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By L. W. Pommier

In the field of minerals processing the concentration of valuable minerals normally takes place in suspensions of ore particles in liquids. The fluid resistance offered by Newtonian liquids to the movement of free settling spheres for the limiting cases of creep flow and turbulent flow is well described by Stokes' and Newton's Laws. However, these equations are not applicable to non-Newtonian fluids, since the relationship between applied shearing stress and resulting rate of shear is a non-linear function, and more than one viscosity parameter is necessary to describe their rheological properties. For this reason, the term "apparent viscosity," which has been commonly used to explain the flow resistance of the non-Newtonian systems including suspensions, may be misleading. A rigorous analysis of the mechanisms of momentum transport of solid-liquid systems is complex indeed. There are many variables that affect the properties of such systems, and precise limitations and assumptions are necessary in order to produce quantitative relationships for predicting their behavior. Very dilute suspensions of spherical particles in Newtonian liquids of identical density as the solids have been examined in the past (12); however, the study of suspensions in which there is a considerable difference in density, between the liquid and the suspended solids has been neglected.

Jan 1, 1968

By J. A. Procarione, M. K. McCarter, Bradley C. Paul

A numerical technique is described which can be used to model the quantity and distribution of air flow through broken rock. The method can also describe gas concentration and temperature distribution. A computer code implementing the numerical approach was developed to assist in determining the feasibility of a proposed modified in-situ leaching method for extracting copper from sulfide porphyry deposits. This mining method requires air to be forced through cells of rubblized ore. The flow serves to supply oxygen for the leaching reactions and to dissipate heat that would other-wise build due to the exothermic reactions. The method substitutes the Carman equation into the Laplace equation. The solution of the resulting non-linear formulation is done on a finite difference grid using SOR with iteration improvement of coefficients. This procedure yields the required air flow pattern and pressure drop. An equation for advection with absorption is then solved over the same grid using explicit schemes to yield oxygen concentration. Finally, a similar technique is applied to the convective heat transfer formula to yield the temperature profile. The model is potentially applicable to such mining problems as the prediction of leakage through caved ground, the prediction of methane concentration in gobbed ground, and the prediction of heat build-up in caved workings.

Jan 1, 1989

By Steven R. Wardwell, Gerald G. Stokes

INTRODUCTION A compaction grouting program to control ground loss, and to minimize surface settlement, was developed for the West Tunnel of the Phoenix Storm Drain Tunnels. The West Tunnel was 4,200 m long, 7.56 m in excavated diameter, and was excavated with shallow cover through the central Phoenix business district. The tunnel could not have been successfully excavated without the compaction grouting program. HISTORY After heated political debate over a ten year period, the Arizona Department of Transportation (ADOT) was granted approval to build the Papago Freeway, an extension of Interstate 10, through central Phoenix. To comply with environmental constraints, a portion of the Freeway was depressed, requiring an excavation 10 meters in depth, thus intersecting surface run-off from the northern section of Phoenix. This required that ADOT and Project Design Consultant, Howard, Needles, Tammen & Bergendoff (HNTB) design a drain system to collect run-off and transport it approximately 4 kilometers to the Salt River through the City of Phoenix. Tunnels were chosen for the main discharge conduits because depth and location of the system made cut and cover construction impractical. A contract for construction of the tunnels was awarded to Shank-Artukovich-Ohbayashi (S-A-O) on May 24, 1984.

