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By S. K. Kawatra, S. J. Ripke

Iron ore concentrates pellets are currently in duration heat hardened (sintered) under oxidizing conditions. Pellets produced from magnetite ore concentrate are completely oxidized during induration and finish as hematite pellets. These pellets then require a blast furnace to reduce the hematite to metallic pig iron. Partially reduced pellets would save energy and fuel during transportation and further reduction. It has been stated in the literature “the most important single factor responsible for the development of strength in fired magnetite pellets containing no additive is the oxidation of the magnetite to hematite and the subsequent recrystallization and grain growth of this constituent.” However experiments shown in this paper reveal that oxidation is not necessary to produce high-strength pellets. Magnetite concentrate pellets developed similar strength when indurated at 1350°C under either oxidizing (air) or inert (argon) atmosphere; the pellets sintered under argon remained magnetite. In addition, magnetite concentrate pellets indurated under reducing (CO) atmosphere also remained as magnetite, but required induration temperature of approximately 150°C less than the other pellets.

Jan 1, 2003

By Mehmet Mesutoglu, Ihsan Ozkan

This study investigates the feasibility of rock bolting support in an underground coal mine gallery with a thick coal seam. The Ömerler underground coal mine working area, owned by West Lignite Enterprise of Turkish Coal Enterprises, was selected for this purpose. Longwall top coal caving is implemented as the production method in the Ömerler underground coal mine. Field and laboratory studies were conducted to determine rock mass and rock material properties, followed by experimental, empirical and numerical analyses based on the acquired data. The obtained design results were evaluated using a resin-grouted rebar rock bolting (RBR) system and steel arch support (SAS) pilot application areas. The numerical modeling results conducted using the fast Lagrangian analysis of the Continua 3D (FLAC 3D v6.0) program indicated less displacement and secondary stress change in the RBR-supported zone compared to the SAS-supported zone. In situ measurements also demonstrated that RBR provided more successful support to the roof during coal production activities. The findings suggest that RBR is a more effective solution when evaluating the feasibility of rock bolting support systems in underground galleries with thick coal seams at the Ömerler Mine. This study emphasizes the importance of more sustainable and safe support systems to enhance operational efficiency in the coal mining industry.

Mar 1, 2025

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 Mustafa Kumral, Samer Hmoud

Effective management of ore loss and dilution is essential for successful grade control and short-term mine planning due to their significant impact on the economic, environmental and technical aspects of openpit mining operations. Factors influencing ore loss and dilution fall into two categories: (1) controllable factors, like mine equipment selectivity and blast design, and (2) uncontrollable factors, such as spatial heterogeneity of ore and waste blocks on a bench. This paper focuses on the second category by applying spatial entropy concept to capture heterogeneity at the scale of selective mining units. In this paper, global spatial entropy is used to assess the impact of spatial heterogeneity between ore and waste blocks on the magnitude of ore loss and dilution, while the local spatial entropy can guide the allocation of blast movement monitoring balls preblast. High values of the global spatial entropy indicate increasing potential of ore loss and dilution, which reduce profit. Furthermore, through a number of case studies, the study investigates the relationship between spatial entropy, cutoff grades, blast movement, dig-limits optimization model running time, and profit. The results show that changes in cutoff grade and blast movement can significantly affect spatial entropy post-blasting and increase ore loss, dilution and profit reduction, revealing the need for controlled blasting at specific bench sections. Additionally, the results demonstrate an exponential increase in profit reduction due to ore loss and dilution with a rising global spatial entropy

Jun 1, 2024

From 2009 to 2012, Aker Wirth, in cooperation with the BGR (German Federal Institute for Geosciences and Natural Resources), developed a concept for the deep sea mining of manganese nodules, which it also examined by means of a profitability analysis. The study was undertaken for the German Federal Ministry of Economics and Technology. The increasing scarcity of resources, coupled with the continuous rise in resource prices, drive the search for new raw material deposits. As a result, the deep sea is coming increasingly under the spotlight. At depths of more than 1,000 m (3,280 ft), large deposits of ores and precious and high-tech metals, such as manganese, cobalt, nickel, copper and gold, along with rare earths (which are important for the high-tech industry), can be found. The extreme deep sea conditions ? enormous water pressure, ice-cold temperatures and complete darkness ? present huge financial and technical challenges to the mining industry when extracting these materials. Aker Wirth has recognized the potential of this important future market early on and advances the development of relevant products.

