Search Documents

By Jeong-Heum Kim, Jan M. Mutmansky

The primary purpose of the research summarized in this paper was to experimentally investigate the performance of numerous mine ventilation network iterative solution procedures under controlled conditions to determine which procedures were most favorable in terms of number of iterations and total computational speed. The solution methods investigated included: (1) the Newton-Raphson method, (2) the Hardy Cross method, (3) the linear theory method, and (4) the second-order approximation method. The first three of these methods have been widely used while the fourth was formulated by the authors to investigate its potential for ventilation network computation. In addition to the four solution procedures, three methods for selecting the branches in the meshes and three methods for ordering the meshes for computational purposes were also studied during the investigation. As a result of these various options, a total of 36 combinations of solution methods, mesh selection procedures, and mesh ordering schemes were studied. To investigate the performance characteristics of these combinations, an identical set of 180 randomly chosen ventilation networks was used in sets of 30 networks at six different levels of complexity. Here, the level of complexity is defined as a function of the size (number of branches) and the density (number of nodes compared with the number of branches) of the network. To eliminate variations in the solution times due to differences arising from the fans, a single fan with a fixed characteristic was utilized in all the experiments. This paper will deal only with the results for the four iterative methods used to achieve solutions to the mine ventilation networks. The minimum-resistance spanning tree method for mesh selection and random mesh ordering was used throughout the analysis. The conclusions will be based upon a detailed accounting of the efficiency of each method using the total time to achieve a solution to each problem and averaged over the total problem set. However, the number of iterations and the time to complete each iteration will be studied in order to better understand the solution procedures.

Jan 1, 1991

By Barnaby J. Watten, Sibrell. Philip L., Brian J. Vinci, Andrew E. Friedrich

We evaluated a new process for remediation of acid rock drainage (ARD). The process treats ARD with intermittently fluidized beds of granular limestone maintained within a continuous flow reactor pressurized with CO2. Tests were performed over a thirty day period at the Toby Creek mine drainage treatment plant, Elk County, Pennsylvania in cooperation with the Pennsylvania Department of Environmental Protection. Equipment performance was established at operating pressures of 0,34,82, and 117 kPa using an ARD flow of 227 L/min. The ARD had the following characteristics: pH, 3.1; temperature, 10 OC; dissolved oxygen, 6.4 mg/L; acidity, 260 mg/5; total iron, 21 mg/L; aluminum, 22 mg/L; manganese, 7.5 mg/L; and conductivity, 1400 uS/cm. In all cases tested, processed ARD was net alkaline with mean pH and alkalinities of 6.7 and 59 mg/L at a CO2 pressure of 0 kPa, 6.6 and 158 mg/L at 34 kPa, 7.4 and 240 mg/L at 82 kPa, and 7.4 and 290 mg/L at 117 kPa Processed ARD alkalinities were correlated to the settled bed depth (p<0.001) and CO2 pressure (p<0.001). Iron, aluminum, and manganese removal efficiencies of 96%, 99%, and 5%, respectively, were achieved with filtration following treatment. No indications of metal hydroxide precipitation or armoring of the limestone were observed. The surplus alkalinity established at 82 kPa was successful in treating an equivalent of 1136 L/min (fivefold dilution) of the combined three ARD streams entering the Toby Creek Plant. This side-stream capability provides savings in treatment unit scale as well as flexibility in treatment effect. The capability of the system to handle higher influent acidity was tested by elevating the acidity to 5000 mg/L with sulfuric acid. Net alkaline effluent was produced, indicating applicability of the process to highly acidic ARD.

Jan 1, 2000

By Fikret Suner

Two rare evaporite minerals, nahcolite (NaHC03) and trona (Na2C03?NaHC03?2H20), are both widely used in the production of sodium carbonate (Na2CO3) The largest natural carbonate deposits (trona) are in Wyoming, USA, the second largest deposits are in Turkey, in a lacustrine basin (Beypazari) where nahcolite and trona are located within the Miocene Hirka Formation. The Hirka Formation is mainly composed of bituminous shale, claystone, marlstone, tuffltuffite, and dolomitic limestones. Although numerous trona beds are interbedded within the above mentioned sedimentary units, there are only two beds of trona that are of economic interest. The lower part of the lower trona bed is rich in double carbonates such as pirrsonite (NA2C03? CaC03?2H20) and gaylussite (Na2C03?CaC03?5H20), indicating a change in Ca/Na ratio in the solutions which are responsible for the formation of soda occurrences. In contrast, the upper part of the upper trona bed is rich in nahcolite. This minerological pattern has been found in every drill- hole in the basin. While both trona beds average up to 10 m (33 ft) thick, the average thickness of nahcolite is approximately 0.5 m (1.6 ft). Also, thin section studies show that many fibrous and lath-like trona crystals contain small radial nahcolite crystals; all of them are euhedral and equi-dimensional. Geochemical studies of major, minor, and trace element contents also support this coexistence. Macroscopic and microscopic observations reveal that the nahcolites principally were formed as a result of an incongruent dissolution process of trona. In conclusion, the presence of nahcolite in the Beypazari basin is due to the dissolution process of trona, particularly during the last stage of evaporation, as a result of fresh water entering the basin. During that time, an increase in the HC03/ C03 ratio resulted in nahcolite-trona phase transitions.

