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The American Institute of Mining, Metallurgical, and Petroleum Engineers, Inc, shall be composed of constituent parts as follows Section 1 Societies-The Institute shall be divided into three or more Societies as follows a The Metallurgical Society of AIME Society of Mining Engineers of AIME Society of Petroleum Engineers of AIME Other Societies may be formed within the framework of AIME or granted member¬ship in AIME as approved by the Board of Directors Section 2 Divisions and Subdivisions-The several Societies of the Institute shall consist of Divisions as follows a The Metallurgical Society of AIME 1 Institute of Metals Division 2 Iron and Steel Division 3 Extractive Metallurgy Division Any Divisions or Subdivisions which may be formed in the future with the approval of the Board of Directors Society of Mining Engineers of AIME I Coal Division 2 Industrial Minerals Division 3 Minerals Beneficiation Division 4 Mining and Exploration Division

Jan 1, 1960

By H. E. M. N. Moursy, Hassan El-Shall, M. N. Ahmad, Ponisseril Somasundaran, C. R. Holland, D. Chandler, G. Walker, Regis Stana, Patrick Zhang

The reaction kinetics of the acidulation of Moroccan phosphate rock (32% P?05) using 45% nitric acid were investigated using the Shrinking Core Model (SCM) , a newly developed Diffusion Model (DM) and a newly developed Combined Reaction and Diffusion Model (CRDM). This was achieved buy conducting a number of experiments in a controlled batch reactor system at different temperatures. The conversion and the reaction rate were calculated at different time intervals. The SCM based on reaction time and the CRDM characterize the acidulation process of phosphate rock. Diffusion of nitric acid and the products of acidulation at the early stages of reaction through the phosphate rock were found to be the rate-controlling step, resulting in a dif- fusion coefficient, which controls the mass transfer process. The reaction kinetics were found to fit first-order kinetics especially during early stages of reaction time. The deviation of the kinetics from first-order at later stages was due to the fact that the overall reaction was mass transfer controlled with respect to nitric acid and products diffusion.

Jan 1, 2002

By Jin Liu

To improve value metals recovery, a new operating concept for the Inco flash furnace was investigated, in which the flash furnace coke addition practice would aim at establishing a reducing barrier on the surface of the molten bath. The descending partially oxidized sulfide mineral particles would reach the supernatant coke layer prior to the formation of slag. The key for success depends on the effectiveness of the coke layer to produce a sulfur deficient matte while maintaining furnace production capacity. The mechanism and rate of carbon reduction of the phases formed in the flash flame, i.e. sulfide (matte) and oxysulfide was investigated by means of crucible graphite reduction test work. The objective of these tests was to measure the rate of solid graphite reduction of matte, oxysulfide, slag, and matte+slag mixture at different temperatures. These tests were expected to permit understanding and explaining the reduction mechanism of each material by comparing the respective reaction rates, product reaction gas compositions, reduced material assays, and other relevant parameters. Direct visual inspection of the test crucibles would provide additional information.

Jan 1, 2006

ENGINEERS AVAILABLE (Under this heading mill be published notes sent to the Secretary of the Institute by members or other persons introduced by members.) Member, technical graduate, aged 35, with experience as surveyor, sampler, assayer, and cyanide superintendent, desires position as engineer or cyanide operator.. No. 157. Member, technical graduate, aged 34, with 12 years' experience, including engineering, construction, and management of gold, silver, copper, and quicksilver mines, desires position of 'superintendent of gold or silver property. No. 158. Technical graduate, aged 24, with experience as mine surveyor and instructor in chemistry, desires position as instructor in mining. No. 159. Member, aged 29, technical graduate, with experience in exploration, prospecting, mining, concentration, smelting, and as assayer and chemist, engineer, and draftsman (smelter) in United States, Canada, Africa, South America., and Australia, desires to become connected with a firm of consulting engineers. Speaks Spanish. No. 1130.

