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By Manuel Gomez

This, report presents correlative data on ash and sulfur forms for Pittsburgh seam coal from Greene County, Pa. Prediction equations were developed for sulfur form distributions and washability characteristics. The equations developed were evaluated with an independent set of exploration core samples. It was demonstrated-that both sulfur distributions and washability characteristic's can be predicted from exploration core data ahead of mining. This information can improve the efficiency of selective mining techniques for the production of a uniform metallurgical coal product.

Jan 1, 1967

By Henrique D. G. Turrer, Robert F. Winterfield, Jobe E. Wheeler, Marcello R. Cru, Baojie Zhang, Pedro P. Pinto

Efficient particle classification is essential to enhance gravity and magnetic separation stages in one iron ore beneficiation plant. This study investigates the use of high-frequency screens to process finely ground iron ore, focusing on two size cuts: 400 μm and 100 μm. Pilot-scale trials were performed on four ore blends using a full-scale screening unit. A total of 48 tests were conducted, evaluating the effects of panel aperture, solids concentration, wash water, and feed particle size distribution. Results showed that wash water improved efficiency by up to 5%, while a solids concentration near 40% w/w delivered high separation performance with minimal ultra-fines in oversize streams. Apertures slightly larger than the target cut sizes yielded better precision in particle separation. Feed particle size distribution had limited impact on efficiency. The investigation successfully identified optimal screening conditions, including aperture selection, to achieve the most effective separation performance for downstream concentration processes.

Feb 22, 2026

By George L. Atwood

The Monsanto Company's Henry Mine is located 17 miles northeast of Soda Springs, in the southeast corner of Idaho. Ore being mined is phosphate rock which lies in the Permian Phosphoria formation. The ore is the raw material for Monsanto's three furnace elemental "P" plant near Soda Springs. The finished product from the furnaces is elemental phosphorus which is subsequently used in the manufacture of detergents, food preservatives, fire retardants and other consumer items. Mining is by open pit methods using 0-9 crawler tractors to rip and push-load rubber tired scrapers in stripping and mining operations. The equipment spread consists of six D-9G crawler tractors, four Caterpillar 6578 scrapers, eight Euclid S-24 scrapers, one Reich T750 drill and one Caterpillar No. 16 patrol plus various support equipment. Stripping is performed on a 12-month basis, two shifts per day. Mining is performed mainly in the summer months on a one shift basis. The hauling season begins May 15 and ends November 1 due to freezing weather and snow conditions. Annual production is 1,000,000 tons of phosphate rock and 2,500,000 to 3,000,000 yards of stripping. Stripping, mining and hauling is performed for Monsanto by Dravo-Soda Springs, a mining contractor. Dravo-Soda Springs owns, operates and maintains all equipment used in stripping, mining and hauling. The contractor is paid for stripping by the bank yard and for ore mined and hauled on a tonnage basis. The Monsanto elemental phosphorus plant began production in 1952 with ore being trucked to the plant from the Ballard Mine. Haulage was over public roads in 15, 24-ton semitrucks. A private haul road was constructed from the Ballard Mine to the Monsanto Plant in 1958 and 1959 to accommodate trucks with two twenty-yard trailers. This 11/ mile road was first used in the 1959 hauling season. In 1965, five new Kenworth tractors with two, 'forty yard trailers were introduced into the haulage system. These units carried 120 ton payloads and were used until

Jan 1, 1972

By M. S. Prasad

This paper presents the process methods of beneficiating various copper-cobalt and copper-zinc ores mined by Gecamines in Zaire. Geology and mineralization of both sedimentary and vein type deposits are described with special emphasis on the production of oxide, dolomitic and sulfide concentrates. Description of different types of concentrators and their treatment methods are discussed for the production of sulfide, oxide and dolomitic concentrates. Production of concentrates, their consumption and stocks are briefly reviewed. Modernization of the concentrators with substantial investments on grinding, classification, flotation methods and tailings treatments are also discussed to improve the metallurgical efficiency of the concentrators with a view to improve the production of concentrates both qualitatively and quantitatively.