Jan 1, 1987

By M. Di Tommaso

Acid Mine Drainage (AMD) is a quite diffuse environmental problem in dismissed sulphur mine districts. AMD is related to sulphur oxidation of abandoned tailings, generating heavy metal-bearing streams with high iron and sulphate contents and low pH. Treatments of AMD aim firstly to increase the pH to reduce its leaching power causing the further release and mobilization of heavy metals through soil solution and groundwaters. Increasing the pH also allows the precipitation of iron (III) (involved in sulphur oxidation and then in AMD generation) and the partial removal of other toxic components such as arsenic. pH adjustment and the reduction of the high concentration of iron are also necessary for the successive biological treatment in permeable reactive barriers (PRB) with sulphate reducing bacteria (SRB) (II step, biologically mediated precipitaton/biosorption) to avoid the rapid plugging of the barrier (for the large amounts of iron precipitates) and to ensure neutral pH conditions for SRB growth. In this work the first step (chemical precipitation) of a combined chemical and biological treatment of AMD is presented. Experimental data from column tests using carbonate stones to remove iron and increase pH were reported. Toxicity tests have been carried out on the treated wastewaters by Lepidium sativum.

Jan 1, 2006

By Charles C. Y. Chen, Duane R. Skidmore

Batch leaching of Ohio #1, Ohio #4A and Ohio #5 coals was conducted with the thermophilic. sulfur, oxidizing microorganism: Sulfolobus acidocaldarius. Four different size fractions of each coal, with median sizes of 508, 200, 90, and 49 microns, were subjected for 10 days to leaching in a 5% coal-water slurry at 72°C and a pH of 2. Release rates and extent of sulfate and iron I1 and I11 ions from the coals during the leaching process and distribution of sulfur forms (sulfate, pyritic and organic) in the coal samples before and after treatment were determined. Data indicated that Ohio #1 coal, with an initial sulfur content of 0.6%, showed no significant changes with leaching in total sulfur content or forms of sulfur. The Ohio #4A coal, with initial total sulfur content near 5%, yielded product, for sizes below 200 microns with 90% reduction of pyritic sulfur while the extent of organic sulfur removal varied with coal particle size. Results similar to those for Ohio #4A were observed on the Ohio #5 coal, which had an initial total sulfur content of 3% and a final sulfur content near 1%. The findings were interpreted in terms of basic biological, chemical and physical mechanisms.

Jan 1, 1989

By Michael G. Nelson, Dot J. Helm, Edmond C. Jr. Packee

Under a Section 319 Water Quality grant from the U. S. Environmental Protection Agency, a working group of government agencies, mining industry representatives, and University of Alaska researchers undertook reclamation of an abandoned mine site on Birch Creek, near mile 101 of the Steese Highway. Stream hydrology, water quality, and the extent and nature of natural revegetation, were measured before the site was then recontoured to decrease the bedload deposition and bank erosion of the streams. Methods for stabilization of streams and rivers appear to have much in common with those for locations further south, but are subject to complications unique to the North, such as the formation of aufeiss, and to a lack of accurate climatological data. Natural revegetation appears to be possible for most northern placer mine sites. However, it may take place more slowly than required by public perception. Preliminary data indicate a surprising lack of nutrients in the floodplains of mined areas. Natural revegetation may be enhanced by selective planting of grasses and willows, which will contribute variously to soil quality, bank stabilization, and short-term appearance. Monitoring of hydrologic and water quality data, and of natural and artificially placed vegetation will continue at the .Birch Creek site through the summer of 1995,

Jan 1, 1995

By D. Gärtner, A. Oster

"""Operational excellence"" provides a sound foundation for RWE Power AG's three strategic pillars (innovation, expansion and organic growth). The maintenance and development of operational excellence in lignite mining operations at RWE Power is supported by the Continuous Improvement Process (CIP) ""immer:besser"" (getting better). Obtaining cost reductions with the aid of the CIP as a kind of value-adding self-renewal process to ensure competitive lignite mining operations and lignite-based electricity generation forms the basis for further increases in profitability and growth. Since the beginning of the 1990s, RWE Power has carried out a multitude of re-engineering projects focusing on the efficient shaping of processes with the assistance of external consultants to enable the Company to cope with the considerable need for change by swiftly implementing major efficiency levers. These also include the required adaptation and development of the organization. Since 2004, the Continuous Opencast Mine Improvement Process (""KVT"") has enabled the mines to tap significant cost reduction potential by virtue of their own strength. But there is another issue that has come into focus: the development of performance enhancement options. In 2006, the ""KVT"" process, piloted in the opencast mines, was included in the RWE Power-wide immer:besser (i:b) programme (Figure 1)."