Jan 11, 1914

By Maria I. De Rosa

Table 10 and figure 4 show the number of fires and fire injuries occurring at the surface of underground metal/nonmetal and stone mines by state during 1990-2001. Table 10 also shows the injury risk rates, employees' working hours, and lost workdays. A total of 12 fires occurred in 11 states; 5 of the fires caused 5 injuries. The yearly average was one fire and 0.42 injury. Five fires with two injuries occurred at metal mines, five fires with two injuries occurred at nonmetal mines, and two fires with one injury occurred at stone mines. None of the fires involved contractors. The Ewhr value was 58 x 106 hr (Irr = 0.017), and the LWD value was 75. Nevada had the most fires (two fires and no injuries). Ohio, New Mexico, Missouri, South Dakota, and Idaho each had one fire with one injury. Of these states, Missouri had the highest injury risk rate value (In = 0.25). Table 11, partly illustrated in figure 5, shows the number of fires, fire injuries, risk rates, employees' working hours, and lost workdays by time period. The number of fires decreased during most of the six time periods. The number of fire injuries and employees' working hours decreased during all of the periods. The Irr values follow patterns similar to those shown by the injury values. Tables 12-17 show the number of fires by ignition source, method of detection and suppression, equipment involved, location, and burning material by time period. Figure 6 shows the major variables related to fires for 1990-2001. Table 18 shows the number of fire injuries per number of fires causing injuries and total fires by year, ignition source, equipment involved, and location.

Jan 1, 2004

By David P. Conover

The effects of stress distribution on pillar stability were evaluated through rock mechanics studies conducted in three underground room-and-pillar coal mines. Results of the three field instrumentation programs are presented to illustrate the use of borehole pressure cells to determine vertical stress distribution in coal pillars at various stages of mine development. Vertical stresses and stress changes are determined using methods developed at the Bureau of Mines. Measured stress distributions are compared against theoretical distributions and possible explanations for observed behavior are discussed. Results indicate that stresses and stress changes calculated from pressure cell data are inconsistent with theoretical values; however, the data are useful for qualitative evaluation of trends over time. Calculated values are found to be sensitive to coal properties and initial setting pressures of the cells.

Jan 1, 1989

By R. E. Hathaway

INTRODUCTION For a number of reasons, the mineral processing industry has been slow to adopt significant automation of its processes. Most ore grinding operations can be operated manually so the pressure for automation to run the process has not been there. Furthermore, many plants had proven over years of operation that manual operation would produce a profit, so why automate? Until recently, the instrumentation necessary for automation of the grinding process was not available. All this has now changed, however. Increasing costs for labor, fuel, power, grinding media, reagants, and capital equipment have not been matched by the market price of the finished product. This cost-price squeeze has made it mandatory to reduce operating costs and improve the efficiency of the process to remain profitable. A number of other factors have added to the direct operating cost to make the problem even worse. These include environmental pressures to conserve natural resources and reduce smelter stack emissions; increased competition--especially from developing countries, varying economic conditions, stockpiling, governmental controls, and to top it all off, lower grade ore bodies. Although many of these other influences are good overall, they take their toll in cost per ton of delivered product. Fortunately, many operators are finding the solution to the cost-price squeeze in instrumentation and process control. Automated processes yield tighter spec. products, require less operating labor, produce more product from a given plant, and reduce unit cost by increasing process efficiency. The plant manager has a "handle" on the plant through process control so that he can change operating modes as economics dictate by changing the set point of some process controllers. He can even respond to economic conditions by switching control strategies from one that maximizes production to one that maximizes recovery, or minimizes unit cost, etc. This has all been made possible by the advent of on-line analytical instruments and practical process control computers. On-stream particle size (1), x-ray, and silica analyzers are now available and in use in many plants. Digital computer hardware and software has become practical so operators can use it without trying to become computer experts. This chapter describes the approach toward design and implementation of automatic grinding circuit control that has proven to be the most cost effective, based upon results from a number of

Jan 1, 1982

World refrigeration expert Professor Gunter Heinrich, at the invitation of the School of Mechanical Engineering at the University of the Witwatesrand (Wits), recently addressed industry representatives in Johannesburg on alternative refrigerants.

Jan 1, 1994

By Goode A, Belby P, Wardell-Johnson M

Within a short time frame of just over two years Wirralie Gold Mine has been developed from a virgin geological discovery to an operating gold mine with a mine life of six years and a production rate of 50,000 oz per annum. Construction work on the 600,000 tpa single stage crushing, two stage milling CIP plant commenced in November 1987 and was completed some seven months later at an overall cost of $A25 million. Since commissioning in May 1988 the treatment plant has exceeded design tonnage, availability and recovery. A second expansion programme in January 1989 will increase plant capacity to one million tonnes per annum with associated bullion production of 80,000 ozs per annum.

