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Traditional fine-grinding mills in mineral processing rely on the use of steel medium for particle size reduction. The IsaMill (a horizontal stirred mill with a 3, 000-L net chamber volume and a rated motor power of 1,120 kW) has successfully adopted a novel product-separator design, which has led to a major revision of medium use in fine grinding. The IsaMill separator has successfully operated with a wide variety of medium types, such as silica sands, river gravel, smelter slag, run-of-mine ore fractions and steel shots, with sizes ranging from 0.5 to 10 mm. Great flexibility in the IsaMill operations has been achieved for processing feed as coarse as 120 µm to a product as fine as 5 µm. This paper discusses the performance of two inert grinding media in the IsaMills that are used at Mount Isa Mines Ltd. for the processing of galena- and sphaleritebearing ores to below 10 µm. They are the copper reverberatory furnace slag (CRFS) and the heavy-medium plant rejects (HMPR). A methodology was also demonstrated for comparing the medium performance and for scaling up the IsaMills from a 1.5-L laboratory batch mill to a full-scale 3,000-L mill.

Jan 1, 2002

By Dewa Putu Agus Sudarmaja, Widi Astuti, Januarti Jaya Eka Putri, Yayat Iman Supriyatna, Himawan Tri Bayu Murti Petrus, Siti Nurul Aisyiyah Jenie, Ferian Anggara, Anisa Maulidia, Vincent Sutresno Hadi Sujoto, Sutijan, Wayan Warmada

With the rapid development of renewable energy technology, such as electricity reserves to store energy for electric and battery-electric vehicles, environmentally friendly energy sources are expected to continue to be tapped to meet mobility and stationary energy needs. Lithium is the most attractive candidate for energy storage development. The presence of lithium in the Sidoarjo mud in Indonesia makes it interesting for use as an alternative source of raw material. In this research work, sulfuric acid was used as a solvent, and hydrogen peroxide as an oxidation agent. Lithium recovery was studied with a leaching process at various temperatures (30, 45 and 60 °C) and sulfuric acid concentrations (0.5, 1, 2, 3, 4 and 6 mol/L). Increases in temperature, sulfuric acid concentration and hydrogen peroxide concentration were found to directly affect the lithium recovery. An increase in sulfuric acid concentration is directly proportional to the recovery of lithium in the solution. Optimum conditions were obtained at a sulfuric acid concentration of 6 mol/L and temperature of 60 °C. The shrinking core model was used to study leaching kinetics, and ash layer diffusion acted as a rate-controlling process. The activation energy of the leaching process was found to be 38.90 kJ/mol.

Jun 1, 2024

By Y. Gonzalez

The reaction kinetics of sulfuric acid leaching of brass foundry flue dusts were studied. As received flue dusts were separated into two size fractions, a coarse fraction (median size 103.5 microns) and a fine fraction (median size 3.4 microns). Experiments were performed on both fractions al-though most results are reported for the coarse fraction. Leaching experiments were conducted in a stirred reactor to determine those factors which govern the kinetic response of the sys-tem. The kinetics were followed by determining the increase in the concentration of dissolved copper with time. The effects of temperature, concentration and oxygen partial pressure on the rate of copper dissolution were examined. In addition, reaction residues were examined by SEM, EDAX and X-ray analysis. An activation energy of 4.8 kcal/mole (20.08 kJ/mole) was calculated. The rate was found to be significantly dependent on oxygen partial pressure and acid concentration. The order of the reaction with respect to oxygen partial pressure and hydrogen ion concentration was approximately 1.50 and 0.37, respectively. Results indicate that the dissolution reaction is limited by diffusion and can be described by the following mathematical expression where ~ is fraction of copper extracted in time t