Jan 1, 2001

By Jerome C. Motz

Los tipos, tamaños y número de trituradores empleados en un sistema completo de reducción varían con factores tales como: el volumen de mineral que debe ser procesado, tamaño de la mina, tamaño máximo del todo-uno, dureza del mineral y de su roca soporte y tamano requerido para el producto final. Básicamente, sólo existen tres tipos distintos de equipos de trituración para la etapa primaria. Estos son: machacadoras de mandíbulas, trituradores giratorios y molinos de impactos, y cada uno tiene unas características operativas propias. Las trituradoras de mandíbulas operan por aplastamiento de las rocas entre dos paredes, una fija y otra móvil, en una cavidad determinada. Han sido ensayadas variaciones en la caída y movimiento, pero la mayoría de las máquinas del tipo Blake tienen un ángulo de trituración de aproximadamente 27° entre las mandíbulas fijas y móviles. El principio de trabajo del triturador giratorio es muy simple. Si un cono es montado en un eje vertical y la parte superior del eje se sitúa estacionaria mientras que la parte inferior gira excéntricamente, el cono tendrá también un movimiento excéntrico. Si el cono se coloca en una carcasa conveniente se movers aproximándose y retrocediendo de las paredes que lo contienen a la vez que gira. Si el cono y las paredes de la carcasa son suficientemente robustas y pesadas, cualquier cosa apresada entre ellas será triturada. Al lado de los incrementos de la eficacia desarrollados por las continuas acciones de trituración y el curvado de las caras de trituración, los trituradores del tipo giratorio tienen las siguientes ventajas: 1. La gran abertura de alimentación disponibles en comparación con trituradores de otro tipo. 2. El alto rango de tamaños y capacidades, que se encuentran entre 600 y 6.000 toneladas por hora. 3. Generalmente no es necesario control en la alimentación. Dumpers de 150 a 200 toneladas descargan directamente en las actuales instalaciones.

Jan 1, 1982

By John F. M. Sims, Charles B. Green

Introduction The gold rush to the Canadian Klondike and then to the Alaskan gold camps is legend to Americans. The early years of the century also saw flourishing hard rock mining at the famous Treadwell and Alaska Juneau gold mines and the bonanza Kennecott copper mine. Over the last two decades, rich discoveries of copper, cobalt, zinc, lead, molybdenum, tin, and silver have furthered Alaska&apos;s reputation as a storehouse of US minerals. Alaska may also contain large commercial deposits of chromite, platinum, and nickel. Yet in 1984, with the exception of the Usibelli coal mine and several hundred small placer gold mines, Alaska&apos;s mineral potential remains largely undeveloped. Uncertain 70s The 1970s were years of uncertainty for the mining industry in Alaska. The discovery in 1968 of a huge reservoir of oil and natural gas in Prudhoe Bay brought the land claims of Alaska&apos;s natives to national attention and focused interest in Alaska&apos;s resource wealth. The Alaska Native Claims Settlement Act of 1971 (ANSCA) created the regional native corporations and granted them selection rights to some 178 000 km2 (44 million acres) of Alaska&apos;s 1.5 Gin&apos; (375 million acres) - about 12% of the state&apos;s land area. The native land selection process would take years. During this time, the status of much of Alaska&apos;s land was in question. Additionally, section d(2) of ANSCA provided for the conversion of an undetermined amount of federal lands into conservation units. These consist of parks, scenic rivers, national monuments, and wildlife refuges. The debate over the d(2) lands was long and acrimonious. National conservation groups lobbied to include as much acreage as possible within restrictive classifications, regardless of its mineral or energy potential. The process was polarized into a preservation-versus-development struggle. To many Alaskans, what was at stake was the ability to develop a viable independent economy, rather than an ongoing role as economic wards of the federal government. The resulting compromise was the Alaska National Interest Lands Conservation Act of 1980 (ANILCA). ANILCA resulted in the placement of some

Jan 1, 1985

By Bruce K. Cantrell

As part of a continuing study of diesel exhaust aerosol characteristics, the Bureau of Mines has applied Chemical Mass Balance (CMB) model source apportionment analysis to evaluate a size selective sampling method for determining the separate diesel exhaust and coal dust aerosol concentrations in a diesel equipped coal mine. Size selective sampling was used under the premise that diesel exhaust aerosol dominates the submicrometer aerosol size range, while most of the mineral dust aerosol is greater than 1 m. A micro-orifice, uniform deposit impactor (MOUDI) was used as the size selective sampling method. Concurrent samples of mine aerosol were collected using a dichotomous, respirable aerosol sampler. This sampler separated the fine and coarse size portions of its sample at 0.7 m. Trace element concentrations for in-mine aerosol sources and the dichotomous samples used by the CMB model were measured using Instrumental Neutron Activation Analysis. The CMB analysis, applied to both the fine and coarse fractions of the respirable mine aerosol, yielded the mass concentration contributed to each by diesel exhaust and mineral dust aerosol sources. The MOUDI size distribution data were analyzed using nonlinear regression to parameterize the fine and coarse fractions of the sampled aerosol and to estimate the contributions to each fraction from both coal dust and diesel aero- sol sources. For the diesel mine visited, analysis of the MOUDI samples yielded a value for the fraction of the fine aerosol attributable to coal dust of 15[]7% fine fraction coal from analysis of the MOUDI samples versus 25[]4% from the CMB analysis. The remainder of the fine aerosol in the mine was diesel exhaust aerosol. CMB analysis confirms the original premise for such sampling. After correcting for the presence of coal dust in the fine aerosol fraction, analysis of the MOUDI size distributions yielded estimates of respirable mineral and diesel exhaust aerosol concentrations that were accurate to with-in 10% of values determined by the CMB apportionment.

