proportional to f3/2 where f is the volume fraction of the precipitate. F-H-P used data by Shaw, Shepard, Starr, and Dorn4 on A1 3-5 pct Cu to support their theory. Roberts, Carruthers, and Averbach5 were the first to make a microstrain study of dispersion strengthening, and they found that for steel the Gensamer relationship was obeyed. Hayman and Nutting6 suggested that in the case of tempered steel 1) the ferrite grain boundary was the primary obstacle and 2) the carbide particles simply formed part of the grain boundary obstacles. They found that the strength varied as G"" where G is the ferrite grain size. Turkalo and LOW' determined the structure of quenched and tempered plain carbon steel using a replica technique; they concluded that the carbide particles did not necessarily lie in the ferrite grain boundaries. When they defined the mean free-ferrite path as being determined by both particles and the ferrite path boundaries, their data obeyed the Gensamer relationship. Meiklejohn and skoda8 showed that the strengthening from iron particles in mercury was a function of the particle size and the distance between particles. Dew-Hughes9 explained the Meiklejohn and Skoda results by the following theory: 1) grown-in dislocations which surrounded the particles were produced by the thermal stresses during the time the material was cooled to the testing temperature, and 2) the observed strengthening was associated with the cutting of the grown-in dislocations by the glide dislocations. Ansell and Lenel10 proposed a theory in which the glide dislocations pile-up or surround the particle until the number of dislocations in the pile-up is sufficient to yield or fracture the particle. The resulting theory says that the yield point varies as p-1/2, where P is the distance between particles. The Ansell-Lenel theory is almost identical to the

Jan 1, 1965

By R. V. Tomkins

Introduction In the past fifteen years, bentonite production in the United States has increased very rapidly. From l50,000 tons in l934, it rose to 400,000 tons in l 942 and to a record 920,000 tons in 1948. Occurrences of bentonite ?in western Canada were described in the early part of the century by Ries and Keele (3), Ells (2), Spence (4). and other officer of the Dominion Department of Mines, and, in Saskatchewan, Professor W. G. Worcester (6), of the University of Saskatchewan has explored and carried out tests on deposits of the material at various times during the past thirty years. Nevertheless, Canadian developments have not kept pace with those in the United States. Exact figures for total Canadian production are not released, but it is probably under l0,000 tons annually. There are a great number of uses for bentonite, but in the United States three alone account for 90 per cent of the consumption. These are for drilling gel, foundry sand bond, and oil decolorizer. The tremendous oil exploration programme on the Canadian prairies has increased the possibility of an adequate domestic market for drilling gel. Therefore, in Saskatchewan, our efforts have been directed mainly to the production of a good drilling gel, with a minor part of the project devoted to activation of bentonites for oil decolorizing.

Jan 1, 1952

By Michel Grenier, David A. Young, Brent Rubeli, Mahe Gangal

This work forms part of a study being conducted into the feasibility of constructing a prototype Load Haul Dump (LHD) vehicle for operation in underground Canadian mines. A series of in-laboratory assessments were performed to determine the relative contributions of a Deutz F4L912W (54 HP- conventional vehicle engine) and a Deutz F2L 2011 (31 HP hybrid vehicle engine) to diesel emissions concentration during mine transient duty cycles and steady-state operating modes respectively. Results indicate significant reductions in the quantity of gaseous and particulate emissions with the application of the hybrid system and a high-efficiency ceramic element filter. This paper describes the engine test protocols and the results of this laboratory study.

Jan 1, 2006

By Mayo D. Hersey

This problem was taken up by the Bureau of Mines as a result of the oxygen explosion at the Jefferson Physical Laboratory of Harvard University , and other similar explosions , in order to supply technical information which would serve as a basis for safety precautions in the future . At the Harvard inquiry the fact was emphasized by the Massachusetts State Chemist , Mr. Walter M. Wedger , that we do not know the limiting pressures and temperatures above which compressed oxygen and lubricating oil are capable of spontaneous explosion . The problem is also of importance to the Army and Navy and to the Bureau of Mines in connection with the hazards incurred in the use of compressed and liquified gases in the repurification of helium. In fact , the risk of fires and explosions due to traces of lubricating oil coming into contact with high pressure oxygen is of serious interest to all concerned with the manufacture and use of oxygen , a commodity which is now being produced and distributed on a rapidly increasing scale . Part of the money expended on this investigation was from Army and Navy funds . The problem was undertaken by the Bureau of Mines at the suggestion of Dr. Charles E. Munroe , Chairman of the Committee on Explosives Investigations of the National Research Council , and referred to the Pittsburgh experiment station of the bureau.