Jan 1, 1988

By Lynn I. Kelley

Industrial-mineral production contributes on the order of CDN$1.5 billion per year to the economy of Saskatchewan, Canada. Potash is the primary industrial mineral produced in the province, followed by aggregate, sodium sulphate, salt, potassium sulphate, peat, clays, silica sand, calcium-chloride brine, and "clinker." Potentially economic deposits of kaolin, diamond, and building stone are under evaluation. All of the major deposits of commodities discussed here are located south of lat 56°30'N. Most of the locations are presented on Figure 1 POTASH Saskatchewan supplies approximately 30% of the world's demand for potash. Eight conventional underground mines and two solution mines produced an all-time high of more than 13 million metric tonnes of potash (KC1) (8 million tonnes K2O equivalent) in 1997, valued at over CDN$1.4 billion. All of the mines exploit the middle Devonian Prairie Evaporite. The four potash-bearing members of the Prairie Evaporite underlie much of southern Saskatchewan and extend southward into North Dakota and Montana and east- ward into Manitoba (Fig. 2). Halite and anhydrite separate the potassic members. The evaporite beds locally are interrupted by solution structures (anomalous ground). Some of the structures may be related to deep-seated tectonic features.

Jan 1, 1999

The value of US industrial mineral output last year was $20 billion, a 12% decline from 1981. A glance at the accompanying table and a review of the comments made by several contributors reveals how poorly many industrial minerals fared during 1982. Bromine and peat posted a 6% increase in production while recovered sulfur showed only a 3% gain over 1981. Pumice, salt, and dimension stone managed to stay more or less even with 1981 production. However, most industrial minerals posted losses for the year from 3% for vermiculite, sodium carbonate, and gypsum to 49% for ilmenite. Naturally, the depressed economy is to blame for decreased production and consumption. Low steel and auto production, lower housing starts, and many other factors impacted the industrial mineral industries. Hopefully the worst is over. Interest rates are down, energy costs are decreasing, the automobile and housing industries show signs of picking up. Perhaps the 1983 annual review will be less gloomy. Many thanks to these industrial minerals technical committee members for their contribution to this year's annual review: G. S. Austin, E. S. Montgomery, H. P. Ehrlinger III, D. B. Jorgensen, R. L. Parratt, J. F. Burst, W. Mallory, N. C. Trivedi, and N. K. Olson.

Jan 5, 1983

By V. Miller

C. Hertzler The difference of means of milling characteristics by rock and ore type is tested in a statistical manner in the paper. The null hypothesis being tested is that the sample means of two different populations are equal. It is assumed that the null hypothesis is true when performing the test. Because of this, acceptance of this hypothesis leads to the conclusion that the sample means of the two different populations are equal (or there is no significant difference between the sample means), not that the samples come from the same population, as the author asserts. Rejection of the null hypothesis leads to the conclusion that the sample means of the two different populations are not equal. A P value for this test is reported in Table 2 and is incorrectly identified as the probability that the two samples came from different populations. Actually, what was recorded was 100(1 - p) where p is the usual p-value. The usual p-value measures, on a scale of zero to one, the extent to which the data agree with the null hypothesis. It is the probability of observing a value of the test statistic (Z score) that is at least as contradictory to the null hypothesis as the one computed from the sample data (McClave and Dietrich II, 1982). The author makes no mention of the fact that were these tests to be conducted together, the chances of determining that the null hypothesis is not true when it is really true are increased considerably (Snedecor and Cochran, 1980). This is a multiple comparisons problem and should be dealt with accordingly. The usual methods involve adjusting the level of the individual tests. Table 2 reports an incorrect 100(1 - p) value in the comparison of concentrate grades for other rock types with .30 to .69 % Cu vs other rock types with .70 + % Cu; 86 % should be 98.6 %. Although the article can be criticized for poor data presentation and discussion of that data, the point the author makes regarding the use of these statistics in the ore reserve estimation process is well taken. It does not appear that inaccuracies in the statistical interpretations have hindered that usefulness. References McClave, J.T., and Dietrich, II, F.H., 1982, "Statistics," Dellen Publishing Co., San Francisco and Santa Clara, CA, 766 p. Snedecor, G.W., and Cochran, W.G., 1980, "Statistical Methods," The Iowa State University Press, Ames, IA, 561 p. Reply by V. Miller After reading Ms. Hertzler's comments, it appears that the differences noted were basically semantics. If the paper is ever presented at a math symposium, some changes will be needed. When deciding the best way to present the data to mining engineers and geologists, I decided that presenting the probability that the two sample groupings came from different populations (reverse to the Null hypothesis) was easier and would require less verbage to explain. Hence, the 100(1 - p) difference which was noted. Ideally, I would have preferred to give the recovery and concentrate grades, then comment which ones had significant differences. Unfortunately, Kennecott Corp. did not grant permission to publish the ore body's actual recoveries and concentrate grades. The conclusion is that there was, as expected, a relationship between the ore's geologic properties and the metallurgical properties, and by determining this relationship, the overall mineral model was greatly improved.