Jan 1, 2013

By D. M. Range

Exxon is reclaiming an 80-hm2 (200¬acre) Wyoming uranium tailings basin in accordance with US Nuclear Regula¬tory Commission (NRC) requirements. The reclamation plan is designed to control the tailings and their radioactive emissions for at least 1000 years. Rigorous modeling and analyses of potential erosion by wind and water were used in planning. The use of riprap was minimized and the need for long¬term reclamation maintenance was eliminated. The resulting design is simi¬lar to regions stable from erosion cre¬ated by nature. The concepts used in the design and implementation have wider reclamation application. Highland history Exxon operated the Highland uranium operations from 1972 until mid-1984. Open pit and underground operations mined the roll front, sandstone ore body. The 2.5 kt/d (3000 stpd) mill employed conventional crushing, grinding, sulfu¬ric acid leach, CCD, SX, precipitation, and drying circuits. The mill produced 10 Mt (11 million st) of tailings. An¬other operator began commercial scale in situ leach operations in 1988. The tailings basin was constructed by deposition of mill tailings in a natural ravine dammed with a 30-m (100-ft) high by 1050-m (3400-ft) long clay core earthen dam built entirely downstream of tailings. Figure 1 shows the layout of the tail¬ings basin in 1984. Tailings were de¬posited along the dam face on the east side of the basin as well as in the west end and along the south edge until the early 1980s. This left a deep water pond in the middle of the basin. The deep pond was largely filled with tailings during the last four years of mill opera¬tions by the construction of two tailings causeways and the operation of a float¬ing tailings line. The pond water volume was mini¬mized by recycling tailings water to the mill and by operating a 32-L/s (500¬gpm) evaporation enhancement spray system during the last few years of mill operations. Regulatory requirements Several conceptual tailings basin rec¬lamation plans were prepared during the mine life. In September 1984, Exxon began preparation of a detailed plan in response to an NRC request. Water, Waste and Land Inc. (WWL) of Fort Collins, CO was selected to help pre¬pare the plan that was submitted to the NRC in December 1984. The design had to meet the require¬ments found in NRC regulations (NRC, 1985). These were based on regulations prepared by the Environmental Protec¬tion Agency (EPA, 1983) at the man¬date of federal law (UMTRA, 1978). The NRC regulations contain strict primary standards and a somewhat more flexible set of secondary standards. The primary standards include requiring that: • radiation from tailings be reduced to background levels; • radon gas emitted by the tailings be reduced to 20 pCi/m2/s from the surface of the reclaimed tailings pile; • tailings not be dispersed into the environment; and • these conditions endure without routine maintenance for at least 200 years and for 1000 years to the extent reasonably attainable. The lengthy time requirements pre¬sented a large engineering challenge. Radon is a radioactive, chemically inert gas with a 3.8 day half-life. It is released from uranium tailings by the radioactive decay of radium, which, in turn, is a radioactive decay product of uranium. Radon emissions from ura¬nium tailings can be controlled with an earthen cover thick enough that the radon gas decays to its solid, short-lived radio¬active daughters before the radon dif¬fuses through the cover material into the air.

