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By R. Fraher

The control of mine roof is of vital interest in underground coal mining not only for reasons of safety, but also to avoid the high costs of cleanup, resupport and production losses that can result from roof falls. The difficulty in determining roof stability mine-wide is due in large part to the broad variation of geologic and structural conditions that typically exist throughout a mining area. To ensure a stable roof over the entire mine, support systems must be selected to minimize the possibility of failure. The magnitude and cost of the support selected is a trade off against the cost of the number of falls that can still occur after it is installed. Because of the large number of variables involved in analyzing and optimizing roof stability versus support, a digital computer simulation was selected to find a cost effective solution to roof support design. A total cost approach is used which includes the actual cost of the roof support as well as the cost of the roof fall in terms of resupport and loss of production (Fraher 1990a). A Monte Carlo simulation program called ROCSIM (Roof control Optimization Cost SIMulation) has been developed for determining the optimum roof control system and design layout for underground room-and-pillar mining.

Jan 1, 1992

By Bruce W. Cavender, Brian L. Granger

The leaching characteristics of non-homogeneous ores have traditionally been difficult to assess in laboratory tests. A result of the large sample size necessary to ensure representativity, this difficulty is most acute for ores containing nuggets or other irregular occurrences of precious minerals. Consequently, small-scale leaching tests are often regarded as indicative rather than conclusive. The predictive nature of leaching tests can be improved by applying the experimental design concepts of blocking and randomization. In the case of column leaching, experimental 'blocks" are constructed with the same number of columns as there are separate test conditions. These blocks are treated identically throughout the test, with differences between columns within each block being the variables of interest. To maximize the uniformity of the blocks, randomization procedures are employed in the preparation of the sample and conduct of the test. The initial ore sample is screened and individual column charges composited from randomly selected increments from each size fraction. The leaching order, daily solution analyses, and processing of leach residues are similarly randomized. The observed differences in leaching performance between the columns in each block are compared to determine the over-all outcome of the test. A statistical analysis using t distributions gives a quantitative measure of the reproducibility of the outcome.

Jan 1, 1995

By T. X. Ren, J. S. Edwards

Although much work has been carried out in order to understand the mechanism of spontaneous combustion and hence minimise its potential risk, the spontaneous combustion of coal continues to be a problem in coal mining. The Department of Mineral Resources Engineering of Nottingham University has been active in conducting fundamental research into the problem of spontaneous combustion of coal. For years standardised adiabatic oxidation tests have been used to identify coals and other carbonaceous materials liable to spontaneous combustion. Recently the apparatus has been computerised and coupled with an expert system known as ESSH to evaluate the potential of self-heating of coal on the basis of intrinsic coal properties, geological settings and mining techniques. A numerical model has been developed to study the aerodynamics of airflow in the wastes of longwall coal faces. The Computational Fluid Dynamics approach is being explored to simulate airflow behaviour and nitrogen inertisation processes within the goaf. Such a model would be used in locating the areas most liable to spontaneous combustion and improving the utilisation of nitrogen. It would also be possible to determine the most likely path ways of combustion/heating products so as to provide an indication of the most appropriate locations for fire detectors.

Jan 1, 1997

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

By Richard M. Crowell, John W. Shield, Todd J. Harvey

Facing lower gold prices, gold mine operators throughout the world are seeking ways to reduce costs for treating refractory sulfide ores. Many gold mines, currently producing oxide ores, have underlying refractory sulfide deposits and the operators could extend the mine life by processing these sulfides. Fortunately, oxidation technologies for refractory ores need not be complex nor costly. GeoBiotics, Inc. has developed heap biooxidation technologies (GEOCOAT™) for treatment of refractory gold ores and concentrates. GEOCOATTM uses simple low cost unit operations common to heap leaching of gold and copper ores. The cost effectiveness and flexibility of GEOCOAT™ heap biooxidation allows mining companies to expand operations to include refract01y ore. In addition, GEOCOAT™ can be utilized to supplement or replace existing oxidation processes such as roasting, autoclaving, or agitated tank biooxidation, and to treat a lower resource grade than is currently economic. The GEOCOAT™ technology involves coating sulfide concentrates onto sized support rocks and biooxidizing the sulfides in a heap configuration. After biooxidation the concentrate is stripped from the support rock, neutralized and leached in conventional cyanidation facilities. Ashanti Goldfields Company Ltd. (AGC) produces over 885,000 ounces of gold per year from oxide and refractory ores at its Obuasi operations in Ghana. Refractory flotation concentrates are treated in roasters and in the world's largest stirred tank biooxidation facility. In August 1998, AGC and GeoBiotics jointly constructed and operated a field demonstration heap of the GEOCOATTM process at the Obuasi operations. The heap contained I 082 tonnes of support rock and 125 tonnes of concentrate. The heap was approximately 19 meters square at the base and 3.8 meters high, and ran for 88 days after which the concentrate was reclaimed. Average gold recovery after biooxidation was 91.5% with a NaCN consumption of 8.6 kg/t. Baseline gold recovery by direct cyanidation of the concentrate was 53.8%.

