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By Soner Top

In this study, a process to separate manganese and iron from manganiferous iron ores by reductive acid leaching followed by magnetic separation was conceived and experimentally tested. In the leaching process, sulfuric acid was used as lixiviant and oxalic acid was used as reductant. The experimental results showed that the manganese and iron separation was optimal when the concentrations of the sulfuric acid and oxalic acid were 0.75 M and 30 g/L, respectively, at a temperature of 80 °C, a solid/liquid ratio of 67 g/L, stirring speed of 400 rpm and leaching duration of 60 min. Under these conditions, 90.49 percent and 6.78 percent of Mn and Fe were dissolved, respectively, from the ore sample in a size fraction of smaller than 106 µm. It was determined that the leaching of manganese from the ores was a second-order reaction with activation energy, Ea, of 53.38 kJ/mol. The leaching residues obtained under the optimal conditions were subjected to wet high-intensity magnetic separation tests to recover the remaining iron content. This separation process produced a concentrate containing 56.20 percent Fe and 1.79 percent Mn with iron and manganese recoveries of 56.83 and 66.73 percent, respectively. To the best of our knowledge, this is the first time that a magnetic separation process was used on a residue obtained from the reductive acid leaching of manganiferous iron ores to recover iron.

By Michael M. Kimberley

Stratiform uranium deposits occur in Canada along the north shore of Lake Huron near the base of the Middle Precambrian Huronian Supergroup. Outcrop of the Buronian Supergroup extends for over 400 km from Sault Ste. Marie in the west eastward through Sudbury and then northeastward to Cobalt. Maximum stratigraphic thickness exceeds 5 km. The predominant Huronian rock type is metamorphosed sandstone (quartzite), with sub- ordinate argillite, conglomerate, mixtite (possibly tillite), volcanic rocks, and minor limestone, Much of the quartzite is highly feldspathic. Substantial uranium ore was first discovered in Huronian rocks in 1953 and has been exploited continuously since the initial mine-mill plant was constructed at the Pronto ore body, 30 km south of the present townsite of Elliot Lake. Only one ore body remains in production, however. It is shared by Rio Algom and Denison Mines 14 km north of the town of Elliot Lake.

Jan 1, 1980

By J. M. Marek

The determination of the proper cutoff grade for an open pit metal mine is based on economic criteria and the operating philosophy of the mining company. Previous papers have dealt with the balance of a fixed mine production capacity versus economic return. Mine production capacity is often controlled by the amount of mining equipment and the working area in the pit. A complete cutoff grade strategy is a balance of economic return of mining higher grade material versus the capital costs of additional production capacity and the operating cost of moving the waste and low grade material to assure proper ore release. An approach to the problem is presented which is based in part on work by previous authors and provides guidelines to proper cutoff grade strategy encompassing the time value of money and variable commodity prices.

Jan 1, 1985

By H. G. Naidu, Raj K. Singhai, S. P. Singh

Introduction The economics of open-pit mining de¬pend heavily on the cost of overburden removal. It is necessary to maintain a pit slope as steep as possible. Steeper slopes lower the volume of overburden to be handled, thus lower the cost of material handling. There is, however, a limit to which the pit slopes in any given situation can be steepened. Steepening beyond this point could jeopardize the pitwall stability, ren¬dering it unsafe. It might even result in curtailment of pit production. Several factors affect the stability of pitwalls. Some are natural, such as topogra¬phy and regional and local geology. Others are man made, such as pit geometry, ground water management, equipment, drilling and blasting, and the type of excavation. Greater control can be exercised in the latter with proper planning and by following a set of procedures based on field experiences. This article deals with one such control¬lable factor - drilling and blasting. The eco¬nomics of steeper pit slopes is discussed, followed by a record of drilling and blasting practices in western Canadian surface mines. Microprocessor-based instrumentation and its application to blasthole drilling is also described. Economics of steeper pit slopes The economics of steeper pit slopes is not a new issue. It should be reinforced, though, that pit slopes must be as steep as possible to ensure maximum recovery of minerals at a minimum stripping ratio. For operating, economic, and safety reasons, different pit slopes are maintained during different phases of pit development. Figure 1 shows a hypothetical ore body, cylindrical shape, with a radius of 225 m (740 ft) and located 30 m (100 ft) beneath the surface. The mining operations are planned at two different slope angles of 25° and 40°. The ultimate pit depth was established on the basis of the criterion of "break even pit." The variation of the stripping ratio at differ¬ent mine levels is shown in Fig. 2. Table I summarizes the eco¬nomics involved for two pit slopes, 25° and 40°. Figure 3 presents the profits during the life of the mine for these two pit slopes. This study shows that steeper slopes for an open pit operation are far more advanta¬geous than flatter slopes. But, what is the optimum slope at any one point in the life of a moving pit? And how can the stable slopes be maintained over the short- and long-term? Canadian blasting practices A survey of blasting practices in western Canadian surface coal mines revealed that, at most mines, hard consolidated overburden is drilled and blasted. Blastholes are drilled with rotary crawler-¬mounted rigs using bits in the 220 to 310 mm-diam (8.5 to 12 in.-diam) range. Typi¬cal hole depths are 12 to 18 m (40 to 60 ft) to accommodate the pit bench heights of 12 to 15 m (40 to 50 ft). For a 12-m (40-ft) high bench, a typical hole will be 14.6 m (48 ft) deep including the subgrade drilling of 2.6 m (8.5 ft). Half of this hole will typically be a powder column and half will be stemmed. For a 15-m (50-ft) bench, a typical hole will be up to 18 m (66 ft) deep including the subgrade drilling. Drilling patterns are usually 9 m (29.5 ft) spacing and 9 m (29.5 ft) burden. De¬pending on local geology and hole diameter selected, drilling patterns of 8 x 8 m, 8.5 x 7.3 m, and staggered patterns of 11 x 5.5 m