Jan 1, 1989

By Anthony R. Przybylek, John M. Visnesky

INTRODUCTION Thermal dryers have been utilized for reducing moisture of bituminous coal since the late 1930's. Since that time, numerous changes in the coal industry have motivated advances in thermal drying technology. The purpose of this paper is to discuss the changes that occurred to dryer design and focus on current practices of fine coal drying. The major topics discussed in this paper are: - Early Thermal Dryer Design and Evolution - Current Drying System Design and Associated Components - Current Economics of Fine Coal Thermal Drying - Future Markets and Design Considerations EARLY THERMAL DRYER DESIGN AND EVOLUTION Background The practice of drying fine coal with heated air has been performed for more than one half century. Throughout that period, both Metallurgical and Steam coal producers have used this method for reducing the surf ace moisture content of their fine plant products. During the late 1930's and early 19401s, vibrating bed and flash dryers were used to per- form the drying duties. Closed water circuits and vacuum filtration had not yet been integrated into process circuitry and centrifugal dryers were in their early stages of development. During that same period, changes in mining practices were reshaping the distribution of coal to our processing facilities. Plants originally designed for the purpose of processing coarser fractions of coal were being burdened with additional quantities of fine coal. Those plants, unable to recover fine coal, would reject these solids to plant bleed. The higher incidences of stream pollution that resulted from these events prompted clean stream legislation and the need for new, more efficient fine coal recovery circuits. The coal processing industry responded to these needs by integrating closed water circuits and filtration into circuits. The fine coal filter cakes recovered exhibited poor handling characteristics by themselves and often required further dewatering to improve their characteristics. Flash dryers at that time were being used to dry the coarser products. For those facilities operating flash dryers, the solution to this new problem appeared to be simple. The filter cake was combined with the coarser (-1/4") coal and then introduced to the flash drying column for drying. The filter cakes, however, presented significant problems for the flash drying systems. Heavy lumps of filter cake could not be dispersed and conveyed by the rising hot air and as a result would settle out in the bottom of the column. This material, known as "drop-out", would then accumulate and require further handling. The solution to this problem was not an easy one. Sophisticated mixing equipment was employed to condition the feed to the dryers. As attempts to minimize drop-out with this approach failed, the only recourse was to collect the material with conveying systems and recycle it back to the plant or dispose of it. The additional costs of handling this material reduced the attractiveness of the flash systems and opened the door for the development of a new, more cost effective coal drying system. Due to the nature of the problems experienced by the flash systems, fluid bed systems were considered. Fluid bed dryers were being used in other industries at that time but had no operating history in the coal industry. Dorr-Oliver transferred technology from other industries and built the first fluid bed drying system for fine coal.

Jan 1, 1988

By A. Wright

Gold Fields Mining Corporation is presently evaluating gold mineralization beneath thick, wet alluvium at the Chimney Creek Project, Nevada. Contamination of reverse circulation samples has been visually apparent since drilling began in the area. Tests have shown that, while contamination can affect assays, reverse circulation holes appear to define mineralized zones almost as well as the diamond drill holes. To improve reverse circulation sample integrity several techniques were tried including a stiff-mud injection system but with limited success. Diamond core drilling continues to be used in conjunction with reverse circulation drilling to further delineate the mineralization.

