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By J. Wagner, P. Eger, G. Melchert

In 1991, LTV Steel Mining decided to close an open-pit taconite mine in northeastern Minnesota using a passive-treatment approach consisting of limiting infiltration into the stockpiles and wetland treatment to remove metals. More than 50 Aft (55 million st) of sulfide-containing waste had been stockpiled adjacent to the mine during its 30 years of operation. Drainage from the stockpiles contained elevated levels of copper, nickel, cobalt and zinc. Nickel is the major trace metal in the drainages. Before the closure, the annual median concentrations ranged from 1.5 to 50 mg/L. Copper, cobalt and zinc are also present but they are generally less than 5% of the nickel values. Median pH levels range from 5 to 7.5, but most of the stockpile drainages have pH levels greater than 6.5. Based on the chemical composition of each stockpile, a cover material was selected. The higher the potential that a stockpile had to produce acid drainage, the lower the permeability of the capping material required Covers ranged from overburden soil removed at the mine to a flexible plastic liner. Predictions of the reduction in infiltration ranged from 40% for the native soil to more than 90% for the plastic liner. Five constructed wetlands have been installed since 1992. They have removed 60% to 90% of the nickel in the drainages. Total capital costs for all the infiltration reduction and wetlands exceeded $6.5 million, but maintenance costs are less than 1 % of those for an active treatment plant. Because mine-drainage problems can continue for more than 100 years, the lower annual operating costs should pay for the construction of the wetland-treatment systems within seven years.

Jan 1, 2000

By Anthony R. Przybylek, John M. Visnesky

INTRODUCTION Thermal dryers have been utilized for reducing moisture of bituminous coal since the late 1930's. Since that time, numerous changes in the coal industry have motivated advances in thermal drying technology. The purpose of this paper is to discuss the changes that occurred to dryer design and focus on current practices of fine coal drying. The major topics discussed in this paper are: - Early Thermal Dryer Design and Evolution - Current Drying System Design and Associated Components - Current Economics of Fine Coal Thermal Drying - Future Markets and Design Considerations EARLY THERMAL DRYER DESIGN AND EVOLUTION Background The practice of drying fine coal with heated air has been performed for more than one half century. Throughout that period, both Metallurgical and Steam coal producers have used this method for reducing the surf ace moisture content of their fine plant products. During the late 1930's and early 19401s, vibrating bed and flash dryers were used to per- form the drying duties. Closed water circuits and vacuum filtration had not yet been integrated into process circuitry and centrifugal dryers were in their early stages of development. During that same period, changes in mining practices were reshaping the distribution of coal to our processing facilities. Plants originally designed for the purpose of processing coarser fractions of coal were being burdened with additional quantities of fine coal. Those plants, unable to recover fine coal, would reject these solids to plant bleed. The higher incidences of stream pollution that resulted from these events prompted clean stream legislation and the need for new, more efficient fine coal recovery circuits. The coal processing industry responded to these needs by integrating closed water circuits and filtration into circuits. The fine coal filter cakes recovered exhibited poor handling characteristics by themselves and often required further dewatering to improve their characteristics. Flash dryers at that time were being used to dry the coarser products. For those facilities operating flash dryers, the solution to this new problem appeared to be simple. The filter cake was combined with the coarser (-1/4") coal and then introduced to the flash drying column for drying. The filter cakes, however, presented significant problems for the flash drying systems. Heavy lumps of filter cake could not be dispersed and conveyed by the rising hot air and as a result would settle out in the bottom of the column. This material, known as "drop-out", would then accumulate and require further handling. The solution to this problem was not an easy one. Sophisticated mixing equipment was employed to condition the feed to the dryers. As attempts to minimize drop-out with this approach failed, the only recourse was to collect the material with conveying systems and recycle it back to the plant or dispose of it. The additional costs of handling this material reduced the attractiveness of the flash systems and opened the door for the development of a new, more cost effective coal drying system. Due to the nature of the problems experienced by the flash systems, fluid bed systems were considered. Fluid bed dryers were being used in other industries at that time but had no operating history in the coal industry. Dorr-Oliver transferred technology from other industries and built the first fluid bed drying system for fine coal.