Jan 1, 1989

By Dean Workman

Located in downtown New York City, this project necessitated an innovative conveyor design for successful muck removal. This paper will discuss the technical challenges and engineering solutions of the muck transport system which consists of every commonly recognized version of a belt conveyor. The elaborate system includes a TBM conveyor integral to the back-up system, a wall-mounted extensible conveyor, a crown-mounted transfer conveyor, a floor-mounted fixed-length conveyor, a pocket-belt vertical conveyor, three cascading overland conveyors, and one radial stacker. The most challenging design element was an over-land conveyor featuring a box truss for muck transportation over Northern Boulevard while simultaneously going under the Queens Plaza railroad viaduct. An overview of the project is provided along with the following topics: developing final plans from the proposal concept, interfacing with existing works, communication with the joint venture, and complying with local regulations.

Jan 1, 2008

By M. S. Klima

Tests were conducted to investigate the potential for recovering anthracite from preparation plant tailings. Samples were collected from tailings streams or slurry ponds of several anthracite cleaning plants. A washability analysis was conducted on each sample. The use of gravity-and surface-based cleaning processes for treating the 28x200 mesh tailings material was examined. Simulations were performed for various cyclone staging arrangements using published partition data for both water-only and dense-medium cyclones. Froth flotation testing was done on both precleaned anthracite and tailings material. Grade-yield curves were determined for various conditions.

Jan 1, 1995

By William W. Kephart

Brockway Glass Company, Inc., manufactures glass containers at fourteen plants located in nine states. The relatively high place value of the major glass batch components dictates that raw material supplies be obtained from numerous sources in the United States and Canada. One objective of our central composition control program is to arrive at a glass composition with similar physical properties at each of our plant locations. This requires that adjustments be made in the batch compositions to compensate for chemical variations from one raw material source to another. Minor chemical variations will occur in sand, limestone, and dolomite from one source to another. Major variations are present in the alumina sources because of the diverse rock types utilized.

Jan 1, 1973

By Jeffrey A. Pontius

The Cresson gold deposit is in the southcentral portion of the Cripple Creek mining district in central Colorado (Fig. 1). The deposit is hosted by a series of 32 Ma (Kelley, 1995) felsic to mafic, alkaline volcanic rocks that were deposited in a diatremal environment. Mineralization in the Cresson deposit is structurally con-trolled. It occurs as disseminated and vein type. Gold deposition is closely associated with potassium-pyrite metasomatism and quartz-telluride vein development(Pontius, 1992). The Cresson deposit is owned by the Cripple Creek and Victor Gold Mining Co. (CC&V), a joint venture between Pikes Peak Mining (67%) and Golden CycleGold (33%).

Jan 1, 1996

By John E. Roberts

People are not a novelty. But the phenomenal numbers thereof, who have approached, settled, and been born in California in the past twenty-five years, are such a problem as to be a novelty in the broadest sense of the word. California, with less than 7,000,000 people in 1940, and over 19,000,000 in 1966. In 5 1/2 years, (1960-65) the southerly 14 counties of California, which together comprise one-half the land area of our state, had an increase in population greater than Colorado, or Arkansas, or Oregon, or any one of 19 other states.