Jan 1, 1987

By C. S. Caldwell, L. C. Thompson

Mine closure and reclamation issues have evolved over the past several years becoming further complicated by stricter standards and by the cmrent economic climate and price of gold. Many mine permits are calling for more stringent remediation of spent ore heaps and tailings. Leachate solution water quality must meet strict discharge criteria for multiple parameters besides cyanide at mine closure. Closure standards are more difficult to meet and most mines do not have an increased budget at the end of mine life to deal with the falling standards. These issues make additional gold recove1y at the end of mine life a critical problem in financing some closure operations. Another issue is the length of time that must be budgeted for heap rinsing operations. Innovative biological processes developed and applied by Pintail Systems, Inc. at numerous mine closures have indicated that bio-processes can significantly enhance gold recovery dur­ing heap cyanide bio-detox over the recoveries expected from rinse operations. If heap rinsing can be eliminated or reduced and if gold recovery can be improved at the end of mine life, many mine closures may be possible with processes that pay for themselves or generate profit. Additional advantages of the bio­detox technology is that it can be engineered to treat multiple contaminants besides Weak Acid Dissociable (WAD) cyanide. Other contaminants treated in heap process solutions include nitrates, ammonia, heavy metals, thiocyanate and total cyanide. At the Cripple Creek and Victor Mining Company Ironclad Mine near Victor, CO, 5 million tons ofrinsed, heap leached spent ore were biologically detoxified to reduce cyanide content, enhance gold recovery during bio-detox and to reduce dissolved metals in heap leachate solutions. During the bio-detox treatment period a substantial improvement in gold recovery allowed the detox oper­ation to operate as a profit center. Final gold recovery was at least twice what was expected from a normal rinse operation.

Jan 1, 1998

By Daniel W. Kappes

The 1.8 million tonne Itos tailings deposit is the result of sever­al years of processing by Comibol of the San Jose and ltos Mine ores at Oruro Bolivia. The historic flowsheet consisted of a flotation circuit in which silver-bearing galena and stibnite were concentrated. Approximately half the silver in the ore was not recovered, and the existing tailings contain slightly over 200 grams silver per tonne, in the form of complex sulfides of silver, antimony, lead, and other metals coating grains of pyrite. An extensive laboratory program was conducted to exam­ine conventional processing l)lethods such as cyanide leaching and regrinding followed by flotation. However, the silver-bear­ing sulfides are in intimate contact with the pyrite and the pyrite comprises 30% of the total tailings weight. Cyanide leaching recovers only 40% of the silver and the cyanide consumption is uneconomically high. Flotation produces an upgraded, but still uneconomic, concentrate. For this reason, chloride leaching had been evaluated by Comibol, and in the mid 1980's Comibol undertook an innova­tive test consisting of two 300 tonne field heap leaches using a cold chloride solution. The tests were partially successful, but could not be applied to the bulk of the tailings which are already ground to 50% minus 200 mesh. In 1992, Kappes, Cassiday & Associates undertook to develop a hot agitated chloride leach flowsheet, and is current­ly constructing a treatment plant under a turnkey contract from a Bolivian company, Baremsa, S.A. Using this process, 85% of the remaining silver can be recovered, with significant by-prod­uct credits from antimony and lead. The project features hot, agitated acidic chloride leaching, filtration, crystallization of lead chloride, and precipitation of metals using iron scrap. Arsenic is removed from the precipitate by caustic leaching, and minor metals by furnace refining. Antimony is recovered as the oxide by cupelation, and silver is poured as <lore from the cupel. Production rate will be 600 tonnes per day. Startup is scheduled for late 1995.

Jan 1, 1996

A Mineração da Trindade (Samitri) is a major Brazilian iron ore producer. Founded in 1939, privately owned Samitri operates six mines with a total production capacity of 14.5 Mt/a (16 million stpy). The mines are located in southeastern Brazil in the state of Minas Gerais. They are operating in the iron ore quadrangle, Brazil&apos;s primary iron ore-producing region. In addition to iron ore production, Samitri has the capability to produce 350 kt/a (386,000 stpy) of ferruginous manganese ore from two mines, also located in the iron ore quadrangle. Historically, about 70% of Samitri&apos;s production of lump ore, sinter feed and pellet feed is exported. Destinations include Belgium, Luxembourg, France, Germany, Trinidad and Tobago, China, Turkey and the United States. Alegria Mine In September 1989, Samitri started a new processing plant to treat itabirite, a low-grade iron ore. The goal of the plant was to produce sinter feed (coarse fraction) and pellet feed (fine fraction). Primary separation of sinter feed from pellet feed was accomplished with conventional vibrating screens making a 0.4-mm (40-mesh) cut. The screen oversize was stockpiled as sinter feed. The undersize was deslimed with hydrocyclones and upgraded by conventional froth flotation. Resulting pellet feed grades ranged from 3% to 3.5% Si02. A change in pellet feed requirements by Samarco Mineração SA, a Samitri subsidiary, called for Si02 levels below 1 % and increased production levels. Moreover, market demands called for a change in sinter feed size distributions. Unfortunately, lower pellet feed Si02 levels could not be achieved with the existing process. This was because of the presence of 35% +0.15 mm (+100 mesh) in flotation feed. Flowsheet development. Laboratory and pilot-plant studies conducted at Samitri&apos;s research facility found that to achieve lower Si02 levels, the +0.15 mm (+100 mesh) in flotation feed must be kept below 5% to 10%.