Jan 1, 1991

By J. Trimble, B. M. Davis, D. McClure

During initial mining operations at the Colosseum gold mine, discrepancies between the average blasthole grades and mill-head grades were quite large. In December 1987, mill-head grades averaged 39% lower than corre¬sponding estimates from the blastholes. The correlation between average blasthole grade and mill-head grade had an r-value of 0.015. The magnitude of the discrepancy and the lack of correspondence between estimated and actual grades led to a series of investigations of the sampling system, the effect of blast movement of the ore, and the estimation methodology. The discrepancy between blasthole grades and mill-head grades had been reduced by about half on a percentage basis at the conclusion of these studies. Ongoing studies have produced changes that reduced this discrepancy to its present level of less than 10%. Geology The Colosseum mine is within the Clark Mountain mining district about 68 km (42 miles) southwest of Las Vegas, NV in San Bernardino County, CA. Gold mineralization at the deposit is hosted in two 152 m x 213 m (500 ft x 700 ft) interconnected, rhyolitic intrusive breccia pipes that intruded Pre¬cambrian crystalline basement rocks and overlying Paleozoic marine sedimentary rocks. Subsequent erosion (deroofing) of all of the Paleozoic section and a portion of the Precambrian section has exposed the pipes to the present level (probably a significant depth in the original breccia pipe system). Lithologies exhibited within the breccia pipe complexes include rhyolitic felsite, and breccias of felsite and wall/ country rock. Breccia types grade from felsite plus Precambrian granite-gneiss fragments in an aphanitic felsite matrix to crosscutting breccias of felsite, Precambrian granite-gneiss, and Paleozoic sedimentary rocks in a matrix of finely comminuted material of the same composition. All the lithologies in and around the breccia pipes exhibit moderate to intense quartz-sericite-pyrite alteration, with sericite and pyrite the dominant visual alteration minerals. Gold at the Colosseum mine is generally submicroscopic and associated with sulfide mineralization, chiefly pyrite. It occurs as free gold, with minor alloyed silver. It is primarily in contact with pyrite in fractures or along pyrite grain edges. Secondarily, it appears as isolated particles in quartz and other gangue minerals but spatially always close to pyrite, and rarely as particles encased in euhedral pyrite. The pyrite mineralization and minor base metal sulfides occur in three distinct styles: as disseminations, as open space or vein/fracture filling, and in breccia clasts replacing dolomite. The pyrite megascopically ranges from absent to 30% of rock volume within the breccia pipes. Although the gold at Colosseum is apparently spatially and geochemically associated with pyrite, there is only a general proportional relationship between pyrite content and gold content. Sometimes, a high volume percentage of pyrite has only geochemically anomalous gold. This makes visual pyrite only a general grade control tool and not a specific tool. This further makes the ore control process at Colosseum all the more difficult. Initial grade control procedure Initial grade control was based on blasthole samples collected in two stages by an "automatic" sample collection device (rocket sampler). A sample was collected for the top 3 m (10 ft) in a hole. The bag was then changed so that a sample was collected for the lower 3 m (10 ft). Each sample was prepared and fire assayed to determine the grade. Large discrepancies were sometimes noted in the assays of top and bottom samples. The samples were averaged together and a plot was made showing the blasthole locations and their associated grades. Discrimination between ore and waste was based on as-

Jan 1, 1989

By S. D. Thompson, L. Adler

J.R. Sturgul Thompson and Adler (1988) present a new simulator for use in designing conveyor belt systems for underground mines. Their model uses a large Fortran program to do a Monte Carlo simulation. Their approach is similar to that done by other researchers, such as Stanford (1965) and Newhart (1977). All Fortran based simulation models use an imaginary clock that is advanced a time increment. At each advance of time increment the program has to determine what is to take place in the system. Since many things can take place at different places in the system, this approach requires that a large amount of logic is needed at each step of the program. This, then, necessitates a great amount of jumping between different portions of the program. In fact, the Fortran program, CMBCS, takes up to 15 to 20 minutes to run on the mainframe and several hours for the PC. Such execution times are typical of Fortran based simulation studies. There have been other examples of Fortran based simulation studies in mining. Cross and Williamson (1969) presented a model of a working open-pit copper mine. Fosinato (1982) presented a model of the bucketwheel excavators for the large open-pit mines in Southern Victoria in Australia. These excavators discharge into a complex conveyor system, which can be thought of as being similar, but more complicated than the one modeled by CMBCS. All previous simulation studies using Fortran as the source code have several things in common: they all are very lengthy, often being of the order of 10,000 lines of code and take a great deal of CPU time to execute. The program of Fosinato takes more than 4.5 hours to run on a large mainframe. Also not mentioned by Thompson and Adler is the actual time it took to write the Fortran code. This, too, can be considerable, as Fosinato's program took several man years to write and de-bug. In order to change it, as the parameters change or as the mine increases, further large amount of time is needed. An alternate to such Fortran based programs is provided by using a special computer language known as General Purpose Simulation System (GPSS). This language is designed to solve a wide variety of discrete event simulation problems and appears to be particularly suited for mining situations as shown by Sturgul and Yi (1987). Although GPSS was introduced as far back as 1961, it has not been used in mining to any great extent. However, the original form was quite crude and inefficient compared to modern versions. Perhaps with the increasing popularity and availability of these recent versions, this will change. Two such versions are GPSS/H for the mainframe and GPSS/PC for the personal computer. GPSS/H allows for interactive usage, such as allowing input data to be entered via a menu. GPSS/PC has excellent graphical capabilities and one can actually view the simulation in progress. In addition, one can halt the simulation and change values of various parameters. If a particular portion of the simulation is in progress via a concept known as "microwindows." These new versions have greatly increased the potential for using the language for mining situations. Mining applications of GPSS have been given by Sturgul and Yi (1987), Sturgul and Harrison (1987), Sturgul (1987), and Sturgul and Singhal (1988). Schriber (1987) presents a summary of the recent advances in GPSS.