Jan 1, 1967

By J. R. Carr

A crucial concern when implementing computer algorithms for geostatistical analyses on microcomputers is speed of execution. Kriging, in particular, typically relies on a Gauss elimination algorithm to solve for weights. Because such an algorithm is required for each estimate, the solution for weights can result in very slow program execution speed on a microcomputer. One approach to enhancing the efficiency of Gauss elimination is demonstrated herein. The upper triangle plus diagonal of the intersample covariance matrix is used in a modified banded Gauss elimination algorithm. Results show that such an algorithm yields approximately a two-fold reduction in execution time for kriging when the number of nearest neighbors used for estimation is large.

Jan 1, 1990

By C. Sampaio, C. Petter, M. Veras

"INTRODUCTION Use of sensor technology applied to the gravimetric beneficiation is intended for the selective preconcentration or separation of various minerals. Sensor Based Sorting (SBS) uses sensors to identify the particles and separate them by a mechanical process (flap) or pneumatic ejection valves (Wotruba et al. 2014). The principle of the SBS technology to separate mineral particles consists of three phases which cover feed,detection and separation. (killmann, 2009 Apud nienhaus et al. 2014). The applicability of the technique in an early stage of beneficiation enables free particles and poor grade to be a part of the process. The mass of tailings to be processed can be removed at the initial stage to save energy, water and equipment maintenance (Wotruba et al. 2014). Lindner (2010) suggested that the use of SBS technology would only will be possible if there exsists an electromagnetic interaction in all the wave spectra interacting with the solid ore. The interaction of electromagnetic waves depends directly on the characteristics of the material. The material receives a spectrum of electromagnetic waves by irradiation and this results in a reduced wavelength with differing types radiation such as fluorescence, reflection, remission, absorption, transmission and transmission. The frequency of the wave spectrum varies between 1000 and 1020Hz, illustrated by the data in Figure 1, which shows that each type of sensor emits a wavelength inversely proportional to the frequency ranging from 10-11pm to 104km, and the frequency that varies between EHz 103 and 1020kHz. The difference between the electromagnetic spectrum variations can only be visible to the human eye at wavelengths with intervals between 380nm and 780nm (Wotruba et al. 2014). One of the main advantages of operating with imaging sensors is the versatility of the images which can take on zero, one, two or three dimensions shown below as a point, a line, an area or a spatial image, as illustrated in Figure 2. According to Schubert (1988), to separate a mineral particle it is necessary to follow the characteristics of mineralogy, physics and chemistry with or without use of water. The use of new technologies can help reduce operating costs and improve utilization of mineral resources."

Jan 1, 2016

By R. L. Virta

Four companies mined ball clay in four States during2004. They were H.C. Spinks Clay (owned by Franklin Minerals) Kentucky-Tennessee Clay (owned by Imerys),Old Hickory Clay and Unimin. Production was estimated to be 1.46 Mt (1.61 million st) valued at about $56.9 million, based on a preliminary survey of the ball clay industry. In 2003, production was 1.31 Mt (1.44 million st) valued at $56.2 million. Tennessee was the leading producer state with 61 percent of production. It was followed by Texas, Kentucky and Mississippi.

Jan 1, 2005

By R. Du Plessis, J. D. Miller, D. G. Kotlyar

The low potential hydrophobic state of pyrite in amyl xanthate flotation with nitrogen is of particular interest with regard to the N2TEC flotation technology currently being used for the recovery of auriferous pyrite at Newmont’s Lone Tree Plant in Nevada. The N2TEC system had been found to operate satisfactorily, but cyanide in the flotation mill water appeared to be responsible for a loss in pyrite recovery. This supposition was confirmed with laboratory experiments and a program was initiated to study flotation chemistry variables by electrochemically controlled contact angle measurements. Experimental results show that activation of pyrite in such cyanide solutions can be achieved more effectively with lead than with copper. Subsequently, based on these fundamental studies, significant improvement at the Lone Tree Plant was achieved by lead activation, in which case the recovery increased to expected levels. The effect of activator is particularly significant not only with respect to pyrite depression by residual cyanide, but also with respect to collector (PAX) consumption and the initial state of the pyrite surface. Each of these issues is being studied in detail and research progress is discussed herein.