Jan 1, 2011

Semiautogenous (SAG) ball mills are being increasingly used today for the coarse grinding of a range of ores, including gold, copper and iron. Larger mines are demanding mills with larger throughputs. This has led to some remarkable developments in mill diameters and installed powers. Conventional design principles - mills with trunnion bearings - have reached their limitations. This article describes one SAG mill with a slide-shoe arrangement and Krupp Polysius' Combiflex twin drive. In August 1997, the world's largest SAG mill with a slide-shoe arrangement was commissioned in Queensland, Australia, operating as a wet mill. The inside diameter is 10.4 x 5.1 m (34 x 17 ft) effective grinding length. The 1.36-ktlh (1,500-stph) SAG mill is driven by two Combiflex drives (twin drives with 2 x 5,500-MW or 7,375-hp installed motor power) in conjunction with a girth gear (outside diameter 11,890 mm or 468 in.). Commissioning of the SAG mill was completed with no notable technical difficulties. The plant runs continuously except for planned maintenance stoppages and process-specific reasons. As this article is written, the machine has been in operation for about 8,000 hours without major problems.

Jan 1, 1998

By G. B. Misra

Heat transfer in partially wet mine airways has been the subject of study for quite some time. Of the available models some are quite empirical while some involve highly complicated computation. Models developed by Amano, Mizuta and Hiramatsu and Banerjee and Chernyak as well as that modified by Misra were evaluated for prediction of the rise in temperature and humidity (mixing ratio ) of partially wet mine airways at four different levels in the Mosaboni Mine of Hindustan Copper Ltd. Measurements of barometric pressure, virgin-rock temperature and changes in dry- and wet-bulb temperatures along a certain length of airway were made. The rise in dry-bulb temperature and mixing ratio of air over the sections were calculated using the three models. The study reveals that calculated values of temperature and moisture content agree within possible rage of experimental error with all the three models in case of three levels with low wetness factor, but the agreement was poor for the relatively more wet airway on the 30th level. Errors in the estimation of fw to the tune of 0.025 can lead to fair error up to 0.94 K in the estimated temperature while alternative ways of estimating the heat transfer coefficient [] can introduce errors up to 1.45 K.

Jan 1, 1987

By S. P. Banerjee

Injection of inert gases to combat mine fires has been practiced in various parts of the world for a long time. Of late, nitrogen has become the preferred inert gas for control of mine fires and safe salvage of longwall equipment as compared to carbon dioxide or fuel gases produced from burning of fuel. In India, carbon dioxide has been used for fire fighting in a number of mines during the 1960s and 1970s. More recently inert gas produced from burning of fuel oil in surface plants has been used in controlling the atmosphere in fire areas in two mines. In yet other cases, liquid nitrogen has either been poured directly into fire areas in liquid form or injected in gaseous form to keep the fire in check. Nitrogen produced in liquid form in an oxygen plant is more costly than nitrogen produced from PSA (pressure swing adsorption) based plants. A PSA-based plant has been installed in one of the Indian mines in 1986 and more are in the offing. This paper reviews the application of various inert gases in Indian mines and discusses the likely trend of inert gas application in Indian mines in the future.