Jan 12, 1988

There are many risks associated with the storage of nuclear or other toxic wastes. One of the major risks is the escape of material from the repository. No matter how well a storage facility is engineered, the potential escape of dangerous material must be studied. This kind of study will include some estimation of the "travel path" of the particles. The travel path of a radio-nuclide particle will depend on the geological and hydrological parameters of the surrounding rocks. These parameters must be estimated from a limited number of samples available within the study area. Any estimation will -of necessity -contain a certain amount of uncertainty or "estimation error". This uncertainty will be carried through to the analysis of potential movement of particles, so that the estimated travel path itself will be subject to some error. This paper discusses an approach which may lead to a quantification of the uncertainty on the predicted travel path.

Jan 1, 1989

By Alan Boughey

The world's largest grinding mill operates at the Cadia Hill Mine in Australia. Svedala Industries supplied the 12.2-m- (40-ft-) diam, 20-MW Cadia gearless, semiautogenous grinding mill (SAG). This was a leap of more than 40% above the largest operating SAG mill. This article reviews the significant items involved in the development, design, manufacture, installation and operation of this mill. Description of the Cadia project The Cadia Hill Mine is located 25 km (15 miles) southwest of Orange, in the Central Tablelands of New South Wales, Australia. The mining lease covers about 3,100 hm2 (7,660 acres) of the Cadia Valley. The Cadia deposit is a large, homogeneous, low-grade, porphyry-type, gold-copper resource. Current reserves stand at 202 Mt (222 million st), grading 0.73 g/t (0.02 oz/st) gold and 0.17% copper. They contain 146.2 t (4.7 million oz) of gold and 347 Kt (382,500 st) of copper. The operation consists of openpit mining and ore processing to produce gold bullion and a gold/copper concentrate. The operation processes 17 Mt/a (19 millon stpy) of ore to produce 91 kt/a (100,300 stpy) of concentrate and 9.1 t/a (293,000 oz/year) of gold. The plant operates two 12-hour shifts a day and had a targeted availability of 94%. The plant facilities include a copper concentrator, o water supply system, an openpit mine, tailings pipeline and dam, and a concentrate pipeline and dewatering facility located in Blayney, about 35 km (22 miles) from the concentrator. Significant features of the process plant include o primary crusher and stockpile, a reclaim conveyor system feeding, the Cadia SAG mill (Figs. 1 and 2) and two 6.7-m- (22-ft-) diam ball mills (Fig. 3). A portion of the gold is extracted in a gravity circuit. The balance is extracted with the copper in the flotation section of the plant. Concentrate is thickened and piped about 35 km (22 miles) to the filter plant and rail loadout facility located in Blayney. Tailings are pumped to a tailings dam and process water is reclaimed from the tailings dam. The water system is designed to control and capture available runoff from local creeks, and to store these waters in dams and weirs. In addition, a major portion of the water for the process plant is treated effluent from the city of Orange and a smaller portion from Blayney Mill selection. The low-grade ore posed a dilemma for Cadia. To achieve a satisfactory economic operation, large-scale processing was required. The selection of the mills, equal to the largest designed and operated at the time, gave a plant configuration consisting of two 11.1- x 5.8-m (36- x 19-ft) SAG mills and three 6.1- x 9.1-m (20 x 30-ft) ball mills. This configuration did not provide the most satisfactory economic model for the project. A review of the single line configuration indicated that a significant reduction in the overall capital cost of the plant could be achieved by increasing the size of the mills and reducing the quantity. Increasing the size of the mills,