Jan 8, 1984

By J. Diao

In order to understand the flotation kinetic behavior of coal and associated mineral matter, a detailed investigation was conducted in a 2-liter Denver flotation cell using three high sulfur coals at three feed sizes, minus 28. 100 and 200 mesh. Subsequently. the flotation products were sieved into different size fractions and analyzed for their ash and pyritic sulfur content in order to delineate the effect of particle size on the rate at which each constituent floats. The results were tested with several different flotation kinetic models and the best model for coal was then chosen. The flotation rate constant and the ultimate recovery calculated from a kinetic model were used to delineate the effect of particle size, collector (dodecane) and frother (MIBC) dosages on the flotation kinetics of coal, ash and pyrite.

Jan 1, 1992

By A. Bruynesteyn

Results are presented from a series of tests designed to determine the beneficial effects of the microorganism Thiobacillus ferrooxidans upon the leaching of uranium from a pyritic uranium ore. Minus 1-inch (-25mm) 0.0711 percent uranium containing, pyritic ore from the Rio Algom Mine at Elliott Lake, Canada was leached with lixiviants of various Eh and pH as well as being subjected to a biologically-assisted ferric iron leach. In all cases, the biological leach gave superior extraction. At pH 1.6, the biological leach gave 100.6 percent extraction and the ferric iron leach 69.8 percent. This improvement in extraction is associated with a relatively rapid biological breakdown of the pyritic material, thus liberating uranium mineralization associated with these pyritic materials and regenerating the ferric sulphate lixiviant in situ.

Jan 1, 1981

By J. D. Miller

The effects of operating variables on water transport to the over- flow, classification of hydrophilic particles and the recovery of copper and molybdenum from a low-grade porphyry ore are presented. Recoveries of copper and molybdenum exceeding 90 percent at a grade of 8 percent Cu and 0.2 percent Mo were realized for minus 100 mesh (1 49 µm) material at retention times of less than 500 milliseconds. The rapid rate of flotation and the effective separation of fine particles for an air sparged hydrocyclone appear to be due to directed interaction of particles in a high force field with numerous freshly formed small bubbles. This interaction occurs in the swirl flow developed by the system and is analyzed in terms of collision and induction time phenomena.

Jan 1, 1982

By D. L. Leach

This report covers laboratory experiments simulating in situ copper recovery from primary sulfide ores in sulfuric acid systems pressurized with oxygen. Copper extraction and acid consumption data are correlated with temperature in the range 50-150°C and quenched pH in the range 0.25-2. Adverse changes in the chemical composition of the leach solution and the formation of gangue-mineral alteration products and secondary sulfides can severely limit the extraction of copper above 90°C. Keeping the pH below the self-buffered value of about 2 and the temperature below about 110°C produces the maximum extraction rates. Forcing the pH too low not only results in excessively high acid consumption but may also limit the extraction of copper at temperatures above 100°C.