Jan 1, 1987

By P. R. A. Andrews

Areas of bubble contact were delineated for topaz and tourmaline in the presence of chloride, fluoride, sulfate, phosphate, oxalate, and citrate ions using dodecylamine hydrochloride collector. The limits of pH and the minimum collector concentration were established using the bubble pick-up technique. Bubble contact did not occur below the p.z.c. (point of zero charge), estimated to be 5.6 for topaz and 6.6 for tourmaline, in the presence of chloride ion. Flotation below the p.z.c. was, however, extended by citrate, oxalate, phosphate, sulfate, and fluoride ions. Oxalate, citrate, and fluoride ions displayed the greatest pH range of activation, followed by phosphate and sulfate ions. Chloride ion acted as an indifferent electrolyte. Topaz was selectively activated with respect to tourmaline by fluoride and phosphate ions. Topaz was activated by fluoride ion between pH 2.4 and 5.6. Tourmaline flotation in this pH range was virtually insignificant. Topaz was activated by phosphate ion between pH 4.9 and 7.4. Tourmaline flotation in this pH range was only partial. Activation appeared to be strongest when anionic protonated species were in maximum concentration. Strong activation also corresponded to very stable metal ion species thought to be formed on the surface of each mineral.

Jan 1, 1991

By C. McLean, S. G. Kolb, A. J. Erickson, S. C. Moore

The Grossschloppen vein-uranium deposit, Bavaria, West Germany, was examined with underground workings during 1980-82 by Esso Erz GmbH, an affiliate of Exxon Minerals Company (EMC). Geologic evaluation entailed dense drilling of a portion of the deposit from workings constructed specifically for the program. Discovered in 1977, the deposit was initially explored by surface diamond drillholes which allowed definition of a 30-60 m wide vein system discontinuously mineralized along a 1000 m strike length and to at least a 450 m depth. The underground program was conceived as a cost-effective procedure to answer questions on vein correlation, grade continuity and variability. A 1200 m decline allowed access for detailed sampling of approximately 10% of the known area of mineralization. Fanned drillholes, logged by gamma probe, were spaced to provide intersections of veins at 10 to 20 m intervals. Six cross-cuts also penetrate the pitchblende and uranophane mineralization which occurs in 0.1 to 2.5 m thick quartz veins. Detailed cross-sections and level plans were constructed for resource estimates of the intensively studied portion of the vein system. The program resulted in the discovery of local, high grade areas and an average grade in the evaluated area nearly double that expected from surface drilling.

Jan 1, 1992

By V. Arad, S. Arad

Besides the conservation and closing activities, an important issue to achieve is the environmental rehabilitation of the mining perimeters, also assuring the continuity and sustainable development of the mining activity in the Jiu Valley. Modifying and depleting the soils by mining works and dumps of mineral waste as a result of coal mining and tailing is one of the important aspects. The paper’s aim is to analyze which are minerals and resources policies of our government and to adopt measures for sustainable development in the coal mining in this region. The measures to reduce the instability risk for the mining works and waste dumps, which could affect the underground mining constructions or the related surface areas, were consider. It is most important to establish the priorities in order to obtain a high efficiency by combining the measures according to the environmental analysis and risk assessment.

Jan 1, 2005

By H. J. Handewith, W. D. Coski

This paper describes the development and testing of the RAMEX excavation system. Incorporating a synergy of low velocity, high energy impact and kerf cutting, this machine offers a competitive alternative to drilling and blasting in hard rock. It is a boom mining machine that incorporates the features of a lightweight roadheader and it can mine nearly any shape opening in hard rock. The RAMEX Tunneler has been developed to satisfy industry demands for a non-explosive, flexible, continuous and rapid underground rock excavation system. It will produce minimum ground vibration in the urban environment and allow the flexibility necessary for thin seam mining conditions. It will also reduce initial ground support costs and has a minimum impact on rock integrity. Developed by RAMEX Systems with cooperative research provided by the U.S. Bureau of Mines, this system, could produce the next step in the continuing evolution of mechanical mining systems.

Jan 1, 1989

By Half Zantop

While the Andean belt of Chile contains porphyry copper deposits of considerable economic importance, only one major and several minor deposits of this type have been discovered in neighboring Argentina. There are no obvious geologic reasons for the country's apparent paucity of important porphyry copper deposits. Instead, a lack of exploration because of poor accessibility and past inadequate mining legislation may partly account for this situation. Western Argentina has porphyry copper prospects ranging in age from late Paleozoic to Pliocene. Many of these prospects have host rock and alteration characteristics that are very similar to those of large Chilean deposits. The economic potential of many of these prospects has not been tested adequately by geoehemical, geophysical, or chilling programs. A recently enacted mining promotion law and updated mining code removes some past exploration obstacles in Argentina, and makes in-depth exploration and evaluation of' the county's porphyry copper potential attractive.

