Search Documents

By S. N. Groudev

A process for the combined chemico-biological removal of iron present as oxide minerals in different mineral raw materials (quartz sands, kaolins, clays, etc.) was developed. In the process, mineral raw materials are leached at about 90 °C with lixiviant containing microbially produced oxalic and hydrochloric acids. The leaching is carried out in mechanically stirred acid-resistant reactors for periods of from 1 to 6 hours, depending on the iron content and the forms of iron in the raw materials being leached. The iron contents of some sands treated by this method were lowered from levels that were in the range of 0.035% to 0.088% Fe2O3 to less than 0.012% Fe2O3, making them suitable for the preparation of high-quality glass. The iron contents of different kaolins were lowered from levels that were in the range of 0.65% to 1.49% Fe2O3 to levels in the range of 0.44% to 0.75% Fe2O3. As a result, the whiteness was increased from values of 55% to 87% to values of 86% to 92%. The iron content of clay was lowered from 6.25% Fe2O3 to 1.85% Fe2O3, and this increased the “fireproofness" of the clay from 1,670°C to 1,750°C. A similar process was used for the leaching of aluminum from aluminosilicates, mainly clays and kaolins. However, in this case, the microbial fermentation fluid containing citric acid was acidified by means of sulfuric or hydrochloric acid or by means of different mixtures of mineral acids. For enhancing aluminum solubilization, the aluminosilicates were heated before leaching at 600° to 650°C for 1 to 2 hours. Over 90% of the aluminum present in different clays and kaolins were leached within 3 to 6 hours. "Silicate " bacteria related to the species Bacillus circulans were used to leach silicon from low-grade bauxite ores containing aluminosilicates as impurities. The bacterial action was connected with the degradation of the mineral structures (by means of microbial metabolites such as organic acids and exopolysaccharides), as well as with the selective separation of the rich-in-aluminum fine fractions, which were retained by the mucilaginous capsules of the bacteria. The solid residues after treatment were characterized by higher values of alumina and silicon module (A1203: SiO2 ratio), and they were suitable for processing by means of the Bayer process for recovering aluminum. The bending strength and other ceramic properties of kaolins were improved by contact with well-developed cultures of "silicate" bacteria. The improvement was caused mainly by bacterial metabolites (exopolysaccharides) that acted as resins during drying.

Jan 1, 2000

By I. E. Odom

An electron optical and microprobe study was made to determine the physical and chemical nature of the silica minerals that occur as natural impurities in bentonite and montmorillonitic fuller&apos;s earth clay products (both granular and powder) and that occur in <10-µm dusts from plants that process these products. The purpose of the study was to determine if discrete "free" silica minerals, such as quartz, could be observed using SEM and EDS methods. No “free" silica minerals could be observed in the SEM images of the granular and powdered products, either among or on the surfaces of the particles. Also, no “free" silica could be identified by either SEM or EDS analyses of the dusts, even though quantitative X-ray diffraction analysis indicated that trace amounts of quartz occurred in all samples. It was determined that these quartz and other silica minerals were encapsulated or coated by montmorillonite. Vigorous washing, however, removed some of the clay coatings, so that small amounts of silica minerals could be seen by SEM and EDS analysis. It was found that all primary crystalline quartz grains in the bentonites have diagenetic overgrowths consisting of a mixture of montmorillonite, opaline silica and possibly microcrystalline quartz. Detrital quartz grains in the fuller&apos;s earth have thin diagenetic overgrowths of opaline silica, also containing montmorillonite. The clay encapsulation of the silica minerals in the products studied, as well as in the potentially respirable (<10-µm) dust produced during manufacturing of bentonite and montmorillonitic fuller&apos;s earth clay products, is considered to be an important factor influencing the potential health effects (e.g.. silicosis) of the contained silica. Even if clay dust containing silica should be inhaled, the montmorillonite encapsulating the quartz and other silica minerals would prevent direct contact of the silica with lung tissue. The clay encapsulation of quartz, as well as the highly hydroscopic nature of montmorillonite, may explain why not a single confirmed case of silicosis has ever been reported among plant workers in more than 65 years of production.

Jan 1, 1998

By J. Bowders

A case study describing the expansion of Marion County landfill is presented. The exapansion area is located on a deep mined area. Finite element analyses were used to predict the surface subsidence and associated tensile strains on the landfill liner of the expansion site. A synthetically reinforced liner system is recommended to minimize the impact of the subsurface movements on the integrity of the liner system.

Jan 1, 1986

By A. Iordanidis

Mining activities have recently been taken place in Vourinos mountain, northern Greece, exploiting the so-called ?olivinite?. A representative sample was collected from this mining area. Mineralogical analysis was conducted using X-Ray powder Diffraction (XRD). A thin section was prepared and analyzed under optical microscope. X-Ray Fluorescence (XRF) was employed for the chemical analyses of major, mi-nor and trace elements. Environmental Scanning Electron Microscopy coupled with energy dispersive x-ray analysis system (ESEM-EDX) was also applied for the characterization of the sample. Forsterite (Mg-rich olivine) and enstatite (Mg-rich orthopyroxene) were the main mineral phases determined by XRD, with a 56% and 44% content respectively. Trace amounts of chromite spinels were also determined. Based on this characterization, the industrial rock proved to be a hartzburgite (a type of peridotite in petrological terms). Optical microscopy and ESEM-EDX provided similar observations and confirmed the above-mentioned results. XRF revealed a composition of 43.31 MgO, 44.25 SiO2, 9.98 Fe2O3, 1.40 CaO, 0.67 Al2O3, 0.22 MnO, 0.01 TiO2 and negligible amount of LOI (<0.1%). Trace elements with high concentrations were Cr (2678 ppm) and Ni (2044 ppm). The comprehensive characterization of this mining product suggested the suitability for various uses, such as refractories, foundry and slag conditioner. Additional experiments are being carried out for environmental applications, e.g. wastewater treatment.