Jan 1, 1975

By L. G. Austin

A program has been written to combine the Klimpel-Austin approximate model for liberation in a binary system with a mill model applicable to ball milling. The model applies for a system where the volume fraction of valuable B in host matrix A is small. The mill model is treated as an equivalent series of fully-mixed reactors. Twenty-three classes of material are considered: pure A; pure B; unchanged AB; 0-10.2 B in AB either as surfac- exposed B or as wholly-contained grains of B; 10-20% B in AB, surface-exposed or wholly- contained;etc.. to 90-100% /b ub /ABm syrface-exposed or wholly-contained. Illustrative cases are given making reasonable assumptions concerning relative rates of breakage and primary breakage distribution functions for each class of material.

Jan 1, 1986

By Francis Martino

There are numerous techniques and predictive models used by geologists and mining engineers to define coal reserves and plan mines. The traditional reserve estimation methods commonly used include the isopach, triangular, polygonal, inverse-distance weighting, weighted-moving-average, and trend-surface methods. Geostatistically-based estimation techniques are now being examined by many geologists and mine planners as a superior alternative to these traditional methods. With today's coal mine development costs often exceeding $100 million, and power generation $1 billion, it is logical that a technique that can indicate the degree of confidence associated with a reserve estimate (fuel supply), and thus also the degree of risk, is highly desirable to the coal producer, the financiers, and those desiring to commit to coal utilization under short or long-term supply arrangements.

Jan 1, 1983

By James A. Hartlage

Work on the design of Kelex 100 was initiated several years ago at the Research Laboratories of Ashland Chemical Company. The effort was specifically directed toward the synthesis of an organic molecule which would strongly and selectively chelate copper ions in acidic solutions and form an oil soluble complex which would be stable in contact with an aqueous phase having a pH of approximately one. This publication describes efforts to establish data which could be used to estimate the economic utility of the new reagent. For helping to accomplish these ends, the author wishes to acknowledge the assistance of Mr. A. V. Henrickson of Hazen Research, Inc. for gathering a portion of the data presented here and to thank the copper companies who supplied leach solutions for studies described in this paper.

Jan 1, 1969

By John T. Atkins

Utah International's Shirley Basin Mine is located in the Southeastern High Plains of Wyoming at an elevation of about 7,100 feet. The Shirley Basin area is a gently rolling semi-arid land. The uranium ore zone occurs at depths ranging from 270 to 350 feet below the surface. Overburden consists of poorly consolidated sandstones, siltstones, and clays. Underground mining was begun in 1960. Conditions were made difficult by high rates of water inflow and ground support problems. The development of a solution mining I or in-situ leaching technique allowed underground operation to be abandoned in 1964. Solution mining provided limited production from 1963 to July 1970. In January, 1969, large scale development of open-pit mining was begun. A 1,200 ton per day mill went "on stream" in March, 1971.

Jan 1, 1973

By S. A. Rholl

This paper describes the Bureau of Mines experimental test program which evaluated the effectiveness of using high explosives to enhance orebody permeability. To determine the effect of stress on the fracture network generated by blasting, fifteen single-hole laboratory tests in granite blocks up to .067 m3 (2.4 ft3) in size were blasted using .12 to .20m (4.8 to 8.0 in) of 1.6xl0-3-, 3.2xl0-3-, or 6.5xl0-3-kg (25-, 50-, or 100-grain) detonating cord. In situ stresses were simulated using polyaxial load cells to apply uniform stresses of 1.72xl06 to 1.76xl07 Pa (250 to 2550 psi). All of the blocks were sawed to expose the blast-induced fracture patterns for analysis.

Jan 1, 1991

By Felicia F. Peng

A laboratory fixed bed reactor was used in a study of generating clean thermal energy from coal fine refuse-limestone agglomerates. The major operation variables, including Ca/S mole ratio, agglomerate size, air flow rate and temperature, which affect the rate of sulfur fixation and combustible material utilization, have been examined. The test results show that 70 percent or more of sulfur fixation and 100% of combustible material utilization can be achieved depending upon sorbent, agglomeration, and combustion conditions. The coal fine refuse limestone pellet combustion process (CLC process) requires a simple flowsheet to convert coal fine refuse as a clean solid fuel, while meeting the strict air quality standards for SO2 emissions into the atmosphere. The CLC process is an attractive method to clean up coal slurry impoundments to protect our environment.