Jan 1, 1997

By Nenad Spoljaric

Greensand, greensand marl, and green earth are names given to sediments rich in the bluish green to greenish black mineral known as glauconite. The word glauconite is derived from the Greek word glaukos, meaning bluish green. The term "greensand" as a rock name for a glauconite-bearing sediment is more appropriate than "greensand marl," a term that has been doggedly perpetuated in the literature. Because of its potash and phosphate content, greensand was mined and marketed as a natural fertilizer and soil conditioner for more than 100 years. The advent of manufactured fertilizers with adjustable nutrient ratios led to a decline in the use of greensand in agriculture. The material has since been recognized as useful in water treatment. Unfortunately, despite large reserves and world- wide distribution, glauconite has not been utilized to any significant commercial extent because no major application has been found for a substance with its chemical composition and properties. This is probably due mostly to a paucity of research on its potential commercial uses. Extraction of potash received considerable attention during and just after World War I. Because of relatively high extraction costs and a generally low potash content (viz., less than 8%), glauconite lost its appeal as a source of this commodity. Historical Background Greensand was used as a fertilizer in New Jersey in the latter part of the 1700s. During the early 1800s its use became more common; applications of as much as 22.5 kg/m2 were sometimes made, although recommendations for agricultural use suggested 4.5 to 11 kg/m2 (Tedrow, 1957). Many crops, especially the forage type, were said to improve with greensand application; however, because of its slow release of potash, large quantities were required. Certain greensands that contain sulfur and sulfide minerals are harmful to plant growth, and these were classified as poison, burning, or black marls. The availability of higher grade potash salts from other mineral sources and the manufacture of prepared fertilizers displaced the agricultural use of greensand during the latter 1800s. During the mid-1800s the greensand industry, centered in a small section of the eastern United States, grossed more than $500,000/y. Toward the end of the century, however, annual production had dwindled to less than $100,000 in value. By 19 10 there were only six or eight greensand producers grossing less than $5,000/y each (Tyler, 1934). There was a brief revival of the US industry during World War I because of the curtailment of foreign potash, especially from Germany. During the latter 1940s and early 1950s greensand was again recommended as a food nutrient for plants and farm crops. Agronomic studies discussed its potential as a soil additive that gradually releases potash and many trace element nutrients essential for plant growth (Tedrow, 1957). Greensand was sold with the idea that it would condition soil and absorb and hold water while its base exchange properties would release trace elements. For a short time glauconite was used in certain parts of New Jersey as a binding additive in the brick industry, and in the 1800s it was used for making green glass (Cook, 1868). In the early 1900s the base exchange properties of glauconite were recognized for water treatment and the mineral gained acceptance as a water softener. Mansfield (1922) does not mention base exchange even though this phenomenon was known in 1916 or earlier. From 1916 through 1922 several patents for the use of glauconite as a water softening agent were granted. A method was also patented for treating greensand to improve it for water softening and ready regeneration with common sodium chloride brine (Borrowman, 1920, Spencer, 1924, Kriegsheim and Vaughan, 1930). Treated glauconite, on contact with water containing magnesia or lime, takes up magnesium or calcium ions and releases sodium ions. This exchange is limited to the outer surface of glauconite grains, and when all the surfaces have absorbed their capacity, the grains must be regenerated. Regeneration, simply stated, consists of treating or backwashing the glauconite with a sodium chloride solution, which replaces the hard water elements with sodium, thus reviving the glauconite. The process has become more sophisticated due to competition among companies in the water softening business. Greensand products for water softening generally consisted of several different grades distinguished by the particular treatment the glauconite was given during processing. The standard greensand water softener was produced from natural glauconite that was only washed and classified. Its characteristics for water softening are given in [Table 1].

Jan 1, 1994

By George S. Austin

Mud is one of the oldest building materials used by man, but it is not only of historic interest. Even today in many parts of the world, particularly lightly forested areas, it is the chief building material. Many authorities believe that at least 30% and perhaps 50% of the world&apos;s present population lives in earthen dwellings (Dethier, 1985, Coffman et al., 1990). Not all of these people are members of primitive societies or live in third world countries. European, African, Asian, and both North and South American countries contain a large number of such structures. In the United States, the American West, and particularly the Southwest, is known for this type of construction. Some of the earliest remains of adobe structures are those discovered in the ruins of Neolithic farming villages in Mesopo¬tamia dating as far back as 7000 BC (Steen, 1972). The word adobe has its roots in Egyptian hieroglyphs denoting brick and evolved to its present form through Arabic and Spanish (Lumpkins, 1977). The Spanish conquest of the New World spread the use of wooden molds to produce a standard adobe brick. Today, the word adobe is used to describe various earth building materials and techniques, usually referring to sun-dried adobe brick now used in the United States, but is also applied to puddled adobe structures, mud-plastered logs or branches (Jacal or waddle-and-daub), pressed-earth blocks, and rammed-earth walls or pisé (Smith and Austin, 1989, Ferm, 1985). Modern mud construction is used in many countries in various parts of the world. In the United States, the states from Texas to California are perhaps best known for this type of construction. Of these states, New Mexico has the dominant reputation for adobe use. Indeed, in New Mexico the Santa Fe style (Fig. 1) has made adobe not only acceptable, but preferred. RAW MATERIALS Adobe soil used by present day adobe producers, and probably past adobe producers as well, is principally sandy loam (50% clay and silt), although clayey silts are used in some areas (Coffman et al., 1990). In New Mexico, the best adobe soils are those developed on stream deposits, particularly Holocene (Recent) terrace deposits and older, loosely compacted geologic formations, such as the Santa Fe Group (Tertiary) located in the Rio Grande Valley. Some modern adobe producers use a mixture of materials from the screened fines of aggregate operations and mud from irrigation ditches in the river valleys combined with varying amounts of sand to produce the proper blend. Mineralogy Bulk Mineralogy: X-ray diffraction analyses of whole rock samples from many parts of the world where adobe is the dominant construction material show the major constituents of adobe are quartz and feldspar, with lesser amounts (in order of abundance) of calcite, clay minerals, and gypsum (Coffman et al., 1990). Adobes from and climates contain considerable calcite; in some cases calcite is second only to quartz in volume. Quartz, feldspar, most of the clay minerals, and some calcite commonly are derived from the mechanical/chemical breakdown of older rocks units. Some of the clay minerals, much of the calcite and perhaps all of the gypsum are precipitated from evaporating water. Clay Mineralogy: Clay-size minerals consist dominantly of clay minerals, but nearly clay-size fractions contain minor amounts of quartz and calcite, and occasionally other nonclay minerals. In earth construction materials from New Mexico, clay-size particles are the most compositionally variable in commercial adobe soils. However, the clay mineral groups in this size fraction consist of about equal parts of expandable clay minerals (smectite and mixed¬layer illite/smectite or I/S) and non-expandable clay minerals (ka¬olinite, illite, and chlorite), with minor quartz, calcite, and feldspar (Smith and Austin, 1989, Austin, 1990). In the and American Southwest, smectite is commonly calcium-rich and the I/S is dis¬organized and randomly interstratified. In a study of 42 New Mex¬ican commercial adobe soils, only two contained chlorite, and vermiculite, sepiolite, and palygorskite were not found. Soils in temperate climates, as are found in the midwestern and northeastern United States, contain an abundance of illite (Potter et al., 1975). The clay minerals in earthen structures of that region and in Europe commonly contain illite with less amounts of kaolinite, smectite, I/S, chlorite, and vermiculite. In contrast, the clay-size fractions of soils of humid tropical areas are typically acid, with kaolinite as the dominant clay mineral and lesser amounts of I/S, illite, smectite, and others (Chamley, 1989).