Jan 1, 2006

By Mohammad Hossein Sadaghiani

"Construction of underground structures in urban areas is a very challenging work. There are generally two methods of construction, an open cut-and-cover and underground method. To eliminate the subsurface and surface disturbance and street traffic problems, underground method is preferred. The stress redistribution caused by underground excavation induces movements and deformation in the earth mass and ultimately at the ground surface. This is more pronounced in large space excavations. Generally pre-supporting system is used to control ground deformation and thus enhance its stability. In this paper an innovative pre-supporting system is presented. Concrete Arch Pre-supporting System (CAPS) is introduced in Tehran Metro in 2002. This method has roots in construction method of an Old Iranian small water tunnels, called Quanat. CAPS is an efficient method for stabilizing large span underground spaces constructed in shallow and soft ground. In this technique underground reinforced concrete elements are constructed around the proposed underground space prior to main excavation. This method is used successfully in over 50 large span underground structures in Tehran Metro stations and road tunnel intersections. CAPS has potential to be used to pre-support the large span underground spaces at any weak ground condition in an urban area.1. INTRODUCTION Underground subway stations can be constructed either by open cut-and-cover or underground construction. Due to the presence of subsurface utilities and street traffics in urban areas, underground construction is preferred to eliminate relocation of the subsurface utilities, and other urban disturbance. Underground construction in an urban area is a very challenging effort. The complexity of the work is more pronounced in a weak and shallow ground. The stress redistribution caused by underground excavation induces deformations in the earth mass and ultimately at the ground surface causing damage to adjacent structures. The need to control ground surface settlements in urban area is widely recognized and new construction methods are continuously developed. Settlements induced by underground excavation may cause serious damages to nearby structures and subsurface underground utilities, (Sekimoto et. al. 2001). In order to improve stability of underground excavation and reduce settlements, various ground treatment techniques are presented, Johnson, et.al., (1983), Carrieri, et.al. (2002), and Ocak (2008)."

Jan 1, 2016

By Yi Luo

Bolting the mine roof is one of the noisiest underground mining operations. The main contribution to the high noise is the drilling of the roof bolt holes, especially in hard rock. A previous research shows that the specific energy consumed in drilling the bolt holes is inversely proportional to the bite depth (i.e., penetration per drill revolution). Less specific energy at a reasonably high bite depth means less energy is wasted in the drilling process to produce heat, bit wear and noise. Therefore, control of drilling operation to achieve a higher bite depth can provide an opportunity to reduce drilling noise from its sources as a proactive noise control approach. In order to explore the feasibility and to develop the proactive technology for reducing drilling noise, drilling tests have been conducted in laboratory settings. The preliminary results show significant reduction in noise dose can be achieved by drilling in medium hard rock at a reasonably high bite depth. This paper presents the findings from the laboratory experimental studies conducted so far.

Jan 1, 2013

By Robin G. Adams

The market prices received by producers of internationally traded commodity metals have been, and continue to be remarkably volatile and unpredictable compared with most other traded goods and services. This volatility results in an erratic financial performance on the part of producing companies, disruptive and often inflationary cost changes for consumers, unreliable tax revenues and export earnings for host governments and severe social and economic pressures on local, sometimes remote, communities which are dependent on the activity in question. Generally speaking, price volatility appears to be associated with the inability of the metal producing side of the industry to regulate the volume of supply in a manner that is consistent with metal demand, which in turn reflects broad macroeconomic and industrial developments around the world. This statement does not purport to apply to all metals or to all instances of volatility of course. Price movements may occur as a result of pure speculation or even illegal manipulation of the commodity markets. Embargoes, boycotts and other politically induced supply changes may contribute. Threatened or actual strikes also cause artificial swings in sales volumes and prices. How- ever, most of these developments are of a short-term nature. Yet metals price cycles may last for several years, as demonstrated by the extended copper price trough from 1981 to 1986. This paper deals with price volatility which reflects long-term economic cycles and the adaptive processes that exist within the industry in response to such cycles. Special emphasis is placed on the role of forecasts and forecasters in these processes. It should be said at the outset that this reflects the personal experience of one forecaster working in private sector commercial consulting practices that are independent of producers, consumers and host governments. However, the experience does cover the entire period since 1973 and has involved working with a substantial majority of the major companies that supply high volume commodity metals. By way of background it should be stated that the post-1973 world economy has provided a challenging environment for forecasters. One cannot be precise about the forecasting record of either industry insiders or outside consultants, financial analysts and academic forecasters because comprehensive records have not been kept, or if they have the results are too embarrassing to be published. Institutions such as the World Bank have published forecasts of the prices of major commodities on a regular basis and these have, presumably, had some influence on financing decisions in the industry. This record along with our own experience and numerous discussions with metal producers, major consumers and host governments permits certain generalizations to be made. The most important of these is that during the early 1980s. from about 1980 through 1985/6, the forecasters system- atically and substantially over- estimated metal prices. While some fore- casters did correctly anticipate a recession in the early 1980s and some consequential decline in prices, the magnitude and duration of the decline was far greater than anyone foresaw. This systematic forecasting "failure" appears to have occurred independently of the forecasting methodology employed. Moreover, its extent cannot be explained by obvious factors such as incorrect macroeconomic assumptions. There appears to have been some kind of structural change in the world mining and metals industry over this period

Jan 1, 1989

By Erich U. Petersen, W. Dan Hausel, William J. Tafuri, Donald M. Hausen, Douglas N. Halbe