Jan 1, 1984

By James McManus

The Lornex mine is located on the Southern slope of the Highland Valley in South Central British Columbia. The large low grade deposit was discovered in 1964; stripping began in 1970, and the mill started commercial production in October of 1972. The orebody is classified as a complex porphyry type having as major constituent quartz diorite, with varying amounts of quartz, grano diorite, and porphyry. A substantial degree of alteration is evident in the host rock, and this became the key factor in major equipment selection for the grinding section. A pilot plant was constructed on the property capable of treating 100 TPD of ore. The ore was mined to be as representative as possible of the entire orebody by sinking a shaft and cross cutting the deposit underground.

Jan 1, 1975

By A. S. Winters

INTRODUCTION Homestake Mining Co. owns and operates the Bull¬dog Mountain mine near Creede, CO. Lead-silver con¬centrates are produced from ores extracted from narrow veins which dip from 1.04 to 1.39 rad (60 to 80°). An¬nual mill throughput is approximately 90 700 t (100,000 st). Ore shoots average 2 m (7 ft) in width and the ore is often weak, muddy, and poorly consolidated. Walls can be either strong or highly fractured along the strike of the veins. Cut-and-fill methods were initially selected to mine the deposit and considerable overhand mining occurred during the initial stages of the mine. Because of un¬stable ground conditions, mine management recognized that other methods would have to be employed. Undercut-and-fill mining was adapted to extract the loose ore from narrow veins and is presently the primary method used. Approximately 80% of the annual pro¬duction is extracted by this system. Loading ramps are positioned along the strike of the vein on 90-m (300-ft) intervals to allow for 45-m (150-ft) long cuts in each direction from the extraction raise. This interval was selected for efficient slusher operation and stope cycling. Loading ramps rather than chutes were selected for ore movement into 4.5-t (5-st) Granby-type cars because of extremely sticky ore conditions. Mined-out areas are filled with deslimed mill sands. Sand is pumped from the mill into two mine storage dams where the water is decanted. Each dam has a 544-t (600-st) capacity. Sand is pumped into the mine at a 30 to 40% slurry through 2195 m (7200 ft) of 51-mm (2-in.) thick wall pipe. When called for, sand is reslurried within the dams to a density of 55 to 65% and pumped to the particular mine opening requiring fill. Dry cement is added to the slurry prior to pumping. The capacity of the mine fill system is 20 t/h (22 stph). DEVELOPMENT Vertical development below the main adit level oc¬curs on 61-m (200-ft) intervals from an internal winze. Main haulage drifts or laterals are 2.7 x 2.7 m (9 x 9 ft) in section and driven parallel to the vein structures. Ground conditions generally dictate that better progress can be realized in the footwall than in the hanging wall. A distance of approximately 21 m (70 ft) is maintained between vein and drift openings to accommodate a smooth 18-m (60-ft) radius curve for crosscuts which are driven on 91-m (300-ft) intervals. Approximately 18 m (60 ft) of tail room is driven past the vein inter¬section to accommodate train loading. Ore and waste are hauled using diesel or electric driven locomotives operating over 18-kg (40-Ib) rail [762-mm (30-in.) gage]. Loading ramps are established above each crosscut¬ vein intersection. Scram drifts are then driven in the vein to determine the width and grade of the ore. If a prior decision to mine the block had been reached, then the scram is driven only far enough [about 12 m (40 ft)] to establish a drawpoint and scraper tail room. In most cases, the drawpoint is a 2.1 x 2.1-m (7 x 7-ft) cutout driven about 2.4 m (8 ft) into the footwall. On the level above, a raise bore station is established in the footwall to one side of the crosscut-vein inter¬section. The station is of sufficient size to accommo¬date a Robbins 41R raise boring machine. Support for the station is accomplished with 1.2 and 1.8-m (4 and 6-ft) roof bolts and wire screening. Once completed, a 152-mm (6-in.) concrete slab is poured and the raise drill set in position to center the 1.5-m (5-ft) diam hole approximately 1.8 m (6 ft) from the vein footwall. A 229-mm (9-in.) pilot hole is then drilled to the scram cutout below and reamed back to full diameter (see Fig. 1). STOPING Initial mining in a typical undercut-and-fill stope be¬gins from the top of the borehole. The initial cut can be either immediately beneath the track level or the old scram drift depending upon the particular situation. Cuts vary but are generally 2.4 to 3 m (8 to 10 ft) high. Support is accomplished by square-set timber, overhead caps, or roof bolts and screening depending upon ground conditions. Once the initial cut is removed, underhand timber and screen is installed and cemented sand fill placed. The stope is then ready for undercut mining. Prepara¬tion work begins with sinking 4.5 to 6 m (15 to 20 ft) on the borehole plus cutting a slot to the vein. As sinking progresses, hanging sets are installed to provide man¬way and service facilities. Hanging sets 1.7 m (5 ft 8 in.) long and 1.6 x 2.2 m (5 ft 4 in. x 7 ft 4 in.) in section are installed with steel hangers similar to shaft sets. Framed 203 x 203-mm (8 x 8-in.) timbers are made up in advance and shipped to the stope during sinking operations. When manway and service facilities are completed to the floor of cut to be mined, generally 3.65 to 4.3 m (12 to 14 ft) below the previous sand fill, the ore is removed to the hanging wall plus a round each way along the vein. Levelers or long stringers are then installed over the raise opening to support the work deck, slusher, and grizzly opening. All waste and ore from the sinking process is fed by gravity to the raise (see Fig. 2). Once the stope is equipped, mining pro¬gresses outward in each direction from the grizzly open¬ing. Stope equipment consists of 907 kg (2000 lb) pull air tuggers, timber skips, 11-kW (15-hp) ventilation fans, 15 to 22-kW (20 to 30-hp) electric slushers, 762 to 1270-mm (30 to 50-in.) scrapers, 67-mm (25/a-in. ) jackleg drills, as well as other drilling and blasting supplies. In narrow stopes [generally less than 2 m (7 ft) wide], no support is required for the underhand caps of the previous fill. Where wide sections in the vein occur, segment braces are installed beneath the caps to insure that the sand support cannot slip away from the hanging wall. Sections as great as 6 m (20 ft) wide and 4.5 m