Jan 1, 2002

By Jean-Michel M. Rendu

Increasingly, geologists use geostatistical methods to evaluate drilling programs. The complexity of the methods varies with the nature of the available sample information, the objective of the statistical analysis and the geologist's experience. Occasionally, estimating the average sample grade is all that is required. In other instances a semivariogram calculation may be considered. If sufficient data is available, the geologist may wish to estimate blocks or panels of complex shape, using the kriging method of evaluation. Historically, the use of geostatistics by geologists has been limited by a number of factors, most significant being the difficulty in accessing and using computer programs needed to complete the necessary studies. Geologists are often working in remote areas from which electronic communication is unreliable, if not impossible. Even when access to computer programs is technically possible, the costs incurred in accessing and using these programs are often considered prohibitive. In many instances using the programs requires an understanding of computer systems far beyond that which a field geologist can justifiably be interested in acquiring and maintaining.

Jan 1, 1984

By Michael R. Walsh, John H. Wright, Eivind B. Jensen

A tunnel boring machine (TBM) was designed and built at the Cold Regions Research and Engineering Laboratory in 1992 for use in Antarctica. The TBM system was tested at the South Pole where a 120-m long tunnel was bored to prove the concept. Modifications were made to the system resulting from this work. In 1999,boring of a 600-mutilidor began as part of the new South Pole Station. Tunneler hydraulics reliability problems prevented the completion of the tunnel before the end of the short construction season. This paper discusses the design and operation of the TBM, proposed solutions to the problems encountered, and current operations of the system.

Jan 1, 2001

By Rod Heath

The Saskatchewan industry has the advantages of an ore body of a size and consistency unmatched anywhere else in the world, and large, efficient mines with production costs that compare favourably with those in other countries. On the minus side, Saskatchewan is remote from most major markets, and producers face a distribution challenge to reach the offshore markets which is the main focus of this presentation. One of the characteristics of the world potash business is that all of the major sources of production are located in the developed areas of the world, while there is no significant production in any of the highly populated developing countries. The major potash producers are located in Canada and the United States, in West Germany, France, Spain and the U.K., in Russia and East Germany, and in Israel and Jordan. There is currently no significant production in any of the countries in Asia (with the exception of Jordan where a plant is currently in the start-up phase), Africa, Latin America or Oceania, and all of these areas must, therefore, import a11 of their potash requirements from the producing countries.

Jan 1, 1985

By Peter G. Chamberlain

Domestic mining will not produce a significant portion of the manganese consumed in the United States unless new mining and processing techniques are implemented that will enable domestic lean ores to economically compete with rich foreign ores. One mining-processing technique that appears promising for certain deposits is in situ and/or heap leaching with aqueous SO2. Since leaching offers economic advantages of low capital costs, low operating costs, and short start-up times as well as environmental and safety improvements over conventional mining-processing, the Bureau of Mines studied the leachability of manganese deposits from several districts. Manganese oxides, hydroxides, and carbonates were leached to 80 to 90 pct recovery in 7 to 40 days of leaching with a solution containing 5 pct SO2 in column leaching tests. Preliminary cost analyses of surface mining-heap leaching, block caving-in situ leaching, and shrinkage stoping-in situ leaching indicate that the products of manganese leaching may be economically produced for the chemical and electrolytic manganese industries. The easiest marketable product to obtain would be a manganese sulfate commonly used as a fertilizer amendment. Manganese carbonates and hydroxides could also be precipitated from the pregnant leach solutions.

Jan 1, 1983

By Henryk Bystron

The underground fire areas are formed as a result of failure of an active fire extinguishing in mines. Such areas bring about a fire hazard for the work safety and for the mining operations. For liquidation of fire areas it is necessary to create conditions under which a possibly small amount of air is flowing to the fire. To achieve this end, isolation methods are used which consist in erection of full stoppings of various types, sealing practically the fire areas. These methods usually fail in oases when the coal seam is thick and the fire area is situated between the intake shaft and the outtake shaft in which a high-pressure fan is in operation. In such oases ventilation methods based on the balance of aerodynamic potentials in fire areas are to be used. For a very long time the isolation and ventilation methods have been developed in the world mining, No sooner than since 1980 the aerodynamic potential has been used in ventilation methods [1,2,6]. So far, the aerodynamic potential was called in different ways and was determined by different formulae which gave its numerical values more or less approximative. Nevertheless, even in situations featuring by a lack of sufficient theoretical bases of the aerodynamic potential the ventilation methods brought considerable profits in the field of safety in mines and of economics in mine ventilation [1-6,9,10]. The increase in the depth of underground mines and their equipment with expensive machines (apparatus) brought about that the Central Mining Institute at Katowice undertook research works which are crowned by the theory of the aerodynamic potential in mine ventilation networks [7]. The use of this theory in the mentioned ventilation methods will increase their effectiveness. Since the theory will be published in the III rd quarter of this year, we cite in the present paper the formulae and relations associated with this theory which are indispensable to present the balance of aerodynamic potentials in the fire areas in their control by ventilation methods.