The geology of Wyoming has been compared to some of the richest precious metal producing regions of the world including South Africa, Western Australia, and the Superior Province of Canada. These areas are all underlain by ancient cratons that contain some of the oldest rocks on the surface of the earth. But unlike the other cratons, the Wyoming craton (also termed the Wyoming Province) has been greatly modified by Laramide tectonics which thrust slices of the ancient craton through younger Phanerozoic sedimentary rocks. Unfortunately, the nature of the Laramide tectonics left vast regions of the craton at the bottom of the Tertiary basins unavailable to direct examination, but the exposed slices in the mountain cores contain some of the best exposures of Precambrian rock in the world. The cores of these uplifts contain extensive regions of relatively unmineralized gneiss and granite that include scattered fragmented supracrustal belts formed of metamorphosed volcanic, plutonic, and sedimentary rock. Some of these supracrustal belts exhibit similarities to the gold- and nickel- rich terranes of Western Australia. Yet these areas remain incompletely explored. Archean (> 2.5 Ga) gold deposits within the Wyoming craton are primarily confined to greenstone belts and related terranes. These supracrustal belts include relatively narrow vein and shear zone deposits with strike lengths of tens of feet to more than a mile. Some exceptional deposits possess widths of greater than 15 feet to possibly as much as 300 feet, but the majority have widths of± 5 feet. These belts include Cu, Ag, W, Fe, and Cr mineralization, asbestos and pegmatites, and anomalous Ni and Sri in addition to gold. The Proterozoic terrane (<2.5 Ga) exposed in the cores of mountain ranges in southeastern Wyoming includes a thick miogeosynclinal metasedimentary succession which unconformably lies on the Archean craton. These metasedimentary rocks are separated from a predominantly metavolcanic terrane to the south by a major Precambrian suture, or shear zone. The metasedimentary rocks north of the shear include quartzite-hosted stratabound Cu-Ag-Au deposits, Au-Ag veins, and Witwatersrand-type metaconglomerates with isolated gold anomalies. The suture includes scattered base and precious metal deposits in shear zone cataclastics. South of the shear zone, the metavolcanic terrane includes scattered volcanogenic Zn-Cu-Ag massive sulfide deposits, at least one Cu-Au porphyry, and two large layered mafic complexes. The northern edge of one of these layered intrusives yielded some platinum and palladium to miners at the beginning of the 20th century. Phanerozoic sediments host many precious metal anomalies. Some Au-REE anomalies occur in Cambrian conglomerates. Ag-Cu-Zn stratabound deposits are found in bleached Jurassic redbeds in the Overthrust Belt of western Wyoming. The broad Wyoming basins include many enigmatic gold occurrences and anomalies that can only partially be explained by detrital transport in fluvial systems. Many of these anomalies were probably geochemically transported and precipitated. Rocks of Late Cretaceous age along the flanks of the uplifts have numerous black sand deposits. These paleobeach placers contain concentrations of detrital heavy minerals with anomalous titanium, rare earths, and in some cases gold. Gigantic base and precious metal deposits are associated with scattered composite intrusives in the Absaroka Mountains in northwestern Wyoming. This deeply incised volcanic plateau includes several Tertiary age copper porphyry deposits with significant Cu, Mo, Ti, Pb, Zn, Au, and Ag resources that rival the porphyry districts of the Basin and Range Province. In the Black Hills of northeastern Wyoming, disseminated Au-Th-REE mineralization is found in association with fenitized alkalic igneous rock. In addition to disseminated mineralization-scattered gold, fluorite, rare earth and copper veins; lead, zinc, and silver replacement lodes; and tin pegmatites are found in this region.

Jan 1, 1990

By Jan Visman, James W. Parkinson

GENERAL INTRODUCTION Sample collection, preparation and analysis are all steps in a single process designed to provide information about a much larger amount of material than the sample itself. For example, information about the ash content of a unit train load of coal may be derived from a single, relatively small sample of perhaps a few hundred pounds (liters) extracted from the entire shipment. These three steps can be thought of as links in a chain, allowing the simple analogy that the strength of the chain, or, in other words, the value of the information is only as strong as the weakest link. Sample collection, discussed in Part 1 of this chapter, is critical because analytical procedures can do no more than reveal the characteristics of the sample presented to the laboratory. If that sample does not properly represent the population to be characterized, neither will the analysis of that sample. Sample preparation is also a critical part of the chain. Any change in the nature of a sample between the time it is collected and the time it is analyzed will appear to reflect the nature of the population. For example, sample preparation procedures that allow air to partially dry the coal sample before reaching the laboratory will cause the sample analysis to indicate a lower moisture content than the sample originally had. Other changes can be more subtle. Second only to moisture loss, oxidation reactions generally cause the greatest problems. Oxidation of a coal sample can alter surface properties, changing flotation response. In extreme cases, combustion parameters, such as caking or agglomerating properties or mineral composition, such as pyrite content, can change. Water samples can also be altered by oxidation. For example, ferrous iron can oxidize to ferric iron, possibly changing water quality parameters such as pH. Sample analysis, the subject of Part 2 of this chapter, is the final link in the chain. Obviously, no laboratory procedure can correct for problems created by improper sample collection and preparation. At the same time, poor analytical techniques will create incorrect information about the total amount of coal from which the sample was extracted (the population) in the same manner as improper sample collection and preparation. The key to gaining correct information about the population is to view these three steps as parts of a single process. Following proper techniques through only one or two steps of the process will not produce accurate information.