Jan 1, 1994

By W. E. L. Minter

Discussion by R.W. Hutchinson Minter&apos;s excellent recent paper (1990) on the great paleoplacer gold deposits of the Witwatersrand and its subsequent discussion (Cheney, 1991) and reply (Minter, 1991) provide an informative, interesting and up-to-date review of the geology of these unique ores and some long enigmatic questions regarding their origin. Clearly summarized by Minter (1990) are the compelling sedimentologic and stratigraphic relationships indicating that the Central Rand Group conglomerates are ancient placers and documenting the fluvio-deltaic conditions of their transport and deposition. As Minter (1991) indicates, Cheney&apos;s discussion broadens Minter&apos;s 1990 coverage of the subject by considering higher and lower stratigraphic units and by adding detailed comments concerning recently published information about metamorphic and age relationships. Cheney&apos;s discussion and Minter&apos;s reply also readdress the old controversy of epigenetic hydrothermal vs. syndepositional placer origin for these deposits. This controversy has been resurrected by: • recent recognition of hydrothermally-altered granites (HAGS) in the older Archean basement rocks that lie unconformably beneath the Dominion Group rocks north and west of the Witwatersrand depository; • recognition of greenschist metamorphic assemblages in the Witwatersrand strata; and • the presence of interpreted epigenetic mineral assemblages and textures in some of the deposits. These relationships are briefly discussed by Cheney who suggests they may result from later, superimposed metamorphism and deformation and do not prove that the gold was introduced by epigenetic/metamorphogenic processes. In his paper, Minter (1990) cited the 1986 abstract of a subsequently published paper by Hutchinson and Viljoen (1988) who also considered many of these aspects and attempted to integrate them all in abroad genetic hypothesis. In addition, they emphasized the additional important problem of determining the source for the Witwatersrand gold. Although this issue was partially addressed by Reimer (1984), Mossman and Dyer (1985) and Reimer and Mossman (1990), the question has been generally under emphasized in earlier work and was, therefore, not stressed by Minter and Cheney. Reimer (1985) also emphasized the importance of weathering of the basin&apos;s Archean hinterland and its affects on Witwatersrand sedimentation. Since the epigenetic hydrothermal hypothesis was essentially discarded in favor of placerist origin, the gold source for these great deposits has generally been attributed to the weathering of lodes in greenstones or granites of the uplifted, older Archean hinterland to the north and west. However, Hutchinson & Viljoen (1988) present evidence that this explanation is inadequate, as well as data in support of an alternative explanation. They suggest that the gold and accompanying abundant, clearly detrital and auriferous pyrite in the conglomerate reefs were derived by erosional degradation of auriferous-pyritic exhalite previously deposited, by shallow marine discharge of hydrothermal fluids, along the fault-bound northwestern margins of the Witwatersrand depository. The auriferous-pyritic exhalites were generated by fluid-rock reactions in the subsurface, particularly in the highly-altered lavas of the Dominion Group but also in adjacent basement granitic rocks. The process envisaged is comparable to that observed in sea floor hydrothermal systems today, although, in this case, it occurred at the margin of a rapidly subsiding, shallow marine or continental basin. Important evidence for this explanation is the anomalous gold content of the ferruginous, shaley strata occurring throughout the Witwatersrand Supergroup and including the distinctive contorted bed that is an iron formation. These layers have been interpreted (Hutchinson and Viljoen, 1988 and 1990) to represent the more distal, finer-grained, mixed chemical/clastic sedimentary strata. Thus, they are the basinal equivalents of the proximal, auriferous-pyritic exhalites that were deposited along the basin&apos;s margin and subsequently reworked by sedimentary processes. This gold source explanation resolves many of the enigmatic questions regarding Witwatersrand geology and the reawakened controversy between proponents of hydrothermal and placerist genetic hypotheses. The theory invokes a hydrothermal origin for the gold in the basin-margin pyritic exhalites, thus explaining the visible hydrothermal characteristics (Figs. 1 and 2). It also provides compelling evidence for placer origin by encompassing all of the fluvio-deltaic processes of clastic sedimentary transport and deposition that are clearly summarized by Minter. These characteristics result from the degradational reworking of the pyritic exhalites along the depository&apos;s margins. Fluvial reworking, transport and deltaic redeposition of this detritus, along with additional detritus carried into the depository from the hinterland, then formed the conglomerate reefs. Other enigmatic aspects of Witwatersrand geology include: • the rarity of magnetite and ilmenite (the two most common heavy minerals in placer deposits) and the contrary abundance of leucoxene in the conglomerate reefs; • the presence of differing varieties of pyrite (authigenic, allogenic, compact and porous) all of which are auriferous; and • the presence of well-rounded pebbles, even cobbles, of pyrite that, in view of the brittleness of the mineral, could not have been transported great distances from a hinterland.