Jan 1, 1999

By Gregg R. Bierei

Arch of Illinois is a mining complex comprised of two surface and two underground mines located in Perry County, in southern Illinois. The company a division of Apogee Coal Co. of St. Louis, MO. Arch of Illinois&apos; two surface mines, the Captain and Horse Creek Mines, use large stripping shovels and cross-pit bucketwheel excavators to remove overburden from three seams of coal, the Herrin No. 6, Harrisburg No. 5 and Danville No. 7. All are considered high sulfur reserves. The two underground operations are managed by Arch of Illinois. They are the Kathleen Mine and the newly constructed Conant Mine, which began production in 1991. Both underground mines are continuous miner operations. They mine the No. 6 coal, which is about 1.8 m to 2.1 m (6 to 7 ft) thick. All coal is currently transported by truck to a central preparation plant located at the Captain Mine. This preparation plant produces 27 kt/d (30,000 stpd) of clean coal. Production in 1992 reached 5.9 Mt (6.5 million st) of saleable coal. Of that, about 66% or 4 Mt (4.4 million st) are produced by the surface mines and the rest the by under ground operations. Change is in the wind for Arch of Illinois. In 1988, 100% of all production (6.2 Mt or 6.9 million st) came from surface operations. By 1995, only 38% of production is scheduled from surface mines. And by the year 2000, 100% of production will from underground mining methods. Dwindling low ratio reserves, compounded by increasing environmental and land acquisition costs, coupled with more efficient underground mining equipment, have started an evolutionary trend toward development of underground reserves. For Arch of Illinois to maintain profitability of remaining surface reserves and to compete effectively mining underground, it must improve efficiency, productivity and cost reductions. These efforts are essential for maximizing current surface reserves and to ensure a continued existence in the presence of clean air legislation and the softening of the high sulfur coal market.

Jan 1, 1993

By Robert S. Mayo

My subject today is "Drilling-and-Blasting of Small Rock Tunnels". I am sure that there will be some people, primarily college students but perhaps Consulting Engineers, who will say: "Drilling-and-Blasting? I thought that went out with the invention of the Tunnel Boring Machine". A few years ago I talked to Dick Robbins and he told me that they did not try to sell a "Robbin&apos;s Mole" unless the tunnel was two miles long. For easy calculations I use 10,000-ft. (3km) A Mole today costs about $2 million which works out to $200 per lineal foot. This is "front money". You cannot place the order unless you are the uncontested low-bidder and then you must wait at least six months for delivery. Of course you can use that time to square up your portals, ready to turn-the-eye, or maybe you have shafts to sink to reach the tunnel grade. But this is "dead-time" and it is a heck of a feeling to have a job and cannot get started on it until you get the Mole. I do not have the American figures on lengths of tunnels but there was an article a few years ago in a British magazine: NEW CIVIL ENGINEER, on this subject. This stated that 75% of the jobs in Britain which were bid were for small tunnels, 10-ft x 10-ft (3m x 3m) or smaller, and that they were less than 5,000-ft (1.5km) in length. This does not mean that 75% of the footage in Britain was less than 10-ft in diameter. It means that 75% of the jobs being bid were less than a mile in length. From my experience in the U.S. I believe that we have substantially the same ratio: so many of the jobs with which I was associated were short, 1,000-ft, 2,000-ft and sometimes 3,000-ft and the bores

Jan 1, 1981

Compared to alternative ways of dealing with coal fines, oil agglomeration has always been a commercial bridesmaid and never yet a bride. Yet interest is high in the process, as shown by a number of pilot plants in operation that are funded by industry as well as government. In the past in the US, alternative ways to handle coal finer than 100 to 200 mesh included flotation, or no treatment except dewatering, or discarding the fines as waste. There are now a number of factors that tilt the economics more toward fine coal recovery. Today, the cost of mining a ton of underground coal runs from $22-30, and discarding even 10% as unrecovered fine coal is a real economic loss. A number of attempts (notably by EPRI) have been made to quantify the effect of coal cleaning on utility boiler operations. Transportation and ash handling costs, frequency of unscheduled outages, fouling and slagging problems, and boiler tube erosion are all lowered by cleaning coal. The advantage of coal cleaning on these parameters has been shown, but the dollar amount varies with many assumptions and is difficult to pin down. The message to a utility, however, has not been lost and the trend toward cleaner steam coal is accelerating. This may lead to fewer installations that only wash the coarse coal, and add back the unwashed fine coal until a maximum contractual ash limit is reached. And, as the proportion of coal fines continues to increase by continuous and longwall mining, it may become impossible to clean the coal sufficiently without cleaning the fines. Flotation and oil agglomeration are the only processes that presently deal directly with cleaning the very finest coal fractions. Flotation has proven its effectiveness in cleaning blackwater and in separating fine coal from mineral matter. It has the achievable potential for far greater pyrite rejection than is now obtained in commercial preparation plants. While flotation will continue to be the predominant method for fine coal cleaning, there are conditions and opportunities where oil agglomeration can be a