Jan 1, 1991

By C. J. Bise

Virtual reality (VR) is a concept in which the human experience of perceiving and interacting with a computer-modeled environment is achieved through the use of sensors and effectors. It is an advanced form of computer graphics. The images are generated based on where the user looks and how the user moves. As personal-computer-based (PC) instruction gains acceptance as an innovative approach to providing safety training, VR can provide a multisensory, rather than merely a multimedia, mode of presentation. New technology continues to be introduced into the mining industry to meet the productivity needs of coal and noncoal producers. There is also an increased need for more innovative training of mine workers. New technology and equipment designs often affect the basic causes of accidents. For example, the development of cabs and canopies for underground coal-mine equipment has reduced the number of accidents related to roof falls and rib rolls. However, safety-training specialists then had to focus on the machines themselves, because the modifications had the potential for causing "pinch-point" accidents. In fact, although the overall physical demands on mine workers have diminished, there has been an escalation in the concerns about trauma from repetitive physical activity. Thus, health-and-safety training specialists have shifted their atten¬tion to the recognition, evaluation and control of workplace hazards (Hancock and Hill, 1995). Recognizing that the spatial relationships of a workplace convey more information about hazards when experienced rather than described, trainers agree that simulation is the method of choice. Unfortunately, until recently, simulation systems were either simple software programs that did not come close to portraying the real experience or were complex hardware systems (such as flight simulators) that did, but at a cost of millions of dollars (Fritz, 1991). However, researchers and training specialists are excited about an emerging approach to work-force education based on virtual reality. VR is also described as cyberspace, immersive simulation, artificial reality, telepresence, virtual environment or virtual world (Meyer and Dunn-Roberts, 1992; Peterson, 1992). It represents a group of concurrent advances in computer and interface technologies. They enable an individual to interact in a multisensory way with a computer-generated environment. To achieve this, a typical VR system incorporates a computer that contains the necessary databases for producing realistic images and audio; a head-mounted display (HMD) or screen for viewing the images; and a three-dimensional controller, joystick or data glove for interacting with the computer-generated environment. If virtual reality is to become an effective tool for the safety training of mine workers, its attributes will have to be measured against other forms of delivery. Accordingly, Carr (1992) listed the following areas in which VR should be considered the method of choice: ? When training mistakes would be costly. ? When the necessary environment cannot be experienced in the real world. ? To build interfaces that are sensible and can be manipulated. ? To make training situations really "real." ?To make perceptible the imperceptible.

Jan 1, 1997

By Robert E. Blackett

To help local governments plan for long-term supplies of sand and gravel for buildings and infrastructure, the Utah Geological Survey is assessing natural aggregate resources (mainly sand and gravel) in the St. George, Utah area. Whereas much of the aggregate in the study area is adequate for road-base construction, high-quality aggregate for Port- land cement concrete is limited due to geology, and also due to land administration and environmental issues. Virgin River alluvial gravels southwest of St. George are generally poor sources of aggregate because they contain clasts of soft sediments and soluble minerals derived from Mesozoic rocks. Elsewhere, widespread secondary calcium carbonate (caliche), as rinds and matrix in gravel, lower the quality of material. Young river-terrace deposits in Fort Pearce Wash, and along the Virgin River above St. George are the primary sources of sand and gravel for the area. Alluvial fans on the east side of the Beaver Dam Mountains, and on the west side of the Hurricane Cliffs are also sources of aggregate. Stream-channel deposits associated with "inverted valleys" north of the towns of St. George and Hurricane, Utah, and older pediment gravels on the west flank of the Beaver Dam Mountains may be future sources of high-quality aggregate. Accelerated urban development during the 1980s and early 1990s has caused St. George and other communities in Washington County to encroach upon sand and gravel pits originally located in rural, undeveloped areas. Urban residents often view nearby sand and gravel operations as undesirable neighbors and pressure elected officials to close the pits. Pit closures effectively eliminate the remaining reserve from the sand and gravel resource base of the area. More- over, resources may also be lost as new homes are built upon undeveloped sand and gravel deposits. The St. George area is particularly susceptible to loss of aggregate resources because high-quality natural aggregate for construction is already in short supply. Moreover, most of the land in the study area is administered by the U.S. Bureau of Land Management and subject to various land-use restrictions. Much of this public land cannot be developed for mineral materials, such as sand and gravel, because of federal regulations protecting rare plants and animals, archaeological sites, and important watershed areas. Of particular concern to resource development are areas designated as Areas of Critical Environmental Concern, Wilderness Areas, and Wilderness Study Areas. Protection for Endangered Species is a critical element in all planning efforts. Those species present in the St. George area or in areas near- by include: the desert tortoise, southwestern willow flycatcher, Virgin River chub, and dwarf bearclaw poppy. As the St. George area is depleted and/or deprived of sand and gravel resources through extraction and urbanization respectively, construction costs will rise unless conservation measures are instituted. To avoid the loss of aggregate resources, the UGS is pursuing a program of data gathering and field surveying of surficial deposits to determine tonnage and quality of the remaining sand and gravel resources. The resource information from this program will help land planners, developers, and government officials make informed decisions concerning urban development throughout the region.

Jan 1, 2001

The response of 174 domestic water supplies to longwall mining of the Pittsburgh coal seam was compared to various physical parameters. Sixty-four percent of domestic water supplies returned to service without the need for intervention, while thirty six percent required intervention to reestablish a suitable water supply. Domestic well response is strongly correlated with the topographic setting. Valley wells showed the least effect while hilltop wells showed the greatest effect. Well response was also found to be correlated to the location of the well above the mining with greater effects observed in zones of surface tension and compression, and fewer effects in zones which are stress neutral. Ground water supplies remained unaffected or were re¬established at 149 of 150 sites, indicating that ground water resources remain available to domestic wells after mining. Results of this study support the stress-relief model of ground water flow in the Northern Appalachian Basin. Longwall mining&apos;s effect on this system is to locally increase fracture connectivity in the near surface zone. This may shift the zones of saturated fractures to greater depths beneath hilltops and hillsides, and increase recharge and storage in valleys. General trends in well response are predict¬able utilizing the stress relief conceptual model, the topographic setting of the water sources, and their location over the mine.