Jan 1, 2001

By John J. Stetler

Introduction Samarco Mineracao S.A. is a Brazilian corporation jointly owned by S.A. Mineracao da Trindade (SAMITRI) and Utah International Inc., with 51% and 49% interests respectively. Headquartered in Belo Horizonte in the state of Minas Gerais, Samarco owns and operates an open-pit mine capable of producing 10 Mt/a (11 million stpy) of iron ore; a concentrator rated at 7 Mt (7.7 million st); a 12-Mt (13.2-million st) capacity pipeline transportation system; a 7 Mt (7.7 million st) filtering and pelletizing plant; and a harbor facility. The mine, concentrator, and initial pipeline pump station are located at Germano, near the town of Mariana, about 300 km (186 mi) north of Rio de Janeiro by air. The pipeline terminal, pellet plant, and harbor are at Point Ubu, 70 km (43 mi) south of Vitória the capital city of Espirito Santo State, and about 300 km (186 mi) by air east of the mine. The 396-km (246-mi) pipeline uses 508-mm (20-in.) pipe except for a 53-km (33-mi) section where 457-mm (18-in.) pipe is used to absorb and control slack flow through increased velocity in one of the areas along the route. Iron Ore Slurry The hematite iron ore concentrate produced at the Germano plant is processed by grinding the average 53% run-of-mine itabirite iron ore to -150 µm (100 mesh) for gangue liberation. It is treated further by cyclones to remove the 10 µm (1250 mesh) slimes to prepare the ore for flotation. During the multistage cationic flotation operation, the silica is reduced to make low silica (-0.9%) concentrate for the direct reduction (DR) iron-making process or to make normal silica (-3.5% +2.5%) concentrate for the blast furnace (BF) iron-making process. The concentrate then is reground further to meet the particle size requirements for pelletizing. The reground hematite concentrate is screened in closed circuit with the regrind mills for close control of the maximum particle size, not to exceed 3.5% +75 µm (200 mesh) for pipeline wear protection. Thickened concentrates at 70% solids are pumped to four 5-kt (5,510-st)(dry) agitated slurry storage tanks that hold the feed before it goes to the first pipeline system pumping station. Pipeline Capacity The pipeline system was designed to transport 12 Mt/a (13.2 million stpy) at 65% solids with an average minimum safe operating velocity of 1.5 m/s (5 fps). The pipeline will transport 5 Mt (5.5 million st) in 1982. The system, therefore, is moving the slurry in 13-kt (14,330-st) batches, dispatched at intervals to move 20 kt/d (22,050 stpd). In operation, the pipeline has exceeded its 93% availability design factor. Pipeline Facilities Pump station No. 1 is located at the Germano minesite 1 km (3,330 ft) above sea level. It has the following principal equipment: • four 5 kt (5,510-st) (dry) agitate slurry storage tanks; • two centrifugal charge pumps (one stand-by); • a 217-m (712-ft) safety loop; • seven mainline positive displacement triplex plunger pumps (189 m3/h, or 832 gpm, at 14.9 MPa, or 2,162 psi, maximum output pressure); • four corrosion-erosion spools; • a control and command center. Operation The output at pump station No. 1, with 66% solids, is pumped 151

Jan 10, 1982

By D. Leppert

Clinoptilolite zeolite may offer attractive, cost-effective alternatives for remedial cleanup of old minesites, smelters and other sites with heavy metal contamination. Previous research demonstrates the relatively strong affinity of clinoptilolite for numerous metals and clinoptilolite displays a particularly strong affinity for lead (Loizidou and Townsend, 1987), one of the most toxic and problematic heavy metals. Since clinoptilolite can also selectively absorb some radionuclides and organic compounds, it may offer practical solutions to a wide variety of environmental problems.