Jan 8, 1980

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 a broad 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.

Jan 1, 1992

By B. K. Parekh, R. H. Mehta

Handling of fine coal is an importance issue for coal as well as the utility industry. Reconstitution in form of pellet or briquet would be desirable if it could be done economically. This paper evaluates the effectiveness of three binders e.g., asphalt- emulsion, corn starch and Brewex, in forming pellets of ultra-fine clean coal. It was found that asphalt emulsion and corn starch were not effective binders for ultra-fine clean coal, however, Brewex provided excellent quality of pellets, which exceeded all the minimum quasi requirements of coal pellets. INTRODUCTION Advanced coal preparation techniques (physical or chemical) are designed to process ultra-fine (minus 200 mesh or 74 pm) coal. The average particle size of the clean coal product is about 25 pm. Currently available dewatering techniques are effective in providing a 25 to 30 percent moisture product of this size coal slurry (Parekh et al. 1989). Handling and utilization of this dewatered ultra-fine size coal is difficult. Reconstitution, which includes pelletization, briquetting, extrusion, etc., of the dewatered coal product provides an ideal answer to handling and transportation problems. However, reconstitution of fines in the coal industry has not been successful due to various reasons. The most important criterion for its unsuccessful application in the coal industry is the cost of the binder (Capes et al. 1965). Various methods of reconstitution, such as briquetting, disk pelletization, extrusion pelletization and roller and die-pelletization, have been tested for fine coal agglomeration (Conkle et al., 1991) (Richards, 1990). Conkle et al., (1992) have provided a detail analysis of each technique in terms of moisture and operating cost requirements. According to Conkle, briquetting requires a product containing very low moisture, whereas pelletization can be done with coal containing up to 30 percent moisture. Currently available dewatering technique provides about 25 percent moisture product, thus, pelletization appears to be an ideal route for reconstitution. In this paper, results are presented on the comparison of three different binders i.e., asphalt-emulsion, corn starch and Brewers condensed solids (BCS or Brewex), for fine coal pelletization.

Jan 1, 1995

By V. N. Kelly

The ability to link geological sections on a graphics screen to form three dimensional geology representation was implemented at Kidd Creek in 1987. This technique affords an opportunity to optimize metal estimation in ore at the Kidd Creek copper, zinc and silver mine, by applying enhanced interpretation methods. On a series of geological sections varieties of ore are outlined. These are linked from section to section to form polyhedra which closely simulate the three dimensional shapes of the different ore types referred to as mintypes. The orebodies are divided into zones, referred to as domains, that approximate the function of variograrns, for which&apos; preferential continuity of mineralization has been estimated for distances in the width, strike and dip directions. These distances define ellipsoidal volumes of influence. The domains are also linked into polyhedra. Both types, coded by ore variety and zone, are mapped into the block model. Tonnage and metal values are calculated for ten-foot cubical cells of the block model by any of several estimation methods under domain control. In estimating the metal grades for a cell the predetermined search ellipse is applied to select only those drill hole assays of the same mintype as the cell designation. Once grade estimations have been determined for the cubical cells comprising the model, a wide range of stored information can be viewed. For example, in plan at any row elevation or on vertical section at any azimuth, as either printed values for the cells or as contoured data. This flexibility is of particular value in checking and refining the geological interpretation to better conform to ore types and grades actually mined. Combined with three dimensional geological modelling, and correctly chosen domains, the use of mintypes implants essential geological control into the system and prevents assays for one ore type from affecting grade estimation in adjacent ore types. The three dimensional geological modelling and block model segments provide information which is effectively applied to the stope design, production, planning and control parts of the system.