Jan 1, 1986

By James R. Arnold, Cindy S. Jones, Michael F. Gleason, John O. Marsden, John G. Mansanti

INTRODUCTION The Chimney Creek Gold Mine (Gold Fields Operating Co. - Chimney Creek) is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. The operation is a wholly owned subsidiary of Gold Fields Mining Corporation, the North American branch of Consolidated Gold Fields PLC, London, England. The plant started up in November, 1987, less than three years after discovery of the orebody and three months ahead of schedule. Ore is mined in an open pit and is processed by combined dump leaching and milling techniques for gold and silver recovery. The mine is set to produce approximately 150.000 ounces of gold and 50,000 ounces of silver per year over a 12 year life at current reserve estimations. The mine was designed and constructed at a cost of $79.3 million with engineering and construction services provided by Davy McKee Corporation, San Ramon, California. Key Gold Fields operating staff were involved in the design of the facility from the start of the project: The Mine Manager, Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist were all involved full time on the project within 5 months of the first ore discovery. Emphasis was directed at optimizing operating efficiency and in particular minimizing labor costs in the plant. It was recognized that a high level of instrumentation and control would be required to achieve this. The risk associated with the instrumentation and control systems implemented was to be minimized by using equipment and systems that had been proven in industry while utilizing the most cost effective, state-of-the-art technology available. The reliability of the overall control system was considered to be critical in view of the cost of downtime associated with the gold extraction plant. BRIEF PROCESS DESCRIPTION The dump leaching process treats approximately 1.2 million tons per year of low grade ore at an average grade of 0.035 oz/ton. Run of mine material is dumped on a lined leach pad and weak cyanide solution is applied by drip irrigation. Pregnant solution run off is pumped to carbon columns in the milling plant for gold recovery and the barren solution returned to the dump leach circuit. Average gold recovery is 60%. This process has little instrumentation and control associated with it. The milling operation treats 700,000 tons annually of higher grade ore (0.200 oz/ton initially, dropping to an average of 0.135 oz/ton after first two years). Recovery is directly related to head grade (fixed tail assay effect) and currently averages 96%. A single pass through a jaw crusher reduces run of mine ore to minus 12 inches. The ore is stockpiled and reclaimed by loader for grinding in a two-stage milling circuit consisting of a SAG mill and ball mill, the latter in closed circuit with hydrocyclones. Cyanide and lime are added into the SAG mill to start dissolution of gold as early as possible in the circuit. The ground product leaves the milling circuit at approximately 78% minus 200 mesh and is fed to an unique "double thickener" leaching-recovery circuit. This circuit has been discussed in detail in a paper by J. G. Mansanti et a1 (1). Two thickeners are arranged in counter- current configuration with three leach tanks. Overflow solution from the first thickener is treated by carbon-in-columns (CIC) for gold recovery with 85% of the soluble gold recovered onto this carbon. Underflow slurry from this thickener is pumped to the leach tanks, with a total retention time of 12 hours, and then gravitates to the No. 2 thickener. Overflow solution from the second thickener is used as a wash in the first thickener. Underflow slurry from the second thickener is treated in a carbon-in- pulp (CIP) scavenging circuit to recover the remaining 15% dissolved gold. Gold-loaded carbon from both the dump leach and milling circuits is stripped in batches using the Zadra hot caustic- cyanide elution process. Gold (and silver) is recovered from the hot strip solution by precipitation with zinc dust and the product recovered on Funda pressure filters. The precipitate is retorted to remove any mercury and then smelted into buttons. The buttons (approximately 80% gold, 15% silver) are shipped to an independent refiner in Salt Lake City, Utah, for further treatment.

Jan 1, 1990

By S. L. Bessinger

Introduction In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway&apos;s complex geometric cross section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Head losses in mine entries are calculated using Atkinson&apos;s Equation. [22H= KPLQ (English) H= KP 3Q (SO (1) 5.2AA] where H = pressure loss, in. of H2O (Pa); K = friction factor, lbf•min2/ft4 (kg/m3); P = perimeter, ft (m); L = airway length, ft (m); Q = airflow quantity, ft3/min (m3/sec); and A = flow cross-sectional area, ft2 (m2) In this equation, the friction factor, K, is an empirical constant that describes the aerodynamic roughness of the airway. Typically, the K-factor for a given airway is determined by measuring the factors H, P, L, Q and A in Equation (1) and calculating K. Tables of friction factors calculated in this way are found in textbooks and handbooks that deal with mine ventilation analysis. Unfortunately, very few K-factors have been measured on longwall faces, and the accuracies of those that have been measured are entirely site specific, because of the wide variety of equipment found on longwalls. The development of a technique for prediction without mine-site measurements of the friction factor for any longwall face, regardless of its configuration, will thus be very useful in the design of ventilation systems for mines in which longwall mining is practiced. Calculation of pressure losses using Atkinson&apos;s Equation (1) and empirically determined K-factors provides accurate and useful approximations in cases where the airways have relatively simple cross sections. However, a careful analysis using the principles of fluid mechanics shows that such calculations are based on two assumptions that are not strictly correct when there are obstructions in the airway. The first assumption is that the air velocity distribution in the cross section, particularly around the perimeter, is uniform. This assumption results from the fact that the tabulated K-factor values found in the literature are based on field measurements with uniform conditions. Such uniformity does not exist in longwall airflows. The second common assumption is that the K-factor, and corresponding head loss, is independent of the Reynolds Number (NR) for a given flow. In fact, this assumption is not strictly correct, and is particularly erroneous in the case of irregular protuberances into the airflow, such as those found on a longwall face. The errors arising from the assumptions may be avoided by using K-factors calculated by a newly devised method, described below, which takes into account the fundamental principles of aerodynamic drag analysis. This new technique has two advantages: first, it is flexible enough to model any longwall, regardless of equipment configuration; second, it employs terminology and equations familiar to those who perform mine ventilation analysis, using K-factors, for which ventilation engineers have an intuitive understanding, rather than drag coefficients. To provide guidance for development of a longwall drag model, data were taken on two modern longwalls operating in substantially different conditions. Pressure measurements at Mine B were made with 200-foot (61-m) sections of 1/8-in. (3-mm) diameter plastic tubing, attached to a Dwyer Magnehelic gauge. Pressure drops were measured in 200-foot increments down the face, and summed to give the drop for the entire face length. This technique was found to produce small, repeated errors because of the number of segments required to span the longwall. At Mine A this problem was avoided by using a single, continuous, plastic tube for the entire face length. The psychrometric properties of the air were measured for both Mines A and B. A calibrated, standard vane-anemometer was used to measure the airflow on both faces. Finally, numerous dimensions were measured on both faces, and face profile drawings were obtained to allow detailed evaluation of the face equipment geometry. From this information, accurate evaluations of the average wetted perimeter and average area of the longwall face airways were made. Since the airflow is not confined to inside the powered supports at all points along the face, a quadratically weighted average of the airflows measured at various stations along the face was calculated: [n2Qavg =Qi Ii / It(2)i=1] where [Q, avg = average airflow for analytical purposes, ft3/min (m3/sec); Q= airflow at station i, ft3/min (m3/sec); 1= length of segment represented by Q, ft (m); h= length of longwall face, ft (m): and n= number of quantity measurement stations.] The quadratic weighting scheme was chosen because the