Jan 1, 1989

By Jesús L. Valenzuela

The solvent extraction for gold from alkaline cyanide solutions has been investigated using LIX 79 guanidine-based extractant. Different variables that could affect the extraction system were evaluated: equilibrium time, aqueous pH and extractant and modifier concentrations. The extraction isotherms of the aurocyanide complex with respect to other cyanoanions were compared and the following order of selectivity Au>Ag>Cu>Zn>>Fe was observed. From the pH isotherms it was found that aurocyanide can be extracted in alkaline media and a better separation with respect to other cyanoaniones was obtained in the range of 10.5 < pH < 11.2. Better selectivity was observed when using LIX 79 and tridecanol at 10% and 5% vol, respectively. Stripping gold from loaded organic was carried out at pH>12, using NaOH and NaCN solutions.

Jan 1, 2003

By J. Hanna, I. J. Anazia

The Mineral Resources Institute (MRI) has developed a unique process for selective fatty-acid flotation of phosphate minerals from siliceous ores. The process takes advantage of preferential adsorption (over silica surfaces) and high reactivity of fatty-acid collectors toward apatite under slightly acidic conditions (pH 5 - 6.5). The process was successful in upgrading various Florida and North Carolina phosphate samples containing 7.7% to 15.3% P205 and 62.3% to 77.2% siliceous gangue (Insol). The resulting concentrates assayed 28.1 % to 33.1 % P205 and 4.1 % to 9.9% Insol, with high P205 recoveries of 81.8% to 92.2%. Depending on the degree of surface contamination of the silica particles in the flotation feed, one or more cleaning steps, with or without sodium silicate, would be required to achieve higher-grade phosphate concentrates.

Jan 1, 1994

The Fosterville goldfield, situated 25 km east of Bendigo, Victoria was at its peak between 1894-1910 with some 47 600 ounces produced (Figure 1). Since that time intermittent mining and tailings retreatment was undertaken until Bendigo Gold Associates delineated a viable resource and commenced an open cut heap leaching operation in early 1991. Unfortunately they encountered financial difficulties and Perseverance Exploration purchased the mine in April 1992. Perseverance operated a successful heap leach operation nearby Nagambie, 100 km east of Fosterville. Based on its acquired heap leaching experience, Perseverance Exploration was able to instigate operational changes to create a successful small gold heap leach operation. The major change was the using of the original multi-use leach pad as a single-use leach pad. The available data allows for a comparison of the two operating systems.

Jan 1, 1994

By G. M. Wilemon

The Bureau of Mines conducted preliminary investigations on the use of an acid-alcohol leaching technique for recovering phosphate values from phosphate matrices. H2S04 and methanol were used as the leaching agents. Starting-matrix particle size and acid-apatite ratio were found to significantly affect the efficiency of phosphate leaching. Phosphate extractions of more than 80 pct were obtained when minus 200-mesh ore was used and the acid-apatite ratio was 8:1 or more. The iron and aluminum oxide impurities (R203) present in the crude acid products were monitored, and R203-P205 ratios are reported. Phosphate extractions in excess of 80 pct also were obtained when phosphatic clay wastes were leached. Typically, leach slurries filtered using test leaf filters ex-hibited filtration rates of 100 lb/(h·ft2) or more.

Jan 1, 1987

Haematite Recovery from a Tailings Stream Laboratory and pilot scale testing were carried out in 1999 on a tailing sample from a North American iron ore plant using gravity, magnetic and flotation techniques. The ore contained 16 per cent iron mainly in the fine (-38 ¦m) size fraction. Two methods of evaluating the separation efficiency were used; Selectivity Index and Z factor analysis. Tabling and magnetic separation produced poor results, with flotation testing showing some promise, generating concentrates assaying more than 40 per cent Fe. High grade concentrates containing more than 67 per cent Fe were generated using the Kelsey Centrifugal Jig, and it was recommended that this technique be pursued at a larger scale to verify the results obtained.