Jul 1, 1923

By Thomas A. Palmer

Free cyanide has been successfully determined in ground and surface water by isolation using a microdiffusion technique followed by a pyridine-barbituric acid photometric finish. The determinable free cyanide includes hydrocyanic acid, ionic cyanide, and the cyanide of weak complexes such as cadmium, lead, manganese, and zinc cyanide. Complexes of copper, iron, nickel, and silver are sufficiently strong so the dissociation into free cyanide does not occur at the microdiffusion pH 7. Cyanate, sulfide, and thiocyanate do not interfere. Samples containing as little as 0.005 mg/L are successfully analyzed. The relative standard deviation is <10%. Samples stored at pH >I2 in the dark are stable at least three weeks.

Jan 1, 1981

By D. L. Buchanan, M. J. Jones

In May, 1982, the University of Western Australia in Perth hosted a three-day international conference on nickel sulphides, followed by a ten-day field excursion to the Western Australian nickel sulphide deposits. The conference was arranged by Professor D. I. Groves, C. M. Lesher and Dr. R. E. T. Hill under the auspices of the International Geological Correlation Programme's Projects 161 ("Sulphide deposits in mafic and ultramafic rocks") and 91 ("Metallogeny of the Pre- cambrian"). This volume contains a selection of the papers that were presented at the Perth conference. Most of the papers are related to the nickel sulphide deposits of Western Australia, but contributions with settings in North America, Finland, Vietnam and Southern Africa have also been included.

Jan 1, 1984

By A. G. Darnley

"The Geological Survey of Canada is devoting consider-able effort to finding ways to make airborne radiometric measurements both quantitative and reproducible. Equipment has been designed by Atomic Energy of Canada Ltd. to provide high count rates in order to minimize statistical uncertainty, this being essential to justify applying terrain clearance and Compton scattering corrections to the data. Correction factors have been determined experimentally and can be applied automatically. Considerable ground work has been undertaken in parallel with equip-ment development in order to determine field parameters.Two airborne gamma-ray spectrometers are described. One is based on three 5-by 5-inch Nal(Tl) detector crystals and is used in a helicopter; the other employs twelve 9-by 9-inch detectors and is used from a Skyvan aircraft. The former is being used to investigate the range of radiometric response from specific rock types, and for the correlation of ground and air measurements; the latter has completed initial trials and will be commencing experimental surveys during 1969.Results are given to show the comparative sensitivities of the various detector sizes, and operational requirements for a gamma-ray spectrometer survey are discussed briefly."

Jan 1, 1970

By Rodney R. Koch

The phys ica l properties of gold should faci li tate its detection. However, the average concentration of gold in the ea rth's cru st precl udes the ability of geophys ical methods to detect the metal directly. In order to effectively pursue geophys ical detection and location of gold mineralization there must also be geological input - i.e. a model. This will lead to the selection of the appropriate geophy ical techn ique(s) in the fo llow-up programs. Interpretation of the fie ld survey data is of con iderable importance in expl oration geophys ics. Analytical techniques such as cont inuation, fi ltering, and spectral analys is, are now routinely empl oyed in the analys is of survey data to aid the interpretation process. The uccessful application of geophysical methods and the use processing techniques is demonstrated by two recent Saskatchewan discoveries. The first, the Star Lake 21 Zone, with reserves of 230,000 tons grading bener than 17.14 git Au associated with a disseminated sulphide quartz-bearing shear zone yielded geophysical anomalies that were distinctive, obvious, and would have lead directly to its discovery. The econd , the Laurel Lake Prospect, with ore grade intersections of 28.0I git Au and 52 .46 git Ag· over 5.49 m, resulted in numerous geophysical anomalies. However, none of these anomalies demonstrated association with the mineralization until processed results indicated that structural features crosscuning regional trends were important cont rol mechanisms for the mineral ization.