Jan 1, 1992

By Ivor Hawkes

INTRODUCTION High energy breaking is an alternative to using ex¬plosives in underground secondary breaking operations. It also is a means of upgrading conventional hand-held breakers, manual sledge-hammer breaking, and scaling bar operations. Major areas of application are in sec¬ondary breaking over grizzlies and at drawpoints. Other applications include breaking down ripping lips in longwall seam mining, scaling in stopes and rooms, general demolition work, and roadway maintenance. There is considerable interest in high-energy impact breakers for use in primary ore breaking, but, as of 1977, all such applications have been only experimental (duToit, 1973; Joughin, 1976; Wayment and Grantmyre, 1976). EQUIPMENT Essentially, a high-energy impact hammer is a boom¬mounted pneumatically or hydraulically actuated breaker. The machine basically consists of a piston that oscillates in a housing and impacts the end of a tool or moil thrust against the rock. The force applied to the rock primarily depends upon the impact energy of the piston-the higher the impact or blow energy, the greater the force and, thus, the greater the rock break¬age. Among drill and breaker designers, a common expression for blow energy is "force of blow." Hand-held breakers are limited to blow energies of about 140 J (100 ft-lb), because the operator is unable to handle heavier machines efficiently or to absorb the recoil energy resulting from higher blow energies. How¬ever, these restrictions do not apply to boom mounted breakers; machines with blow energies on the order of 4000 J (3000 ft-lb) and higher are available commer¬cially for underground use. There is considerable evi¬dence to show that increasing the blow energy also in¬creases the efficiency of the breaking operation; i.e., more rock is broken per unit of energy expended (Grantmyre and Hawkes, 1975). Thus, there is a trend to higher blow-energy machines, particularly where high¬strength rocks are to be broken. In relation to rock breaking, the blow rate of boom¬mounted impact breakers is not as important as it is for rock drills. This is because the breaker must be moved over the work surface between blows. The blow rate is governed eventually by the power supply, and typical blow rates range between 200 and 600 blows per minute. As a general rule, light blow-energy machines have higher blow rates than heavier machines. Table 1 lists most of the boom-mounted impact breakers that were available commercially during 1977, and it gives details of the blow energies and machine weights. Restrictions are placed on the blow energy by the machine weight and size, and by the strength of the boom. Typically, boom-mounted impact hammers have a blow-energy to mass ratio of about 1.5, with lower values for lighter machines and higher values for heavier machines. In addition to supporting the hammer weight, the boom also has to absorb the recoil energy of the blow, which can be on the order of 1400 J (1000 ft-lb) for large hammers operating in a horizontal mode. Interesting exceptions to the general run of impactors are the Joy HEFTI hydraulic hammers. In these machines, the piston impacts onto a fluid cushion that is positioned between the piston and the impact tool. This approach allows very high piston velocities, over 30 m/s (100 fps), to be used without the risk of break¬ing the piston or impact tool. Steel on steel impacts must be limited to impact velocities of about 10 m/s (35 fps) due to the high impact stresses generated; thus, increased blow energies can be achieved only by increas¬ing the piston size. The Joy 514 HEFTI®, listed in Table 1, has a blow energy of 27 100 J (20,000 ft-lb), but, as of 1977, the machine has been used underground only on an experimental basis. Using a fluid cushion between the piston and the impact tool allows the use of light pistons, reducing the overall machine weight. The recoil energy, which must be absorbed by the boom for a given blow energy, is directly proportional to the piston to machine mass ratio, and operating with light pistons provides an addi¬tional benefit in reducing the requisite boom size. Both pneumatic and hydraulic hammers are avail¬able commercially. Although hydraulic hammers are a relatively recent development, they already outnumber the pneumatic machines in use. There are many reasons