Jan 1, 1989

By M. Szatkowski

The studies conducted by Laplante and coworkers (1983,1984) with flotation of galena and quartz, showed that the rate of flotation reaches a maximum value for a certain intermediate value of the air flow rate. A similar, though less pronounced, effect of the air flow rate on the selectivity was also observed. Luttrell and coworkers (1985) showed that the size of air bubbles has a significant effect on both rate and selectivity of coal flotation. These two studies used flotation tests in which no froth layer was formed. This note describes effects of the air flow rate and bubble size on the rate constant and selectivity of coal flotation, using a mechanical laboratory machine with a froth layer allowed to form. Experimental A raw, dry ground Illinois coal was used in flotation tests. The size distribution and ash content of the samples are : Particle size (µm) Weight (%) Ash (%) +600 4.30 41.7 600/200 43.05 19.0 200/74 22.25 16.0 - 74 30.40 24.0 average 20.8 No. 2 fuel oil with a concentration of 200 g/t of feed and M.IBC with concentrations of 11, 22, and 33 mg/L of pulp (128, 256, 384 g/t, respectively) were used as flotation reagents. Tap water with a natural pH of 7.0 to 7.8 was used throughout the tests. No other reagents were used. The starting concentration of coal in the pulp was 86 g/L. A standard flotation machine with the pulp volume of 5 L was used in experiments. Three different air flow rates were employed, resulting in three air escape velocity values of 14, 30, and 70 cm/min. The air escape velocity (AEV) was calculated as the ratio of air flow rate (Qa) to the slurry surface (Ss) in the machine. For a comparison, the upper limit of air escape velocity values used in commercial size flotation machines is between 70 and 100 cm/min (WEMCO, 1983). A machine consisting of a generator of fine bubbles and a separatory column, described elsewhere (Szatkowski and Freyberger, 1985), was used for some tests where bubble size lower than that generated in the standard machines was required. Prior to a flotation test, a batch of coal was mixed with water in the flotation cell for at least 10 minutes to ensure that all coal was thoroughly wet. Then the specified amount of fuel oil was added and the pulp was conditioned for five minutes. The specified amount of MIBC was added 15 seconds before the end of conditioning. After air flow was opened, flotation products were collected for 0 to 30, 30 to 60, and 60 to 120 second intervals. Content of the combustible material in samples was determined by subtracting the ash content from 100%. A photographic technique was used to determine initial distributions of bubble size generated by the flotation machines. Bubbles generated in water containing the prescribed amount of the frother but no solids were photographed against a black background. The average diameter and standard deviation were then calculated for each set of flotation conditions. The method gives results reproducible within k 5% of the average bubble size. Table 1 contains information about the measured bubble size distributions expressed by the average bubble diameter (Db) and the standard deviation ( SD ) . Results and discussion The general formula for calculating the flotation selectivity is: o = recovery of coal/recovery of ash (1)

Jan 1, 1988

By Gaither H. Kitchell

Introduction Fire assaying is an ancient method of determining the amount of gold and silver in geological materials. Today&apos;s commercial, mining, school, and private fire assaying laboratories still use this basic assaying method. One of the primary reagents used to collect gold and silver from the ore is litharge (PbO), or lead oxide. The Federal Occupational Safety and Health Administration&apos;s (OSHA) and the Nevada Division of Occupational Safety and Health&apos;s (OSH) have a lead standard. It states that a person exposed to airborne lead at concentrations greater than 50 µg/m3 (1.4 µg per cu ft) of air over an eight-hour period has gone beyond the permissible limit for health safety. It is not uncommon to get airborne lead concentrations of 1400 µg/m3 (40 µg per cu ft) in the flux/sampling area or 900 µg/m3 (25 µg per cu ft) around the gas/ electric furnaces during the fusion and cupelling phases. Therefore, any individual working in these areas for more than 30 days per year without proper protection is a likely candidate for lead poisoning. Lead In A Human Lead enters the body as a toxic substance by breathing it into the lungs and upper respiratory tract and by getting lead particles into the digestive system through ingestion. Smoking, eating, drinking, chewing tobacco and gum, or wiping a lead-contaminated hand across the mouth are all potential ways forgetting lead into the body. Lead taken in large quantities can actually kill a person in a matter of days. This is rare, however. The real hazard of lead overexposure is for the unsuspecting person who works in an unsafe lead environment for months or years. This individual can slowly receive damage, sometimes permanent, to the central nervous system, gastrointestinal tract, kidneys, blood, gums, and reproductive system. Safe Blood-Lead Level The Lead Standard states that the maximum safe blood-lead level is 40 µg/100 g (40 µg per 3.5 oz) of whole blood. It recommends 30 µg/100 g (30 µg per 3.5 oz) if a person is of child-bearing age. People not working around lead products normally have a blood level of less than 10 µg/100 g (10 µg per 3.5 oz). A person not adequately protected can exceed 40 µg/100 g (40 µg per 3.5 oz) in a few weeks by fire assaying only 100 samples per day per week. Studies show that some individuals have developed lead-related diseases with blood-lead levels less than 80 µg/m3 (2.3 µg per cu ft). Lead Poisoning Symptoms There is a long list of symptoms associated with lead poisoning. The most common are: a prolonged, sweet, metallic taste in the mouth; loss of appetite; anxiety and irritability; weakness and dizziness; nausea and tremors; and insomnia. Preventive Measures Based on the lead standard, a number of preventive measures can be taken to ensure that fire assay laboratory personnel frequently exposed to lead are protected against overexposure. First, the actual lead concentration within the assaying area should be determined. In Nevada, one can contact the Department of Industrial Relations, Division of Occupational Safety and Health for assistance in monitoring any laboratory. The agency will also recommend ways to protect against lead. This agency does the monitoring and provides an educational class on lead safety. This type of agency can be found in other states. Commercial consulting firms are also available, but they are expensive. Regardless of the technique used to determine lead concentrations, it is most likely that the average fire assay area exceeds the 50 µg/m3 (1.4 µg per cu ft) lead concentration limit. This is considered unsafe.