Jan 1, 1990

By K. J. Reid, I. Iwasaki, T. Sabelin

Introduction Occurrences of platinum group minerals (PGM) in mafic and ultramafic rocks in association with Cu-Ni mineralization are well known. Yet for the Duluth Complex, a large body of dominantly mafic igneous rocks in northeastern Minnesota with significant Cu-Ni mineralization (Bonnichsen, 1972; Boucher, 1975; Watowich, 1978; Weiblen and Morey, 1980), little has been published concerning its platinum group element (PGE) contents and mineralogy. In a previous study of massive sulfide rock, sperrylite (PtAs2) was found in nickel arsenide and recovery of arsenide minerals by flotation has been investigated (Iwasaki et al., 1986; Ryan and Weiblen, 1984). Recently, an announcement was made of the discovery of significant quantities of platinum and palladium [peak assay values of 9 g/t (0.27 oz per st)] in an interval of a drill core with disseminated sulfide mineralization 16 km (10 miles) southeast of Ely, MN (Anon, 1985; Sabelin, 1985). This paper describes PGM from the Pt-Pd-enriched zone and evaluates their potential for concentration. Geologic setting The Pt-Pd zone occurs in the South Kawishiwi in¬trusion near its basal contact at 734 to 737 m (2408 to 2418 ft) depth. The South Kawishiwi is one of several minor mafic intrusions that constitute the western margin of the Duluth Complex near Ely, MN (Fig. 1). It is com¬prised of a 70.1- to 350.5-m (230- to 1150-ft) thick sul¬fide-bearing basal unit of troctolite with minor amounts of anorthosite, norite, picrite, and oxide layers overlain by a troctolite unit with interlayered anorthosite (Bonnichsen et al., 1980; Foose, 1984; and Foose and Weiblen, 1986). The basal unit, in which the Pt-Pd zone is found, is markedly heterogeneous in rock type, lacking any coherent internal stratigraphy, and shows evidence of both pre- and post-consolidation faulting (Foose and Weiblen, 1986). Geochemistry Chemical analyses were made on 0.3-m (1-ft) intervals of crushed drill core material (1/4 core, BX size) at 731.5 to 740.7 m (2400 to 2430 ft) depth and the results are presented in Fig. 2. The analytical techniques used were atomic absorption spectrometry for silver, chromium, copper, and nickel; a coulometric titration method for sulfur; and a combined fire assay-DC plasma technique for gold, palladium, and platinum. Figure 2 shows the significant enrichment of platinum and palladium at 734 to 737 m (2408 to 2418 ft) depth; it corresponds to a stratigraphic interval rich in oxide minerals. Average values for platinum and palladium contents over the 3-m (10-ft) interval are 2520 and 2580 ppb, respectively, or 5.1 g/t (0.15 oz per st) platinum plus palladium. Platinum plus palladium contents from individual 0.3-m (1-ft) intervals are as high as 9120 ppb or 9.1 g/t (0.27 oz per st) and average 7.4 g/t (0.22 oz per st) in a 1.5-m (5-ft) interval in the richest part of the zone. The Pt-to-Pd ratio is nearly 1:1 over the entire 9.1-m (30-ft) interval and the platinum and palladium correlate closely with chromium, which ranges from 1.9% to 5.3%. Copper, nickel, and sulfur contents for the drill core interval are also given in Fig. 2. In¬terestingly, the Pt-Pd high corresponds to a depletion in copper, nickel, and sulfur. Mineralogy of the Pt-Pd zone The Pt-Pd zone occurs within an oxide-rich zone or layer that consists essentially of oxide minerals, plagioclase, and olivine. The oxides may constitute as

Jan 1, 1987

By Donald C. Violetta

During 1963 more than 200 million tons of coal was consumed by the public utility industry. Ash residual, from the burning of this coal has caused disposal problems. In order to prevent air pollution, dust collection systems have been installed to suppress the fine ash emitted from boilers before it is discharged into the atmosphere. The collected ash, called fly ash, must be removed from the dust collection facilities and disposed in some manner, usually at a substantial cost to the utility. Many years ago fly ash disposal was not a serious problem because the demand for power was relatively low and coarse coal was burned on stoker type traveling grate machines. The low burning rates utilized on these grates did not produce large quantities of air-borne coal ash while the mechanics of this system of burning resulted in a clinkered residue of cinders. Both of these conditions resulted in the production of only small quantities of fly ash. As the demand for electricity increased, a method of burning pulverized coal was developed and was introduced to the public utility industry in 1918 at Milwaukee, Wisconsin.1 This method of firing was considered a significant advance in steam electric generation as it (1) permitted the use of coals that could not be economically fired by other methods and (2) has generally resulted in improved efficiency, greater generating capacity and lower costs.2 The utilization of larger generating units using pulverized coal substantially increased the quantity of fly ash which resulted in a' more severe disposal problem.