Jan 1, 1982

By Chester A. Rowland, David M. Kjos

INTRODUCTION Mineral ore comminution is generally a feed preparation step for subsequent processing stages. Grinding, the fine product phase of comminution, requires a large capital investment and frequently is the area of maximum usage of power and wear resistant materials. Grinding is most frequently done in rotating drums utilizing loose grinding media, lifted by the rotation of the drum, to break the ores in various combinations of impact, attrition and abrasion to produce the specified product. Grinding media can be the ore itself (autogenous grinding - primary and secondary), natural or manufactured nonmetallic media (pebble milling) or manufactured metallic media - steel rods, steel or iron balls, or a combination of autogenous media and steel balls (SAG milling). This chapter covers rod and ball mills utilizing manufactured metallic grinding media. MILL DESIGN The interior surface of rod and ball mills exposed to pulp and/or grinding media are protected from wear and corrosion by rubber, metallic or a combination of rubber and metallic wear resistant materials. Rod and ball mills essentially draw constant power, thus are well suited for use of synchronous motors with power factor correction capabilities as drive motors. A net of approximately 120 to 130 percent of running torque is required to cascade the charge in these mills. The pull-in torque is about 130 to 140 percent with the pullout torque to keep the motor in-step (in-phase) generally in excess of 150 percent. When rod and ball mill are started across-the-line the starting and pull-in torques result in inrush currents exceeding 600 percent which possibly result in high voltage drops. To deliver 130 percent starting torque to the mill the motor design must take into account the maximum anticipated voltage drop. Motor torque decreases as the decimal fraction of the voltage available squared. E.g., a motor rated 160% starting torque with a 10% system voltage drop will deliver 160% x or 129.6% torque to its output shaft When it is not possible or practical to start a fully loaded synchronous motor across-the-line it is possible to utilize the motor&apos;s pull- out torque to start the mill. By using a clutch, normally an air clutch. between the motor and the mill, the motor is brought up to synchronous speed before the clutch is energized. If the motor has an adequate amount (175 or greater) of pull-out torque the pull-out torque starts the mill without major disruptions on the electrical system. Since the energy release at initial cascade of the mill charge is an inverse function of acceleration time, a minimum acceleration time of 6 to 10 seconds or more is recommended to prevent damage to the mill or the mill foundation. Economics at the time of plant design and mill purchase determine the drive to be used. The simpliest drive is the low speed synchronous motor with speeds in the range of 150 to 250 RPM connected to the mill pinionshaft by either an air clutch or flexible coupling. Using a speed reducer between the motor and pinionshaft permits using motors having speeds in the range of 600 to 1000 RPM. In this speed range, if power factor correction is not required induction motors can be used; squirrel cage where there is no restriction on inrush current; slip ring where a slow start and low inrush current is required. Air clutches can also be used to ease starting problems with squirrel cage motors. In some areas of the world induction motors and starters are less expensive than synchronous motors at a sacrifice of motor efficiency and power factor correction. Dual drives, that is two pinions driving one gear mounted on the mill, become economical for ball mills drawing more than 3500 to 4000 horsepower (2600 to 3000 kilowatts). Further developments of the low frequency, low speed synchronous motors with the rotor mounted on the mill shell or an extension of one of the mill trunnions could improve the cost picture for these "gearless drives", making them practical for large ball mills. The percent of critical speed, which is the speed at which the centrifugal force is sufficiently large to cause a small particle to ad- here to the shell liners for the full revolution of the mill is given in mill specifications. Critical speed is determined from the following: Where D is mill diameter inside liners (specified in meters). Cs is critical speed in RPM. When D is specified in feet

Jan 1, 1998

For all the attention given to mining activities in Central and South America, Mexico, Indonesia and Papua New Guinea, it is sometimes easy to forget the activities taking place in one of North America&apos;s most important and largely untapped regions -Alaska. The Alaskan minerals industry employed about 3,406 people during 1995. That is a 23% increase from - - 1994. It also contributed about $719.8 million to the state&apos;s economy during 1995, a 10% increase. During the past few years, Alaska has seen significant increases in exploration and mining activity. Fairbanks Mining, an arm of Cyprus Amax, began producing gold during 1996 at its Fort Knox Mine, as did Nevada Goldfields&apos; Nixon Fork gold mine. Kennecott&apos;s Greens Creek lead-silver-gold mine also reopened and USMX began operations at its Illinois Creek gold mine. In addition, several impressive exploration projects were under way in Alaska. Placer Dome reported good results at Donlin Creek project in southwest Alaska. Likewise, Newmont Exploration is getting serious about its True North gold project. And Nerox is nearing production at its Jonesville coal project in Alaska&apos;s interior. In November, about 400 miners gathered in Anchor- age for the Alaska Miners Convention. Throughout the conference and exhibition, attendees listened to updates on these precious- and base-metal and coal exploration and mining operations. They also learned that the state wants more miners to head north and help develop its natural resources. Alaska is ripe for a smelter The mining industry -Alaska&apos;s miners in particular - has a powerful ally in Sen. Frank Murkowski (R-AK). He is the chairman of the Senate Energy and Natural Resources Committee. During his luncheon address, Murkowski said it appears that it is economically feasible to extend the Alaska Railroad to northwest Alaska, taking advantage of the large coal and hard-rock mineral tracts in the Ambler mining district. He also renewed his push for building a major smelter in Alaska to process ore that would come from the Ambler mining district. Murkowski believes that Alaska&apos;s economy can be expanded by extending the railroad from Nenana to the Point Lay coal fields to open coal development. Along the way, the railroad would pass through the Ambler district, helping to develop the hard-rock deposits. An outgrowth of this additional mineral development, Murkowski said, could be a new smelter in Alaska. It could use power from the Tyee hydroelectric facility near Wrangell or from the Makushin geothermal project near Unalaska. Either site, he added, offers access to plentiful, relatively low-cost power, at tidewater to facilitate ore deliveries to the Pacific Rim. The Congressional Research Service has determined that there are enough mineral deposits and coal beds to make it possible to finance the extended railroad, he said. The 1,200-km (744- mile) rail extension would cost about $4.2 billion. The North Slope of Alaska contains up to 3.6 Tt (4 trillion st) of coal, Murkowski said, that contains less than 0.25% of sulfur and less than 4% moisture. It aver- ages 12,300 Btu in heat content. "That is a tremendous resource," he said. The Ambler mining district, meanwhile, has at least eight identifiable hard-rock deposits, Murkowski continued. These contain zinc, lead, copper, nickel, chromium, tin, barite, gold and silver. "All eight of these can generate ore concentrates at rates similar to Red Dog&apos;s cur- rent 725 ktla (800,000 stpy)" of zinc, he said.

Jan 1, 1997

By D. Diggins

Experimental apparatus for the determination of powder wettability is described. The principle of the technique is based on the measurement of the equilibrium capillary pressure in a packed bed of particles. Both the macroscopic solid-solution-vapour contact angle, e, and the specific surface area, A(m2g-1), of the powder, can be obtained. Experiments on angular quartz particles have been performed. A wide range of situations have been covered: hydrophobic and hydrophobic particles wetted with water, a variety of organic liquids and solutions of surfactants. The results show that reliable and reproducible values for 0 and A for angular and relatively coarse powders can be obtained with the proposed equipment for advancing liquid fronts in the absence of soluble surfactants. In the latter case and for receding liquid fronts in general, only qualitative trends have been obtained.