Jan 1, 1989

By J. J. Antony

The Cactus Gold Mine is a joint venture comprised of Middle Buttes Partners Ltd. (CoCa Mines, Inc.--operator, T. Congdon and W. Carey) and Compass Mining, Inc. The project includes 2.4 sq. miles of contiguous mining claims and leases in the Mojave Mining District, Kern County, California, approximately B5 miles northeast of Los Angeles. The property is readily accessible year round by all-weather roads and is within 6 miles of the main line of the Southern Pacific Railway.

Jan 1, 1988

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 Kevin Schraden, Andrea Pezzi

environmental and social/political changes the world is currently experiencing. Three of major challenges facing the mining industry today are: water scarcity, unpredictable weather conditions that generate a risk of tailings dam failures as well as the correlating social pressure and government regulations that prevent the development of new or expanded tailings dam projects. It is estimated the total number of active, inactive and closed storage facilities globally is over 8,500 containing over 217 km3 of tailings (46% are from copper mining)1 which at some point will need to be addressed. Technologically, there are several solutions for tailings management, including conventional ponds, thickened tails, paste backfill and filtered dry stack tailings. The use of pressure filters, for tailings dewatering is currently being evaluated and often selected as the preferred technology where dry stacking is the chosen solution for residue disposal. This is mainly due to the pressure filters ability and flexibility to address and achieve the current challenging aspects of tailings processing including: • Extremely high capacities of >80,000 dry tons/day • Lower ore grades which require finer grinding and result in a higher amount fine fractions and clay content • Residual moisture targets required for the geotechnical specifications for dry stacking

Feb 1, 2024

By W. E. Benton

The name Perlite applies to both a hydrated volcanic glass containing from 2% to 5% combined water and to the lightweight material that is produced from expanding the sized volcanic glass. Upon heating, Perlite expands to about twenty times its original volume which results in a material of low density, low thermal conductivity, and high sound absorpt'ion. Perlite ore is mined by open pit methods and is crushed and sized to provide specific gradations for expanding into products used in acoustical tile, built- up roofing, filtration, lightweight concrete and horticultural applications, among others. In 1983, the U.S. production of crude Perlite came from 12 mines and 6 western states, with apparent consumption and production estimated to be about 6% lower than 1982. The average price per ton was slightly higher than in 1982.

Jan 1, 1984

By Sam H. Patterson

Australia has led the world in bauxite production since 1976, and only the Republic of Guinea has larger bauxite resources. India has a small integrated aluminum industry, and deposits containing more than 1 Gt (billion metric tons) of bauxite were discovered in the 1970's. The Peoples Republic of China has large bauxite resources scattered in several regions, but it is a monohydrate type that is difficult to process for aluminum. Indonesia and Malaysia both have long histories of bauxite mining, and both have limited resources of high-grade bauxite. Districts containing bauxite resources in the 10 Mt to 150 Mt (million metric tons) range are scattered in countries along the eastern rim of the Pacific Ocean, on islands in this ocean, and on the western and northern rims of the Indian Ocean.

Jan 1, 1982

By Randhir Sehgal

Heavy media processes have been used for cleaning +28 mesh coals for a number of years, Two applications, i.e., the vessels and the cyclones are widely used in the U. S. for this purpose. The basic reason for the popularity of these systems is their extremely high efficiencies for particle sizes above 28 mesh, However, application of heavy media technology to fine coals below 28 mesh, has been limited due to the following reasons: o Inability to make size separations effectively at sizes below 28 mesh o Necessity of feeding large quantitites of magnetite to the separators o Difficulty of recovering magnetite in the presence of large quantities of coal o High levels of coal contamination in the recovered media, leading to high circulating loads and ineffective control of media gravity

Jan 1, 1981