Sep 13, 2010

Magnetic separators are commonly used as an effective and low-cost method to concentrate paramagnetic or ferromagnetic minerals. These versatile separators are often used in multistage circuit configurations to further improve concentrate recovery and/or grade. Although these circuits are generally very efficient, field studies suggest that the layout of multistage circuits can be less than ideal for many industrial operations. To illustrate this shortcoming, a theoretical analysis has been conducted to compare different circuit configurations for magnetic separators. The resulting knowledge base has been used to establish generic engineering guidelines for the layout of multistage circuits. The potential improvements in separation efficiency that may be realized through the application of these guidelines have been demonstrated using a simple case study.

Jan 1, 2004

By C. H. Reynolds, Robert C. Pruess

Late in 1954 exploratory drilling indicated a substantial orebody on the Rattlesnake claims of Woodmont Inc., a Continental Uranium Inc. subsidiary. Woodmont holdings are located about 35 miles southeast of Moab, Utah. Only a few miles north of Big Indian Valley and other Chinle-sand uraninite deposits, the claims are in an entirely different geologic environment. Detailed development drilling continued until the Deep North orebody was outlined, covering an area approximately 300 x 450 ft. All exploratory drilling and about 80 pct of development drilling was wet core. Chemical assays for each hole sample, applied at 1-ft intervals against the measured core length, were put into a conventional ore calculation formula. Forty-six holes were drilled to delineate the Deep North orebody.

Jun 1, 1956

By Jin-Chun Kim

Nanosized W based powders (W, WC, WC-Co) were produced by chemical vapor condensation (CVC) process. Two different solid W(CO)6 and Co2(CO)8 metal-organic materials were used as precursors. CVC experiment was conducted in the temperature range of 400-1100. As-CVC WC1-x powder which had non-stoichiometric composition was carburized by using of the CH4-H2 mixed gas in temperature range of 750-850. The CVC W and WC1-x powders showed loosely agglomeration structure, and the Co powder had intricate long stand structure because of intrinsic magnetic properties of Co phase. Carbon content of the CVC powders controlled by gas-phase carburization was well matched with theoretical WC stoichiometry (6.13wt%).

Jan 1, 2006

By J. L. Huiatt, G. M. Potter, R. B. Tippin

The US Bureau of Mines and Great Salt Lake Minerals and Chemical Corp. have developed a froth flotation process for concentrating potassium salts from Great Salt Lake solar evaporites containing about 6% K2O. The crude evaporites, which were predominately halite (NaCl) with smaller amounts of kainite (KCl.MgS0, 2.7Hp0) and schoenite KeS0,MgS04.6H,0), were conditioned in a saturated brine solution., thus converting nonfloatable kainite to the floatable mineral schoenite. Schoenite was then selectively floated with a medium-molecular-weight fatty acid. Continuous small-scale pilot-plant testing indicated that 80% of the potassium could be recovered from the crude evaporites in flotation concentrates assaying about 13% K2O.

Jan 1, 1976

Measurable physical characteristics of placer deposits include both fluvial sedimentological features and their related heavy mineral suite. These two important identifiable factors are interrelated and generally compliment each other furnishing a complete geological geochemical tool in cost- effective placer evaluation. Heavy mineral suite composition and character will change downstream from the provenance area and will compliment the related sedimento logical environment. Careful attention to gold particle characterization during the exploration phase will indicate new source additions, enhance deposit evaluation techniques and greatly assist in maximizing metallurgical efficiencies. Recognition of the various characteristics within these fluvial systems should enhance the total contained geochemical suite evaluation considering a finite source. Failure to recognize and measure these two important features very frequently leads to frustrated and unrewarding placer evaluation objectives.

Jan 1, 1986

By P. D. Soares, H. R. Bullis, R. A. Hureau

"Echo Bay&apos;s Lupin Operation has produced more than 2.15 million ounces of gold since start-up in 1982. The mining method used to the£ present time has been longhole open stoping using trackless equipment. Because direct measurement of the open stopes has not been possible until recently determining the exact quantity of unplanned dilution in previously mined stopes has not been possible. Estimates of dilution hove been mode using comparisons between production tons and grades against ore reserve tons and grades.This paper discusses tracking of grade and tonnage measurements from reserve estimation through production. Comparisons between estimated and actual tons and grade are presented~ along with a brief description of methods used at the Lupin operation to control dilution."

Jan 1, 1994