Jan 1, 1980

By Bernard L. Merlo

INTRODUCTION Modern mine management is constantly striving to develop improved mining methods which will not only increase productivity, but will also provide a safer work environment and make the most efficient use of our natural resources. Progress can be seen in this area by tracing mining methods through history. Work that was originally done with hand tools progressed with the age of mechanization to conventional mining, matured to continuous and longwall mining, and the advent of short¬wall mining, a combination of continuous and longwall, has recently been seen. This new method of shortwall mining provides a promising method for safe and ex¬pedient coal extraction. This study focuses upon the planning, design de¬velopment, and operation of the first producing short¬wall unit in the US which was operated in the Pittsburgh coal seam in northern West Virginia. MINE DESCRIPTION The mine is located in northern West Virginia at Grant Town in the heart of the high volatile Pittsburgh coal seam. Originally opened in 1901, the mine has be¬come one of the largest in the US, encompassing an area of approximately 103.6 sq km (40 sq miles). The seam being mined was the high volatile Pittsburgh seam which ranged in thickness from 1.8 to 2.96 m (6 to 91/2 ft). Three types of mining methods were employed: con¬tinuous mining, longwalling, and shortwalling, with each type accounting for 50%, 30%, and 20% of total pro¬duction, respectively. A total of 630 contract and super¬visory personnel were employed to produce the mine's designed annual tonnage of 1.8 Mt (2 million st). There were ten producing sections, eight continuous miners, one longwall, and one shortwall, operating a total of 22 unit shifts per day. Eight shafts, varying in depth from 80.2 to 194.2 m (263 to 637 ft) provided entrance to the Pittsburgh coal seam. These shafts were used for ventilation, portaling, coal hoisting, supply, and emergency escape hoisting. Ventilation was achieved by means of three exhaust fans which provided 707.9 m3/s (1.5 million cfm) of air to the mine and removed approximately 141 584 m3 (5 mil¬lion cu ft) of methane per 24 hr period. The room-and-pillar method was used for secondary recovery with designed barrier blocks between panels of 61 m (200 ft) in width. A double row of blocks was left around each gob area to provide efficient bleeder action

Jan 1, 1982

By Ted H. Eyde

Specialty sand and bentonite are produced in the plateau province from the Bidahochi Formation near Sanders, Apache County, Arizona. It appears that deposits of both sand and bentonite formed during the Pliocene in fluvial and lacustrine environments, respectively, along the west side of the Defiance Uplift. The bentonite is an alteration product of airborne vitric ash that fell or was washed into lakes. In the Cheto district, the individual bentonite horizons, which range from less than a foot to more than 10 feet thick, are restricted to the position of the medial volcanic member. Bentonite production began at the Allentown mine in 1924 and the Chambers mine in 1926. In 1933 strip mining of the extensive deposits in the Cheto district commenced. Production increased each year to a peak of 270,000 tons in 1957. As a result of the introduction of synthetic zeolites into petroleum refining, production declined. About 40,000 tons of bentonite are currently produced from the district each year. Bentonite is shipped out of State for processing into desiccants, thickeners, and acid-activated clay products. The specialty sand was derived from Permian sandstones and deposited by streams draining the Defiance Uplift. Although stream channels are usually restricted to the upper member of the Bidahochi, stratigraphically above the bentonite horizon, a few actually cut through the bentonite. The well-sorted sand, suitable for use as a proppant in hydraulically fractured oil and gas wells, is localized in elongate lenses along the margins of the channels. These lenses range from 5 to 50 feet in thickness and often extend thousands of feet in both width and length. The sand is about 97 percent silica, has a yield of 40 percent in the minus 20 plus 40 mesh size fraction, and has a roundness of 0.6 to 0.7 on the Krumbein scale. Production of this sand began about 1961. All of the current production, which amounts to about 40,000 tons per year, is sold in the petroleum-producing area near Farmington, New Mexico.

Jan 1, 1987