Jan 2, 1981

By Y. Topkaya, R. E. Johnson, B. R. Palmer, Bruce E. Kurtz, Anatoli Omelian, John I. Choi

This paper relates to hydrometallurgical studies on the recovery of soluble aluminum values from aluminum-bearing ores, i.e., Georgia clays. At least 50% of the soluble alumina values were recovered by leaching the calcined clays from aqueous solutions, comprising simultaneously contacting a stream containing an effective amount of a reducing agent such as sulfur dioxide and a stream containing an effective amount of an oxidizing agent such as chlorine gas for a time sufficient to dissolve alumina. The recovery of alumina from this aqueous slurry of the calcined clay by leaching with sulfur dioxide and chlorine gas showed much better results than that by leaching with concentrated sulfuric acid and/or concentrated hydrochloric acid at 100O C. The leaching reactions between the aqueous slurry of the calcined clay, chlorine gas, and sulfur dioxide were strong exothermic reactions for several hours, requiring no external heating.

Jan 1, 1993

By Brian K. Failon

Scale formation is a problem that affects many mineral processing systems because the processes require the use of large volumes of water. Scale composition may vary from circuit to circuit, but the scale deposition mechanisms and needs for prevention are often somewhat similar. Likewise, there is a degree of commonality in the prediction, detection, and elimination of scaling conditions in various water systems. This paper briefly reviews some of the conditions which lead to scale formation, as well as some of the common problems associated with unchecked scale growth. The primary focus of the paper is on the variety of chemical scale inhibitors available for use in mineral processing circuits, and the actions of each type. The means of predicting and detecting scaling tendencies are also addressed, and include both direct and indirect methods, i-e., stability index determination and chemical analysis of deposits.

Jan 1, 1986

By F. N. Kissell

The Bureau of Mines used sulfur hexafluoride (SF6) tracer gas to evaluate the effectiveness of gob ventilation and/or sealing practices at three coal mines, each having different problems associated with their mined-out areas. The purpose of these ventilation studies was to better understand whether current techniques employed for ventilation or sealing are successful at minimizing the potential for gob fires and explosions. The work performed at each mine is discussed: One is a longwall operation that uses ventilation to carry off hazardous gases, and two are room-and-pillar operations, each with far different concerns, that seal off the gobs to isolate them from the main mine ventilation. In all cases, employment of the Bureau's SF6 tracer gas technique resulted in answers to questions raised by the mine operators concerning the effectiveness of their gob ventilation or sealing practices in preventing mine fires.

Jan 1, 1987

By L. E. Allen, J. F. Abel, D. R. Morgan

The Waste Isolation Pilot Plant (WIPP) is a geologic repository for disposal of transuranic waste and is operated by the U.S. Department of Energy (DOE). The repository is located at a depth of 655 m in the Permian age salt beds of the Salado Formation, 40 km east of Carlsbad, NM. A key component in the design of the WIPP repository is the installation of concrete structures as panel seals in the intake and exhaust drifts after a panel has been filled with waste containers. The DOE has proposed several panel closure design modifications to the EPA. EPA approval of the changes is required since the details of the designs are specified in EPA’s final rule as a condition for WIPP certification. The Environmental Evaluation Group, in its role of providing technical oversight on the WIPP project, has evaluated the proposed modifications. An analysis of this evaluation is discussed.