Jan 8, 1985

By Stanley W. Ivosevic

Introduction Increasing investor and industrial demand for gold is not being matched by new mine output from traditional sources. This forces the exploitation of alternative natural and industrial resources to supplement traditional sources. A traditional natural resource is either high- to medium-grade ore, of gold or silver with coproduct gold. Or, they are medium- to low-grade base metal ores having byproduct gold. Traditional ores are also frequently extracted at relatively great expense by selective underground mining. Alternative natural resources include low-grade, near surface ores and the products of mining-dumps - and of onsite processing - tailings. Traditional industrial resources are newly mined gold or that obtained from plant sweepings. Their alternative analogues include gold in old scrap and chemical sludges and precipitates, notably those incidental to the electronic industry. These alternative gold resources have several attractions. They are untapped and abundantly available. Having been overlooked by prior metals suppliers, title to many alternative resources is easily acquired. Or, the value of such a resource may go unnoticed until its gold is rendered exploitable by an advance in extractive technology or an other approach. This article addresses the effect of large tonnage, low-grade lode ores on gold supply. Exploiting these ores is rendered commercial by their amenability to bulk mining by modern large-scale mining and metallurgical operations requiring little selectivity. Placer gravels, perhaps the earliest type of gold ore mined, also are bulk minable. But these fall outside the definition of being lode deposits. Most lode bulk mining is from surface open-pit mines. Some, however, is from underground by such large-scale mining techniques as room and pillar, vertical crater retreat, and end slicing. The low-grade gold ore being discussed averages 2.8 g/t (0.082 oz per st). Output How greatly do bulk minable, low-grade resources effect supply? 1982 was a somewhat healthy year for gold mining and exploration in North America in spite of the general depression in the mining industry (Table 1). The wildly fluctuating price of gold bullion had stabilized at an annual aver¬age of $12.08/g ($376 per oz) - (Handy and Harmon base price). Half of the 1.3 kt (43 million oz) of 1982 world gold mined were from South Africa&apos;s high-cost, medium-grade, selective underground mines. Of the remaining half, 20%, or about 130 t (4.2 million oz), was mined in North America. This includes Canada and the US - the third and fifth largest gold producers in the world. Of that North American production, more than 62.2 t (2 million oz) of gold, or about 50%, came from bulk mining of gold ores with or without co-product silver. Most of this was in the US, where bulk minable gold-silver lode ore produced nearly 31 t (1 million oz). This amounted to 60% of the nearly 46.6 t (1.5 million oz) produced in the US during 1982. An additional nearly 7.7 t (250,000 oz) of gold were produced as the byproduct of bulk mining of cop¬per ore, for a total of 34 t (1.1 million oz) of gold, or 75% of US gold production by bulk mining. To further illustrate this, US placer, mine dump, and related operations produced an insignificant 4% of US gold output at the time. Exploitation of bulk minable gold deposits is becoming increasingly important worldwide. Most new Australian gold mining announcements are of bulk minable developments. This trend will increase in North American mining as more large tonnage, low-grade operations come onstream in Canada, where much current production is from underground. It will increase with the general climb in Mexican gold mining. And it will grow with expansion of the 12 t/a (400,000 oz per year), gold production of the Dominican Republic. Production Metal price and operating costs make these large tonnage, low-