Jan 1, 1986

By J. A. Cooper, M. R. Petersen, P. O. Jackson, J. C. Langford

The organic, inorganic, and radiological characteristics of airborne aerosols have been measured as a function of particle size in controlled atmosphere test chambers and operating uranium mines. Concentrations of benzo(a)pyrene in two mines ranged from 26 to 57 ng/m3 of air. The carbon chain length of adsorbed n-alkanes was correlated with particle size. Normal mining activities produced an ore dust aerosol with mass median aerodynamic diameter (MMAD) greater than 2 µm. The elements Na, Al, Si, K, Ca, Ti, V, Fe, and U exhibited elemental ratios similar to bulk ore and had comparable MMAD&apos;s. The S, Zn, and Pb were higher in aerosols than bulk ore and were associated with smaller MMAD particulates. Radon daughter particle size distributions were influenced by the kinds of particulates

Jan 1, 1981

The Engineers Society of Western Pennsylvania, William Penn Hotel, Pittsburgh, Pa. K. F. Treschow, Secretary Since 1880 this society has been publishing Proceedings containing papers on a wide variety of engineering and related subjects The Proceedings are issued monthly throughout the year (except for the months of August and September each year) and the monthly Proceedings from February of one year to and including January of the next are also collected in volumes The latest is vol 47, for 1931 The rate of subscription for the monthly Proceedings is $5 per year throughout the Postal Union, single copies (monthly), 50 cents each The price of the yearly volumes from vol 17 to date is $5 each, unbound. The Secretary will quote prices for

Jan 1, 1933

By James R. Arnold, John G. Mansanti

BACKGROUND Gold Fields Mining Corporation started their Chimney Creek operation in late 1987. The dump leaching operations commenced in October and the mill startup took place during the last half of November. The milling process at Chimney Creek is almost a conventional CIP plant with the major exception being the installation of a second thickener as a washing stage prior to CIP12 &apos; A simplified flow sheet is shown in FIGURE 1. Most of the gold in the mill is liberated by grinding in cyanide. After grinding in a conventional SAGball sill circuit, the cyclone overflow reports to the thickeners. The thickener overflow from the first thickener is pumped through 5 stages of carbon columns for gold adsorption. Slurry from the thickener underflow is leached, washed, re-thickened and pumped to CIP. After approximately 6 hours of retention time in CIP, the slurry is pumped to tailings for final deposition. The carbon from CIP is transferred to the third stage of carbon adsorption and then advanced as normal through the columns and strip circuit. Loaded carbon is stripped at approximately 1500C (3000 F) and 55 kPa (80 psi). After 4-5 hours of stripping, the grade of four tons of carbon is reduced from 5,140-6,860 g/t (150¬200 oz/ton) to typically 17 g/t (0.5 oz/ton). Zinc precipitation is used to recover the gold from the pregnant solution. After a water wash, the stripped carbon is transferred to acid wash or to the kiln feed tank for reactivation. All strip lots are acid washed in 3.0% HC1 and all stripped carbon passes through the kiln for thermal reactivation. Then the carbon is quenched, sized and stored in tanks or bins for transferral to CIP or to the carbon columns. Loaded carbon from the dump leach circuit is treated in a circuit consisting of five carbon columns which are parallel to the mill circuit. No distinction is made between the mill or dump leach carbon once the gold has been stripped from them. CARBON FINES HANDLING Once the carbon has been transferred from the carbon columns, every effort is made to minimize carbon attrition and to recover the carbon fines that have been generated in the process. Wemco Hidrostall pumps are used for transferring all carbon in

Jan 1, 1990

This is the final report on Project CR-102 under the auspices of the Department of Environmental Resources, Commonwealth of Pennsylvania .1 The project was activated on January 5, 1970, with financial support from Bituminous Coal Research, Inc., the Pennsylvania Coal Research Board, and the Water Quality Office of the Environmental Protection Agency (WQO-EPA)2 through Grant 14010 EST to the Commonwealth of Pennsylvania Work on the project was conducted according to the procedures outlined in BCR Research Program Proposal RPP-164R, dated April 24, 1969, and a subsequent Revulsion of Scope, dated February 25, 1970. This report covers the entire project period from January 5, 1970, through March 25, 1971.

Jan 1, 1971