Jan 1, 1966

By C. R. Pelley

This paper is intended to describe the utilization of Western subbituminous coal as used in a utility-sized steam generator. It compares this coal's characteristics with those of a typical Midwestern and Eastern coal as a base and reviews the various design considerations when blends of these coals are used. The effects of blending various low-sulfur Western coals with Midwestern or Eastern coals to achieve a predetermined sulfur dioxide emission and the resulting alteration in the coal characteristics as they apply to utilization are also reviewed. The blending of coals as one means of complying with sulfur dioxide regulations will take on increasing importance as the various deadlines for compliance with new and existing regulations approach. We must first consider the availability of these low-sulfur coals; and insofar as the Eastern low-sulfur coals have been mined for many years, we have focused our attention on the availability of the coal reserves west of the Mississippi River in the United States, as this is the area to which many people are turning in order to meet this requirement.

Jan 1, 1976

By P. E. Lenz

In Nevada, large deposits of diatomite were formed during the Miocene / Pliocene in deep, alkaline lakes. Associated sediments are predominately fine-grained lacustrine with varying fluvial-clastic layers. Tephra layers are common, and probably served as a source of soluble silica for the extensive diatom colonies. Four companies are currently producing diatomite from Nevada deposits. Nevada ranks second to California in annual domestic production. Nevada diatomite is used in filter aids, fillers, absorbents, catalyst supports, and many other products. Economics and mineral reserves indicate that diatomite production in Nevada will be significant well into the next century.

Jan 1, 1993

By Dayu Long, Changhong Li, Yu Wang, Juzhou Li, Yongyue Hu

To address the problem of seepage failure that affects the stability of the fine-grained tailings dam, two kinds of prevention and treatment measures were proposed involving composite flocculants and composite curing agents. The orthogonal test was conducted to determine the optimal ratio flocculant (ORF) and the optimal ratio curing agent (ORCG). Then, numerical models were established to evaluate the effects of the prevention and treatment measures on the seepage and stability of the tailings dam. The results show that the mechanical properties and microstructure of the tailings are significantly improved by adding the ORF and ORCG. Furthermore, numerical studies suggest that the addition of the ORF and the ORCG to the tailings improves the factor of safety of the tailings dam, but the addition of the ORCG also increases the height of the phreatic line. It is also speculated that the addition of the ORF can prevent the formation of through failure zone in the tailings dam. Therefore, the optimal measure to prevent seepage failure in the tailings dam is adding an appropriate amount of the ORF, as it reduces the height of the dam’s phreatic line and lowers the risk of seepage failure. The typical area susceptible to seepage failure can be treated by adding the ORCG. However, it is important to ensure adequate drainage of the area and to maintain the phreatic line at a reasonable buried depth.

Mar 4, 2024

By C. Vass, A. Noble

"The chemical treatment of mine-influenced waters is a longstanding environmental challenge for many coal operators, particularly in Central Appalachia. Mining conditions in this region present several unique obstacles to meeting NPDES effluent limits. Outlets are often located in remote areas with challenging terrain where conditions do not facilitate the implementation of large-scale commercial treatment systems. Furthermore, maintenance of these systems is often laborious, expensive, and time consuming. Many large mining complexes discharge water from numerous outlets, while using environmental technicians to assess the water quality and treatment regime multiple times per day. As an alternative solution, this paper describes the ongoing research and development of automated chemical treatment systems for mine water discharges. The current work explains the development of a bench-scale prototype that has been used to evaluate various monitoring, notification, and control strategies. Results from this testing are compared to present day practices to show the costs and benefits of implementing an automated treatment and monitoring program. INTRODUCTION Background The National Pollution Discharge Elimination System (NPDES), per Section 402 of the Clean Water Act, requires permitting of all point source discharges from coal mines. This program dictates that all water leaving a permitted boundary, whether it be process water or storm water, must leave through a monitored NPDES outlet. These permitted NPDES discharge points, also denoted outlets, outfalls or monitoring points, must meet a rigorous set of effluent limits to maintain compliance. Common effluent limits monitored by coal companies include but are not limited to iron, manganese, aluminum, total suspended solids (TSS), and pH. Raw mine water is often deficient in meeting these effluent limits, and as a result, various treatment processes are required to bring the water into compliance prior to environmental release. Depending on the mining and geologic conditions, water monitoring and treatment programs can constitute a significant operating cost for the mine operator. In Appalachia, acid mine drainage (AMD) constitutes one of the most significant and widespread water quality challenges that must be treated per NPDES guidelines. AMD is a general lowering of water pH that occurs when water and sulfide minerals interact within oxidizing conditions in coal and metal mining. Acid Rock Drainage (ARD) is a similar term that applies more generally to non-mining causes such as highway construction and other large-scale excavations. Today, mine influenced waters (MIW) is becoming more popular as the general term referring to all types of mining-induced water issues. AMD has been researched substantially over the last 30 years, with investigators studying various causal mechanisms and treatment strategies (e.g. Gazea, 1996; Ziemkiewicz, 1998; Skousen, 2000; Kleinmann, 2001; Johnson, 2003; Akcil, 2006; Kalin, 2006). This research has produced several pragmatic results, which provide numerous methods to predict, prevent, and treat AMD."