Jan 11, 1984

By J. Gallegos, R. Faber, T. Cox

"An innovative method to reduce acid rock drainage (ARD) and sediment loading of a recreational lake was used in the Red River watershed in northern New Mexico. The Red River drains a mining district that contains multiple naturally occurring hydrothermal alteration zones that release ARD and sediment to the river during episodic summer rain storms. A lake is near the mouth of the watershed and is filled with water from the river through a sluice gate and diversion. An automated actuator was installed on the sluice gate to close when the river is impacted by high-suspended sediment. Water quality probes were installed at the sluice gate to monitor the river water for turbidity, specific conductivity and water depth. If high turbidity is measured, the sluice gate closes automatically. Remote access is setup with a computer view screen to review the real-time system status. Once the turbidity drops below a threshold, the sluice gate is reopened remotely. The automated sluice gate has successfully reduced the amount of ARD and suspended sediment entering the lake. This reduction has resulted in an improved aquatic habitat for the fishery and benthic macroinvertebrate community. BACKGROUND The Red River is located within the Sangre De Cristo Mountain Range in northern New Mexico. The river drains an area of approximately 190 square miles and spans elevations from 6,600 feet at its confluence with the Rio Grande to its headwaters near 12,000 feet (Figure 1). The river generally flows to the west along most if its length.The Red River watershed includes several hydrothermal alteration zones. The alteration zones were created by intrusion of high-silica granites and subvolcanic porphyries, with associated hydrothermal fluids, into the volcanic rocks of the Questa Caldera. The hydrothermal alteration is also responsible for the molybdenum deposits in the area. Rapid chemical weathering and physical erosion of pyrite-rich hydrothermally altered rocks has resulted in the formation of alteration scars. The alteration scars are characterized by steep, erodible slopes with little vegetation and incised gullies. Debris fans from the drainages containing active alteration scars have formed and typically consist of poorly sorted sand and gravel that is distinctly pale yellow in color. Some of the more prominent scars are located in the Hansen, Straight and Hottentot Creek drainages downstream of the town of Red River."

Jan 1, 2018

By Ernest H. Gilmour

Introduction An earth science teacher at the junior high school level has the potential of working with 450 students per year. And these young people are at an age where they can be properly informed about the importance and necessity of a viable, domestic minerals industry. It was with this philosophy that a course was developed at Eastern Washington University for science teachers entitled, "The Minerals Industry of the Pacific Northwest." The two-week course is sponsored by the Northwest Mining Association and the Idaho Mining Association. It was first taught during the summer of 1971 and will be taught during the summer of 1986 for the sixteenth straight year. It is a two-week mini-version of the very successful six-week course taught at the Colorado School of Mines. In 1971, Russ Chadwick was President of the Northwest Mining Association. He suggested that such a course should be developed and offered at Eastern. It was determined that tuition scholarships would be necessary. This would ensure that a reasonable number of teachers would register for the course. Say a teacher has five classes a day, with 30 students in each class. Some 150 students come into contact with that teacher each term. With three terms a year, an instructor works with about 450 students per year. So, with 20 teachers in the course, it was possible that 9000 students a year could learn from our informed teachers. And now, after 15 years, more than one million students have learned about the minerals industry from teachers who have completed the course. The course is designed to take the teachers through all aspects of the minerals industry. Most of the lectures are given by the working geologists and engineers who talk about their jobs in the industry. This is extremely important for the credibility of the course. The teachers realize that the lecturing professionals really believe in what they profess. Industry fundamentals The course begins with a lecture on the need for minerals in a modern society. Most of the teachers do not realize how critical minerals are to every aspect of our daily lives. The lecture is followed by the US Bureau of Mines movie, "Minerals Challenge." An exploration geologist then discusses what is involved in the exploration for a mineral deposit. This includes a case history that the geologist has actually been involved in

Jan 9, 1985

By J. E. Gebhardt, H. Toro, K. Y. Lee

Industrial-scale test results indicate superior metallurgical performance of column flotation compared to conventional cell flotation for copper cleaning. A technical comparison of bubble generation systems for column flotation is made. Two types of bubble generation systems are considered: the filter-cloth type spargers and the Turbo-Air® sparger, which utilizes an air-water mixture fed to the column through injection tubes. Examples of the improved technical performance available with an external sparger system are presented using data from plant tests conducted for specific applications with several different mineral ores. INTRODUCTION Column flotation is now a valuable unit operation which mineral processing engineers employ in the design and operation of flotation circuits in nearly all mineral applications. The acceptance of columns in the mid- 1980s by the minerals industry was paralleled by a rigorous approach to developing a scientific understanding of the design, scale- up, and operation of column flotation units (Finch and Dobby, 1990). Since then, column flotation has become a standard unit of the cleaning stage in many new mineral concentrator projects and has replaced older mechanical cells to improve process efficiency. The size of column cells has continued to grow from the initial 0.9- to 2.1 -meter diameter range, and it is not uncommon to find column cells today being 2.7- to 3.6-meters in diameter. With the success of the smaller units, the scale-up to larger cells has been proven and demonstrated by the performance of the large-diameter cells (Murdock and Wyslouzil. 1991). The use of columns in flotation circuits has also been accompanied by developments in modelling. control, and optimization of column performance (Lee et al, 1991). The bubble generation system is the heart of the operation of column flotation cells. The most common methods for generating bubbles can be divided into two types: the internal and external spargers. Internal spargers are typically a porous material, such as sintered glass, porous rubber or filter-cloth, or maybe a single- or multi-nozzle sparger. External spargers implies that a gas and a liquid (either water or slurry) are brought into contact outside the column and that mixture is then introduced into the column. Several types of external spargers are available today, and innovative designs for externally mounted air sparging systems have been a significant factor in the successful application of columns. Early types of bubble generation systems.