Jan 1, 1993

By Pierre Longtin

"Replacing the aging roadway access to the iconic Golden Gate Bridge, thereby ensuring traffic safety and robustness in case of seismic occurrences, provided an opportunity for major design improvements as well as the development of a parkway that would include the construction of four new tunnels. With an eye to the future, the Doyle Drive replacement project aimed to meet the highest standards for environmental stewardship beyond regulatory requirements. Understanding that lighting equipment is the tunnel‘s subsystem that uses the most energy, the operator was looking for a sustainable solution to light the tunnels. An Intelligent Lighting Control System (ILCS), designed specifically for road tunnels, was acquired to individually control the almost 2000 luminaires lighting the four tunnels. This paper will show the results in terms of energy and operational savings generated over the nine months of operation between the commissioning and the conference in April 2016. INTRODUCTION Electrical & mechanical (E&M) equipment represents an average of 10 % of a tunnel total construction cost. Once the civil infrastructure completed, E&M equipment drives the main part of the expense during a tunnel useful life. Table 1 below describes the impact of various types of energy consumption and their cost impact in tunnels of less than 500 m. It is clear that lighting is the sub-system that requires the most energy. A good lighting strategy is the key to reducing energy consumption due to lighting and a viable solution that will bring important savings in tunnel operations.With a price of $0.14/kWh in California and a planned increase between 3% and 6% per year during the period 2013-2040, the tunnel operators could not afford to overlit the tunnels—particularly when they have to operate it for the next 30 years. However, they also need to maintain the right lighting level to ensure driver safety all the time. The Presidio Parkway project includes the Battery tunnels, which are 240 m (790 feet) and 230 m (750 feet) in length, and the Main Post tunnels, which are 280 m (920 feet) and 315 m (1030 feet) in length."

Jan 1, 2016

By James R. Craig

The Central and Southern Appalachians were the major sources of gold in the United States throughout the first half of the 19th century and have continued to yield gold to the present day. Small amounts of gold were probably extracted by the Indians and the Spaniards but the first documented report of gold in the region was made by Thomas Jefferson who found a four pound lump of gold ore, which yielded 0.85 troy oz., in the Rappahannock River of Virginia in 1782. Gold was subsequently discovered in North Carolina between 1793 and 1799, in Virginia in 1806, in South Carolina in 1827, in Georgia in 1829 and in Alabama about 1830 (Becker, 1894-5; Pardee and Park, 1948). The gold belts of the Central and Southern Appalachian Piedmont and Blue Ridge Provinces are shown in Figure 1. The superimposed circles indicate the mount of production from some of the more productive and well known areas. Although North Carolina produced far more gold than the other states, the most significant concentrated area of production is the Dahlonega District in Georgia which yielded more than 500,000 troy ounces. After discovery, gold production increased rapidly as shown in Figure 2. Production was at its maximum between 1830 and 1852 with combined totals generally exceeding 30,000 troy ounces per year. Production dropped sharply after the California gold rush in 1849 as experienced miners moved westward in search of bigger and easier stakes. After the Civil War briefly halted gold mining in the Appalachians, production rose to an average of about 8,000 ounces per year from 1865 until 1878 and to an average of about 13,000 ounces per year from 1881 until 1915 when it again fell to near nil (Pardee and Park, 1948). Minor resurgences occurred in the 1930&apos;s and 1940&apos;s. Although there has been no reported production in recent years, it is likely that "weekend panners" still remove hundreds of ounces from streams annually. The total reported production for the area through 1959 is in excess of 2.5 million troy ounces.

Jan 1, 1981

By R. L. Taylor, S. D. Abbott

"INTRODUCTION Mines are filled with a mixture of gases, some of which are toxic, flammable, explosive. Others can be early reporters of potential problems. It has previously been shown that analysis by gas chromatograph can the most accurate conventional analytical tool for common mine gas mixtures when used correctly. There are some other detection methods (e.g hand held sensors), but they do not detect many situations which are immediately dangerous to life or health (IDLH). Examples include explosive hydrogen mixtures, toxic carbon dioxide levels, and some explosive mixtures of methane with oxygen. (Reference our previous paper). There are several reasons for mines to consider onsite GC analysis. (1) Timeliness of results. Samples can be taken and analyzed immediately. Samples do not have to be shipped to a lab and wait for results. (2) Ability to run greater number of samples a lower for the same cost. This allows mine to increase the number of samples taken to be more informed about the gases/conditions of the mine without the worry of financial constraints. (3) Fewer sample media needed, and there is no wear and tear or cost from shipping them offsite. When samples are shipped offsite, those sample media cannot be used to take until the lab returns them when the analysis is done onsite, the sample media can be used again almost immediately. However there are several challenges associated with using GC to analyze mine gas samples onsite. To date there has been a clash of the requirements for a successful analysis, the capabilities of the technology, and the practicalities of implementation. The purpose of this paper is to discuss some of the difficulties in getting accurate reliable data and doing so in an economic way. This whole measurement activity is a cost center for a mine, so there are economic concerns for the mine to address. These concerns are mainly staffing costs, capital costs, maintenance, and consumables costs. A properly staffed and equipped analysis lab, including burdened overhead, costs about $500K per year. At a sampling rate of 10 per day, this is about $140 per measurement for a proper onsite lab."