Jan 1, 1982

By R. A. Bergman, P. G. Thornhill

Development of the Falconbridge Ferronickel Process for treatment of lateritic nickel ores began with laboratory work in 1955, the year the Company commenced exploration of its concession in the Dominican Republic. Process development centred on selective reduction originally directed to producing calcines amenable to ammonia leaching, but later found suited to electric smelting. The reduction step of the process was developed in a 25 tpd shaft furnace pilot plant located in the Dominican Republic, to which was later added an electric melting furnace. After almost 3 years of development work on this scale the plant was expanded to 100 tpd for fixing design data for a commercial plant. This took the form of the 30,000 ton per year Nickel Ferro-nickel Complex of Falconbridge Dominicans, which started up in 1971.

Jan 1, 1979

By S. R. Bryan, J. D. Stephens, D. R. Rothbard

The distributions of gold and associated elements in mineral grains from a pyritic gold ore were studied by optical microscopy, and by SEM and SIMS imagery. Products studied were a sulfide flotation concentrate, a calcined sulfide concentrate, a cyanide leached calcine, and a cyanide leached pressure oxidation residue. Individual pyrite grains are compositionally zoned and solid solution gold occurs in some, but not all zones. Most of the gold in calcined products is present as metallic coalesced grains that are smaller than 100 angstroms. Gold lost into cyanide leached residues from the calcines tends to be in “semi-dense” grains that cannot be distinguished from gold-free grains except by SIMS analysis. The gold in cyanide leached pressure oxidation residues occurs as solid solution gold in unoxidized pyrite remnants.

Jan 1, 1990