Jan 1, 1993

By C. J. Johnson, C. J. Bise

Introduction Coal extraction by continuous miners (CM) is currently the most common underground method in the US industry and accounts for slightly more than two-thirds of the nation&apos;s deep mining production (National Coal Association, 1987). Even if longwall mining should become more commonplace, it can proceed only after ventilation and access entries have been driven by CMs. Since an area of concern continues to be the effects of the dust generated on the health of mineworkers, this paper discusses the relationship between the elemental compositon of mining-generated airborne dust sampled from the immediate ventilation return of a CM and laboratory-generated dust derived from channel samples taken from the mines. There are several potential contributions of this type of study to the coal mining industry. First, after more fundamental knowledge of the causes of Coal Worker&apos;s Pneumoconiosis (CWP) is learned, the laboratory-generated respirable dust could be used to identify a potentially hazardous coal seam. Also, this study could possibly aid in understanding the fundamental causes of CWP by producing mining-simulated samples of coal dust that could be used in epidemiological studies. Further, assuming that there is no difference in the elemental composition of a drill-core sample and a channel sample from the same location, a mining company could predict a new mine&apos;s respirable dust elemental composition in the immediate ventilation return by using exploratory drill core samples of the roof, coal, and floor rock to prepare the laboratory dust. Ventilation engineers could then use engineering design and control measures during premine planning to reduce the incidence and severity of CWP by better ventilating the potentially hazardous coal seam. If this proper planning prevented any future changes to the ventilation equipment and mine design, much time and money could be saved. Scope of work To investigate the variability of the chemical characteris tics of respirable dust, airborne dust samples from eight underground coal mines located in the eastern and midwestern United States were collected with eight-stage Sierra Model 298 Marple cascade impactors, as well as 25 channel samples of mined material. Each channel sample was removed from the middle of the coal face before mining occurred. Sampling of the mining-generated dust was conducted by Lee (1986) by sampling the entire working sections, primarily for characterization purposes, to obtain information on the locational variability of dust characteristics. Research performed for this study used the elemental analyses of the mining-generated dusts he sampled in the immediate ventilation returns of CMs. The procedure that was used to produce the laboratory-generated respirable dust was based on the Hardgrove grindability test since it reflects the pulverizing characteristics of coal. This test was chosen for several additional reasons. First, it is repeatable and reproducible. A consistent amount of input energy is used as well as a specified size range of feed material to be crushed (the channel samples). Second, it is thought to generate secondary dust in a way similar to that of the crushing and grinding of the coal and rock as they pass through the arc-shaped cutting path of the CM&apos;s cutter head. The potential effect on dust generation by this secondary grinding mechanism may be at least as much as that produced by primary fragmentation, which is dust produced by the cutting action of the bit against the coal or rock (Roepke, 1984). Finally, the Hardgrove grindability test is well known and is used in the coal industry to guide mineral processing engineers in estimating the capacity of mills used to grind coal. One hypothesis of dust researchers in the Generic Technology Center for Respirable Dust is that the elemental as well as the physical characteristics of coal mine dust will make a difference in the incidence and severity of CWP. Coal mine dust is generated not only from coal, but also from any rock partings contained within the seam or any roof or floor material mined with the coal. Thus, coal mine dust may not have the same elemental characteristics as the coal being mined. Given that hypothesis, mixtures proportional to each thickness mined of roof, coal, and floor rock derived from the channel samples of the face areas from which the respirable dusts were generated by the CMs were used to produce the feed material that was pulverized in the Hardgrove machine.