Jan 1, 2003

By C. A. Rowland

Introduction Ball mills are lined drums, either cylindrical in shape or modified cylinders that have either one or both ends of the shell, consisting of conical sections, that rotate about the horizontal axis. Fig. I I shows a cylindrical mill, Fig. 12 a conical ball mill, and Fig. 13 a Tricone ball mill (Hardinge tradename). Steel or iron grinding media, generally in the shape of spheres, are used to grind the ore to the specified product size. In order to obtain more contact area for grinding and to simulate the shape of worn balls, balls have been made with two concave surfaces diametrically opposite each other. Some concentra¬tors, such as Erie Mining Co., have used slugs cut from worn and broken rods to supplement the balls in ball mills and save money otherwise lost as rod scrap. Cylindrical and conical shapes have been tried instead of balls, but balls remain as the most common shape grinding media used in ball mills. Ball mills were a logical development from the earlier pebble mills that used hard natural pebbles such as flint pebbles or sized ore pebbles (obtained from the ore itself) as grinding media. In the early 1900s36 it was found that when cast iron or cast steel balls were used in place of flint or ore pebbles, the mills drew more power and gave greater production capacity. Advances in technology have resulted in the manufacture of ball mills up to 18 ft diam inside shell, drawing up to 8,000 hp. Ball mills are employed to grind ores, especially the more abrasive ores, to finer sizes than can be produced economically in other size¬reduction machines such as roll crushers, hammer mills, and impactors. Ores can be ground dry-dry grinding-or in a slurry-wet grinding-using ball mills. Dry grinding nominally refers to less than I %v moisture by weight. If the moisture content increases by several percent, dry grinding capacity is significantly reduced as shown in Table 17. The usual range of solids content in wet ball-mill slurries is from 65 to 80% by weight. Wet grinding is used to prepare the feed material for unit opera¬tions such as flotation, magnetic separation, gravity concentration, and leaching that require a slurry of liberated valuable mineral and unwanted gangue particles. Dry grinding" is employed to produce feed for agglomeration, pelletizing, and pyrometallurgy processes that require feed that is dry or nearly so and for finely ground industrial mineral products used in the dry state. Dry grinding is also used when minerals cannot be dewatered economically to the required moisture level or when the ground product reacts unfavorably with liquids. For example, cement clinker must be ground dry. Dry grinding requires about 30% more power than wet grinding for comparable size reduction .28 The total power required in a dry¬grinding ball-mill plant including drying may be double that required for a wet-grinding plant. Grinding-media and liner consumption in dry grinding reported as pounds of metal consumed per kilowatt-hour per ton of ore" is 10-20% of that used in wet grinding. The Wabush pellet plant, Point Noire, Que.3o reported ball consumption dropped from 6.3 lb per ton of ore ground to 2.5 lb per ton of ore ground when they converted from wet to dry grinding, and a 30% increase in power consumption. A number of comparisons made on wet and dry grinding of cement raw materials show metal consumption in dry grinding to be 10% of that in wet grinding. The capital costs for wet grinding are generally lower than for dry grinding. When thickening and filtering of the wet-ground product are required, dry grinding may have a lower capital cost. With open-circuit grinding the ball-mill discharge passes directly to the next processing step without being screened or classified and no fraction is returned to the ball mill (Fig. 14). In closed-circuit grinding the ground material, undersize, in the ball-mill discharge is removed either using a screen or a classifier with the oversize being returned to the mill for additional size reduction (Fig. 15). The over¬size material that is returned to the ball mill is called the circulating load. Open-circuit ball-mill grinding requires more power than closed¬-circuit grinding for products containing similar amounts of top-size material. The less the amount of oversize allowed in the product, the longer the ore must remain in the ball mill when grinding in open circuit. This increases the production of extreme fines and thus the consumption of more power. The power required for open-circuit ball-mill grinding can be estimated using the multipliers listed in Table 18 and knowing the power required for closed-circuit grinding to yield the desired product particle size. For example, assuming the desired grind size is 90% passing some specific top size, open-¬circuit grinding would require 1.40 times the power to achieve similar results as closed-circuit grinding.

Jan 1, 1985