Jan 1, 2016

By R. T. Webber, D. Moffett, D. W. Averill, G. H. Kassakhian

INTRODUCTION In Canada, radium-226 is removed from uranium mining and milling effluents by the addition of barium chloride to precipitate barium-radium sulphate. Large settling ponds are provided to effect solid-liquid separation. However, the barium-radium sulphate precipitate [(Ba,Ra)S04] is a fine, slow-settling crystalline solid which is not easily removed by simple sedimentation systems. Consequently, settling pond effluents may contain appreciable quantities of particulate or suspended radium-226, depending upon the specific pond system and/or the time of year considered. Also, current technology is expected to be incompatible with ultimate disposal goals since safe recovery of the barium-radium sulphate sludge from the settling ponds may be difficult to achieve (IAEA, 1976). Under the auspices of Environment Canada, a Government-Industry Mining Task Force established a Radioactivity Sub-Group in 1974 to assist in the development of effluent regulations and guidelines for the uranium mining industry. Among the recommendations of the Sub-Group was the need to seek more effective methods of radium removal, including development of physical--chemical treatment systems as an alternative to settling ponds. In response to the Sub-Group recommendations, a bench scale feasibility study began in March 1976 at the Wastewater Technology Centre (WTC) of Environment Canada. The bench scale work was expanded and accelerated in 1977 with financial assistance provided by the Atomic Energy Control Board. The results of this study indicated that coagulation and flocculation of the (Ba,Ra)S04 precipitate greatly increased settling rates (Wilkinson and Cohen, 1977). Subsequently, a research program with a larger scope was initiated in January 1978; this program is jointly managed and funded by government and industry as indicated in Table 1. The principal goal of the Joint Government-Industry Program is the development of a physical-chemical process to reduce the radium-226 content of uranium mining and milling effluents. The following effluent radium-226 activities were established as targets for the program: 0.37 Bq/L (10 pCi/L) as total radium-226; 0.11 Bq/L (3 pCi/L) as dissolved radium-226. Process development work was undertaken at bench scale at the WTC and at

Jan 1, 1980

By N. A. Abdel-Khalek

An experimental study was carried out to compare the results of laboratory column flotation versus conventional flotation for the cleaning of fine coals of different rank and their oxidized samples as well as discrete size fractions. Surface oxidation caused a significant reduction in the flotation yields achieved with both methods. However, the column flotation produced concentrates? of lower ash content. The recovery -size curve achieved with column was similar in shape to that obtained with conventional cell, with superior results in the region of fine fractions. The influence of some column operating parameters was also discussed.

Jan 1, 1990

By M. Adnan Goksel

The hydrothermal agglomerating process described (U.S. Pat. 3,235,371, (1966)) is different from the conventional high temperature processes. Green pellets or briquettes made with mixtures of moist iron ore concentrates and burned or hydrated lime fines are hardened in an autoclave at 198°C and under steam pressure for 1/2 to 5 hours, with or without previous drying. Under the neutral conditions, oxidation of magnetite does not occur as in the conventional processes. The resulting agglomerates are hard and resistant to abrasion and weathering, and they do not disintegrate in the Linder test. The equipment required for the process is not complicated, and favorable operating costs are projected because of the low temperature and efficient heat recuperation. A low capital investment makes the process attractive even when the annual pellet production is below a million tons.

Jan 1, 1968

By William R. Yernberg

Caterpillar introduced 21 new products during Caterpillar Press Days ‘96, held in June at the company’s proving grounds near Tucson, AZ. The event, billed as the largest product introduction ever held by Caterpillar, was attended by press from around the world. Most of the new products were demonstrated at Caterpillar’s Tinaja Hills Training Center in the Sierrita Mountains south of Tucson. Under the scorching Sonora Desert sun, the press was treated to demonstrations of the new equipment. After a display of the equipment’s power and capabilities, the press was given an opportunity to operate the new models and, in some cases, compare them with the old.