Jan 1, 1989

By C. J. Johnson, C. J. Bise

Coal extraction by continuous miners (CM) is currently the most common underground method in the US industry and accounts for slightly more than two-thirds of the nation&apos;s deep mining production (National Coal Association, 1987). Even if longwall mining should become more commonplace, it can proceed only after ventilation and access entries have been driven by CMs. Since an area of concern continues to be the effects of the dust generated on the health of mineworkers, this paper discusses the relationship between the elemental compositon of mining-generated airborne dust sampled from the immediate ventilation return of a CM and laboratory-generated dust derived from channel samples taken from the mines. There are several potential contributions of this type of study to the coal mining industry. First, after more fundamental knowledge of the causes of Coal Worker&apos;s Pneumoconiosis (CWP) is learned, the laboratory-generated respirable dust could be used to identify a potentially hazardous coal seam. Also, this study could possibly aid in understanding the fundamental causes of CWP by producing mining-simulated samples of coal dust that could be used in epidemiological studies. Further. assuming that there is no difference in the elemental composition of a drill-core sample and a channel sample from the same location, a mining company could predict anew mine&apos;s respirable dust elemental composition in the immediate ventilation return by using exploratory drill core samples of the roof, coal, and floor rock to prepare the laboratory dust. Ventilation engineers could then use engineering design and control measures during premine planning to reduce the incidence and severity of CWP by better ventilating the potentially hazardous coal seam. If this proper planning prevented any future changes to the ventilation equipment and mine design. much time and money could be saved. Scope of work To investigate the variability of the chemical characteris tics of respirable dust, airborne dust samples from eight underground coal mines located in the eastern and midwestern United States were collected with eight-stage Sierra Model 298 Marple cascade impactors, as well as 25 channel samples of mined material. Each channel sample was removed from the middle of the coal face before mining occurred. Sampling of the mining-generated dust was conducted by Lee (1986) by sampling the entire working sections, primarily for characterization purposes, to obtain information on the locational variability of dust characteristics. Research performed for this study used the elemental analyses of the mining-generated dusts he sampled in the immediate ventilation returns of CMs. The procedure that was used to produce the laboratory generated respirable dust was based on the Hardgrove grindability test since it reflects the pulverizing characteristics of coal. This test was chosen for several additional reasons. First, it is repeatable and reproducible. A consistent amount of input energy is used as well as a specified size range of feed material to be crushed (the channel samples). Second, it is thought to generate secondary dust in a way similar to that of the crushing and grinding of the coal and rock as they pass through the arc-shaped cutting path of the CM&apos;s cutter head. The potential effect on dust generation by this secondary grinding mechanism may be at least as much as that produced by primary fragmentation, which is dust produced by the cutting action of the bit against the coal or rock (Roepke. 1984). Finally, the Hardgrove grindability test is well known and is used in the coal industry to guide mineral processing engineers in estimating the capacity of mills used to grind coal. One hypothesis of dust researchers in the Generic Technology Center for Respirable Dust is that the elemental as well as the physical characteristics of coal mine dust will make a difference in the incidence and severity of CWP. Coal mine dust is generated not only from coal, but also from any rock partings contained within the seam or any roof or floor material mined with the coal. Thus, coal mine dust may not have the same elemental characteristics as the coal being mined. Given that hypothesis, mixtures proportional to each thick¬ness mined of roof, coal, and floor rock derived from the channel samples of the face areas from which the respirable dusts were generated by the CMs were used to produce the feed material that was pulverized in the Hardgrove machine.

Jan 1, 1990

By K. R. Weller

In recent years, a variety of new or substantially modified grinding machines have been offered to the mining industry with claims of improved performance relative to standard grinding machines especially rod and ball mills. The incentive for developing such new machines is high, because costs associated with grinding are a very substantial part of the total costs of processing run-of-mine ore, and gold ores are no exception in this regard. The incentive is also high because of the very low conversion of input energy to comminution energy in most tumbling mills. If the most energy-efficient means of mechanical size reduction is slow compression of single particles, then this energy amounts to only about 3% of that used in tumbling mills. It would seem, therefore, that there is a huge potential to do better. This paper examines three types of grinding machine for which claims of higher energy utilisation ( amongst other beneficial claims) have been made, namely the high pressure grinding rolls, the vibrating drum mill and various styles of stirred mill. It presents evidence concerning the claims, examines the range of particle size and particle size reduction over which benefits might reasonably be expected to accrue, discusses the major issues in contention and briefly describes the research the CSIRO Division of Minerals has in progress, both independently and collaboratively, the latter mainly with the Julius Kruttschnitt Mineral Research Centre (JKMRC) via the Cooperative Research Centre for Mining Technology and Equipment (CMTE) and the Australian Mineral Industries Research Association (AMIRA). There are many criteria which may be chosen to assess relative performance of grinding machines. Those used below are:

Jan 1, 1995