Jan 1, 2016

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 F. C. Bond

History The breaking and shaping of rock was one of man&apos;s earliest occupations. In the Paleolithic Age long before the dawn of history, arrow¬smiths and the makers of stone axes, hammers, knives, scrapers, spears, and borers were highly respected members of society. In early historical times stones for building blocks, roads, and city walls were shaped by slaves and convicts, who also did most of the mining. However, great artists erected beautiful stone sculptures, while gifted architects planned imposing temples and monuments. Until well into the 19th century nearly all rock was broken laboriously by hand. The small rock required by John MacAdam for his macadamized roads in England in the 1820s was produced by women and boys seated alongside the roadside with hand hammers and legs wrapped with rags. Eli Whitney Blake, a nephew of the Eli Whitney who invented the cotton gin, developed the first successful jaw crusher before 1870. The gyratory conical crusher soon followed. Comparative tests established its large capacity advantage over the jaw crusher, as well as its greater cost for a given feed size. Both types have been in use for more than 100 years. Crushing rolls appeared before 1900. Thomas A. Edison made very large diameter rolls which were excessively long; they failed because of shaft deflection. Various types of disk crushers and edge runner rolls appeared about this time. The older methods of reducing rock were adaptations of other processes. The stamp battery of dropping weights effected crushing by simulating heavy hammer blows; the much earlier arrastre, in which heavy stones were dragged in a circular path over the ore by animal power, came from the prehistoric method of grinding grain between two rubbing stones, while the jaw crusher was adapted from simple squeezing devices. But the tumbling grinding mill was not just an adaptation; it was an invention, because it required thinking on a somewhat higher order-there was no prototype. Its nearest antecedents were probably the small closed tumbling drums used in England more than a century ago for cleaning and polishing small iron castings. The date of the first tumbling mill actually used to grind rock is unknown, but it was later than the American Civil War (1861¬1865). It was almost certainly a closed or batch mill in which rock was placed and rotated until it reached the desired particle size. It could have been operated either wet or dry. The first published refer¬ence to such a batch mill was one introduced by Alsing in England (1870) for the grinding of calcined flints for pottery work.21 There are several rather indefinite reports of grinding mills in the early 1890s, including an overflow ball mill in the Helena and Livingstone reduction plant in Montana which may have been the first of its kind. 11 Many of the first mills, which were called tube mills, used hard rock both as grinding media and as mill lining. The rock used was preferably stone from the Normandy and Danish beaches, when it could be imported. This remarkable siliceous stone was already widely used for grist mills throughout America, and its resistance to wear was greatly respected. The decade of the 1890s saw the development of tumbling mills with continuous feed and discharge and their extension into different industrial uses. By 1895 some experience had been accumulated. Iron grinding balls were being tested and the proper speed of rotation was being determined. The Clark Patent tube mill was featured in an E. P. Allis bulletin of 1890, which may have been the first published description of a tube mill. More than 1,000 Gates tube mills had been built by Allis-Chalmers before 1913. Many of these were used in gold mining, espe¬cially in South Africa. The 5 x 22-ft size was particularly favored for grinding portland cement; the use of tumbling mills in the manufac¬ture of cement began about 1900. A great deal of attention was paid to the mill lining. Metal was then relatively expensive, and the general approach was to trap some of the rock grinding media into mill lining pockets. This rock would then absorb the wear and protect the metal lining. In the first ten years of the 20th century there were several different types of pocketed liners, with different manufacturers advancing the superior claims of their patented arrangements. The Osborne liner, developed in South Africa, was probably the most successful. 21 Another item which attracted much attention was size classifica¬tion within the mill and in ancillary equipment attached to the mill and rotating with it. The Krupp type, with interior screens protected inside the mill lining, was developed quite early in Germany, possibly before 1890. The Dorr reciprocating rake classifier (1907) had not yet been invented, and many strange and impractical screening and classifying devices were proposed. In these unsatisfactory machines the two separate processes of size reduction and classification were combined into one operation. It was many years before recognition came that a machine is most efficient when it is designed for one specific purpose. There was much industrial wastage before the opera¬tions of grinding and classification were finally separated. After 1900 the grinding of portland cement raw material and of cement clinker required large numbers of tumbling mills. Most of the raw material was then ground dry. This was also the heyday of gold mining. The old stamp mills that were used in great numbers for grinding gold ore did not grind sufficiently fine to liberate all of the gold, and the new tube mills were installed following the stamps. After 1910 larger diameter tum¬bling mills with larger grinding media were developed. These could receive the finely crushed ore directly, and the inefficient stamp batter¬ies were gradually eliminated. The Rand in South Africa was the greatest gold producer, treating immense quantities of rather low-grade but consistently free milling gold ore. The first tumbling mills, or tube mills, went into operation there in 1904. They were so successful that within a few years no new stamp batteries were installed in the district, even though old ones continued to pound away until after 1950.3 The early tube mills on the Rand all employed Normandy or Danish pebbles, which had to be imported at considerable expense. Their reported wear was as low as 4 lb per ton ground. 4 Many of the mills were lined with the same tough Danish stone cemented into place, while others used the pocketed steel Osborne liner. It was in 1907 on the Rand that an important test was made using hard native ore for grinding media in place of the expensive imported pebbles. 3 This ore, called banket, did not wear as well as the renowned Danish pebbles, but the cost per ton of grinding was definitely reduced. Many of the tube mills on the Rand were soon grinding with native ore. This was the beginning of the development now called autogenous grinding, in which the ore grinds itself. This is treated under a separate heading. See Subsection 3C, Chapter 4. Gold mining was important in America also, and its grinding history follows that of the Rand. Danish pebbles were replaced by native ore in Santa Gertrudis, Mexico, in 1913, 4 and in Consolidated Gold Fields, Nevada, in 1914. 5 Other properties followed suit. However, it gradually became apparent that the capacity of a given mill could be almost doubled if rock grinding media were re¬placed by cast iron or steel grinding balls. In order to increase plant grinding capacity many rock media mills were converted to iron grind¬ing media in the second decade of the present century. In some mills the motor size was doubled; other mills were cut in two and another motor was provided for the second half. Grinding mills began to assume a more modern appearance. Crushing rolls were formerly much used following crushing in jaw or gyratory crushers and preceding grinding in ball mills. How¬ever, the roll surfaces wore rapidly, and skilled maintenance was required to obtain even wear. Rod mills could take feed of the same crusher product size and reduce it finer. The first rod mill was con¬structed by Mine and Smelter Supply Co. and was tested in Canada