Jan 1, 1996

By Kadri Dagdelen

The importance of factoring mining dilution into feasibility stage ore reserve estimates is discussed. An example that relates the issue of mining dilution to the size of smallest volumes that can be mined selectively is given. Relationship between the size of the selective mining units (SMU's) and the variance of their average grade is presented. The concept of variance reduction and its use in factoring mining dilution into geostatistical ore reserve estimation is described. An approach to estimate variance reduction factor by way of conditional simulation is introduced.

Jan 1, 1996

By J. J. DeMull, B. A. Boster

Amax's Sleeper mine uses conventional cyanidation milling and heap leaching techniques to produce gold. Since milling operations commenced in March 1986, average ore processing rates have improved from the design rate of 455 t/d (500 stpd) to more than 720 t/d (795 stpd). Heap leaching was added to the process facilities in September 1986 at a design rate of 1.5 kt/d (1650 stpd) and current heap leaching rates exceed 2.7 kt/d (2950 stpd). Presented are descriptions of the Sleeper processing facilities with emphasis on the crushing and grinding circuits, a review of the selection of crushing and grinding equipment, and a summary of operating performance during the first 16 months of operations.

Jan 1, 1990

By B. J. Scheiner

The Bureau of Mines, U. S. Department of the Interior, is conducting research on a dewatering technique for fine particles generated during the mineral beneficiation process that will recover a portion of the water now being lost using conventional waste disposal methods and significantly reduce the time required for land reclamation. The technique utilizes a flocculant, polyethylene oxide (PEO), that forms strong stable flocs that can be dewatered rapidly on static and/or rotary screens. In field tests, phosphate clay waste has been dewatered from a nominal 2.5 percent solids to greater than 20 percent solids. In other field tests, coal-clay waste has been dewatered from a nominal 3 percent solids to greater than 55 percent solids, in addition, small-scale continuous tests have shown that bentonitic clays can be dewatered from 12 percent solids to 40 percent solids.

Jan 1, 1983

By H. Rush Spedden

A process has been developed and piloted to reject copper, lead, iron, phosphorus, bismuth, antimony, elemental sulfur, and insol from molybdenite concentrate without decomposing the contained molybdenite and rhenium. The impure concentrate is first heated with concentrated sulfuric acid at 260 to 290°C for 1 hour. This bake not only oxidizes the contaminating sulfides, but also activates the concentrate for later rejection of insol by flotation. Two stages of leaching follow the sulfation step. The first is an acidified water leach which extracts the oxidized copper, iron, and phosphorous; the second leach rejects the lead, bismuth, and antimony into a warm concentrated solution of sodium chloride. Final upgrading is achieved by conventional froth flotation. The chemistry of the process as well as results obtained from its implementation in a 5,000 lb/day pilot plant are presented.

Jan 1, 1971

By J. L. Huiatt

The U.S. Bureau of Mines and Great Salt Lake Minerals and Chemical Corp. developed a froth flotation process for concentrating potassium salts from Great Salt Lake solar evaporites containing about 5 percent K. The crude evaporites, which were predominately halite (NaCl) with smaller amounts of kainite (KCl.MgSO4.2.7H20) and schoenite (K2SO4MgSO4.6H20), 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 percent of the potassium could be recovered from the crude evaporites in flotation concentrates assaying about 11 percent K.

Jan 1, 1975

By William C. Cassun

The important role of pit slope design in the overall economics of open-pit mining ventures has long been recognized and well documented.1 2 Any economic analysis of a pit slope design must reflect the trade-off between the benefits and the increased risk of slope failure inherent in steeper slope angles. The conventional practice in pit slope design is based on the factor of safety analysis.' The approach uses average values to arrive at a deterministic solution to slope stability. Past analyses based on this approach generally have had several drawbacks: 1. The risk from slope failure could not be quantified. 2. Only relatively simple pit wall geometries were considered. 3. The analyses were performed solely for the benefits of steeper slopes, seldom discussing the potential costs of slope failures. 4. Analyses were seldom considered in a year-by-year mine planning context. Recognizing that average values do not reflect the uncertainty in geomechanical parameters, a probabilistic philosophy to slope design has evolved. 4,5 However, this approach had not been fully developed to include complex slope geometries and mine planning consideration. In order to develop an integrated, pit slope design procedure, a research project was initiated at the University of Arizona during 1973 in conjunction with the Canada Center for Mineral and Energy Teleology. This research resulted in the development of a probabilistic design procedure. The remainder of this paper gives an overview of the probabilistic design procedure and a discussion of the requirements for the benefit-cost analysis of a pit slope design. Finally, the results of a case study undertaken at an operating mine are given.

Jan 1, 1977