Jan 1, 1992

By Ehsan Alavi Gharahbagh

Tool and cutterhead wear is a crucial issue in soft ground tunneling, especially because tool inspection, maintenance, and replacement are done under difficult conditions, and often under pressurized faces or under ?hyperbaric interventions?. Overall, tool inspection and changing is a time consuming and costly step, which needs to be quantified and the related costs accounted for in any construction contract for scheduling and costing of the tunnels at design stage or bid preparation. However, lack of a universally accepted index or measure for soil abrasivity in the geotechnical reports has made it difficult, if not impossible, to define cutter wear for this purpose. Soil abrasion and its impact on cutting tools for tunneling machines under dry or wet conditions have recently been under spot light due to construction claims in several of the recent tunneling projects. The objective of this paper is to examine soil?s physical, mechanical and tribological properties related to the interface of the soil and excavation tools and discuss ways to measure soil abrasivity in a quantitative and repeatable manner. The paper will discuss the efforts that are underway to develop a new soil abrasivity index and the fabrication of a new testing device for this purpose.

Jan 1, 2010

By Fred N. Kissell

INTRODUCTION During the last few years, several new ventilation developments have attracted the interest of mining engi¬neers. Some of these developments are applicable pri¬marily to hard-rock mining, while others are more applicable to coal mining. STOPPINGS Parachute Stopping The parachute stopping is a new type of quick-erect stopping that is intended for temporary use in hard-rock mines (Kissell, Thimons, and Vinson, 1975). As shown in Fig. 1, the stopping is shaped very much like an ordinary parachute, with a canopy of impermeable fabric that is sewn to regularly spaced straps running to a common point. To erect the stopping, the straps are attached to a fixed anchor point such as a roof bolt, and the edge of the canopy is lifted into the moving air¬stream. The airstream pops the parachute canopy into place, and the differential air pressure across the stop¬ping holds it in place, forcing the fabric against the walls, roof, and floor of the mine opening. The principal advantage of the parachute stopping is that it requires only a few minutes to install, making it a great time-saver for emergency use or for day-to¬day changes in ventilation during the production cycle. However, the parachute stopping does require some minimum air velocity to lift it and some minimum differential pressure to hold it in place. For a fabric weighing 0.27 kg/ m2 (8.0 oz per sq yd), the minimum air velocity is about 0.5 m/s (100 fpm), and the mini¬mum differential pressure is about 0.05 kPa [0.2 in. water gage (WG) ]. There is always some air leakage around the stop¬ping, mainly depending upon the extent to which pipes or other obstructions encumber the airway and prevent good sealing. Leakage of a few cubic meters per second (a few thousand cubic feet per minute) can be expected, unless foam is used to improve the seal at the edges of the canopy. Quick-Fix Blowout Stopping The quick-fix blowout stopping is a variation of the parachute stopping (Thimons and Kissell, 1976), and it is used in the proximity of blasting operations. This type of stopping is designed to be blown out easily by the blast forces, and it may be reinstalled quickly and easily. The long high-strength straps of the parachute stopping are replaced by groups of short straps that tear easily. These straps are attached at six equally spaced locations around the perimeter of the canopy. To erect the stopping, one strap of each of the six groups is fastened to the mine wall, roof, and floor by using spads, by setting pins with a powder-actuated gun, or by tying the straps to some firm anchor point. Once the straps have been attached, the differential air pressure across the stopping, which must be at least 0.025 kPa (0.1 in. WG), forces the stopping perimeter against the mine walls, thus creating the air seal. It is the self-sealing feature of this stopping that makes it a significant time-saver. Only a few attachment points are needed; in many cases, four attachment points are sufficient, since the stopping naturally tends to form a seal with the airway surfaces. When nearby produc¬tion blasting exerts excessive forces on the stopping, one or more of the straps tears away from its attachment point, protecting the stronger canopy from damage. Damage-Resistant Brattice The damage-resistant brattice is a stopping that is designed for use in mines such as salt and limestone mines where the differential pressures are low and the roof is relatively flat. As shown in Fig. 2, the damage-resistant brattice consists of a series of brattice panels that are hung vertically and joined by Velcro® connections. When the brattice is subjected to strong blast forces, the Velcro® connection peels apart and allows the panels to open without incurring damage. The Velcro® connections can be resealed by hand within a matter of minutes. Such damage-resistant brattices have withstood the blast effects of 318 kg (700 lb) of ammonium nitrate-fuel oil (ANFO) explosive detonated as close as 91 m (300 ft) from the brattice. Ordinary brattice cloth is used for the panels, with a 51-mm (2-in.) wide strip of Velcro® hooks sewn along one edge of the length, and a 51-mm (2-in.) wide strip of Velcro® pile sewn along the other edge. Both the hooks and the pile are sewn onto the same side of the brattice cloth. The resulting Velcro® seal formed be¬tween adjacent panels is perpendicular to the brattice itself, and the leading edge of the seal can be directed either toward or away from the blast forces; the brattice works equally well in either case. To hang the brattice, panels of brattice cloth about 0.9 m (3 ft) longer than the height of the airway are cut from a 1.8-m (6-ft) wide roll. The additional 0.9 m (3 ft) of brattice cloth allows 0.3 m (1 ft) for attachment to the roof by means of a board, with 0.6 m (2 ft) for forming a good air seal at the floor. Each brattice panel is wrapped once or twice around a 51 X 102 mm (2 X 4 in.) or 25 X 76 mm (1 X 3 in.) mounting board that is 254 to 305 mm (10 to 12 in.) shorter than the width of the panel. For convenience in

Jan 1, 1982