Jan 1, 1985

By Gus H. Goudarzi

The apatite deposits in east-central Bahia, Brazil, are about 50 km east of Serra do Jacobina, between lat. 11° and 12° S. and long. 39°40&apos; and 40° W. Apatite was found and examined at 15 localities in a mineralized belt that strikes about N.25° - 30 W. over a distance of about 100 km, but the size of the deposits has not yet been determined. Rocks of the region consist of a sequence of Precambrian migmatites and quartz feldspar paragneisses containing units of pyroxene-rich rocks, lenses of sillimanite-bearing quartzite, amphibolite zones, and calc-silicate marble lenses. Pegmatites and granites of probable anatectic origin (anatexite) are abundant. Faults are numerous. Clastic sediments of Tertiary and Quaternary ages cover most of the area; thus, outcrops of the Precambrian rocks are sparse. Trenches and pits have exposed mineralized zones as much as 10 m wide at some localities within the region. A layer of apatite at least 1 ½ m thick is exposed in a trench at Lagoa do Mel. At Rio do Peixe, vertical shafts have penetrated a mineralized zone, to a depth of as much as 13 m, of which about 25 percent is apatite. Apatite is massive or forms minute crystals, stringers, and disseminated grains in alkalic host rocks. At Apolinario, euhedral crystals of apatite are as much as 50 cm long and 30 cm in diameter. Associated minerals include vermiculite, biotite, scapolite, opal, chalcedony, epidote, montmorillonite, chlorite, kaolinite, sericite, magnetite, and limonite. Apatite is everywhere intimately associated with pyroxene-rich rocks, alkalic syenite pegmatites, and alkalic granite pegmatites. At Panelas-Retiro, apatite is associated with pegmatites, crystalline calcite (vein calcite), vermiculite, and pyroxene-bearing rocks. Analyses of samples show that the apatite is radioactive and rare-earth-bearing. Traces of vanadium and titanium were detected by spectroscopic analysis. The apatite was probably derived from phosphatic zones in the original Precambrian sedimentary rocks and has been crystallized and some mobilized by high-temperature regional metamorphism (higher amphibolite facies) into veins, pegmatites, and along certain preferred beds and fractures. Radiometric and geophysical surveys, in addition to conventional geological investigations, are planned to further investigate the nature and economic potential of the deposits.

Jan 1, 1972

By Omar A. Muhtadi, David A. Milligan

7.1 INTRODUCTION The circulating fluid (generally an aqueous cyanide solution) is the major active component of heap leaching. This solution extracts precious metal from the heap and is then processed to recover the precious metal. The solution continuously recycles through the heap to recover more precious metal. With each cycle through the heap, water and chemicals are lost. Make-up water and chemicals must be added to balance losses from evaporation, seepage, entrainment, and chemical reaction in the heap, as well as any other processing losses. Adding more water than needed can result in a build-up of water as inventory in the various ponds and can eventually require that water be discharged from the operation. This is an extremely undesirable condition, and can often be very expensive in both investment in water treating facilities and operating costs. Most heap leach facilities operate under a zero discharge permit; thus, a situation requiring water discharge can result in permit violation. Chemical solution control must be managed to maintain the optimum pH, dissolved oxygen content, and cyanide content, while preventing "scaling". Calcium-based precipitation is a common scaling problem in mill reclaim waters. A buildup of these deposits can result in severe reduction of water flow, loss of pump efficiency and eventual interruption in metal production. In addition, problems of plugged sprinkler heads, blinded leach pads and fouled carbon (in carbon recovery systems) can be and often are experienced when scale control is not maintained. The remainder of this chapter discusses chemical solution control and scaling prediction/prevention in the context of a "closed" system. Sizing of ponds to accommodate precipitation and runoff, collection systems, and other similar design considerations are not discussed here. These topics are deferred to later chapters, e.g. Chapter 10 considers design of collection systems while Chapter 12 discusses pond sizing. 7.2 CHEMICAL CONTROL Chemical composition of the circulating solution must be controlled by the operator if the heap leaching process is to be successful. The primary or essential chemical control parameters in any heap leaching operation are:

Jan 1, 1988

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

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

Jan 1, 1980