Jan 1, 1988

By William T. Folwell

INTRODUCTION - CURRENT RAISE BORING PRACTICES Raise boring is a comparatively new tool to the underground mining and construction industries. Underground raise boring, in this country, had its beginning in the early 1960's, in the iron mines of Michigan. These first raises were bored 40 in. and 48 in. in diameter with maximum lengths of approximately 200 ft. Since the late "sixties", the raise boring machine has been accepted by the mining industry as an important machine to increase excavation rates, provide adequate ventilation airways, improve safety, and reduce costs, to name a few of the benefits. The construction industry is now accepting raise boring techniques for hydro-electric power projects, water supply projects, tunnel projects, waste water projects, etc. Raise drills are now capable of boring raises to diameters as much as 12 ft, in the softer rock types. Raise lengths range from 200 ft to as long as 1700 ft. Bored raises, in general, are bored at dip angles ranging from 45º to 90º, depending on the requirement of the raise. The general practice is to drill the pilot hole down to an opening below and up ream the pilot hole to the required diameter. The pilot hole (9-7/8", 9", 11", or 12-1/4" diameter) is generally drilled with compressed air to remove the cuttings. In the event that ground water is encountered and removal of cuttings is impeded, a foaming agent is added to assist in lifting the cuttings. When excessive water is encountered, water is used as the circulation medium and the cuttings are removed by pumping. During the reaming cycle, the cuttings fall by gravity and are removed from the bottom of the raise by conventional methods.

Jan 1, 1997

By Morris Loshinsky, George Ziegler

NEW ROUTES - 63rd STREET CROSSTOWN The New Route construction along East 63rd Street in Manhattan began in 1969 with the East River tunnel connecting Manhattan and Queens. The 63rd Street Line, once in Queens, is planned to connect to another new route called the Queens By Pass or Super Express (See Exhibit 1) which would connect to the Queens Boulevard Line at Forest Hills. The purpose of the Queens By Pass was to alleviate subway congestion to and from the Borough of Queens, whose population had rapidly expanded during and after World War II. The existing Queens Boulevard subway lines (the "E" and the "F" trains) are almost filled to capacity at their terminal stations with passengers from Long Island. Passengers boarding trains enroute from the local Queens stations have difficulty boarding crowded trains, and, at times, cannot board the trains. The New Routes, thus, were to give additional train capacity into Manhattan and thereby relieve the crowded conditions along the Queens Boulevard corridor. PROFILE - METHOD OF CONSTRUCTION The portion of the New Route passing east-west under East 63rd Street on Manhattan (See Exhibit 2) had to snake beneath intersecting structures running north and south, such as the Con Rail tracks below Park Avenue and the 80 ft deep Lexington Avenue Express track structure along Lexington Avenue. Further, the Lexington Avenue Local tracks are in a separate structure which has a high profile close to the street surface. Therefore the 63rd

Jan 1, 1981

By J. A. Brierley, S. R. Hutchins, C. L. Brierley

Introduction Many gold and silver ores are refractory to conventional processing technologies. Thus, precious metals recovery is uneconomic. These resources can consist of ores, flotation concentrates, mill tailings, and other reserves. The refractory nature of these materials can be ascribed to such components as iron sulfide minerals, silica minerals, tellurides, organic compounds, and sulfosalts. Precious metals locked within iron sulfide minerals account for a large number of refractory gold ores. These minerals consist principally of pyrite or arsenopyrite. They must be oxidized to liberate the encapsulated precious metal and allow its contact with the leaching agent. Carbonaceous gold ores represent a unique class of refractory precious metal ores in general. Not only does gold encapsulation by iron sulfide minerals occur (Hausen, 1985), but these ores also contain organic matter that interferes with gold recovery by cyanidation. Direct interference can be attributed to either occlusion of gold within organic matter or formation of a stable gold-carbon complex. The magnitude of this problem is debatable. Wells and Mullens (1973) examined two carbonaceous gold ores and found no association between the gold and carbonaceous material. A more common problem is indirect interference, whereby the aurocyanide complex formed during cyanidation is sorbed by the native organic matter and is not available for recovery. This phenomenon is referred to as "preg robbing" (Osseo-Asare et al., 1984; Hausen and Bucknam, 1984). Guay (1981) summarized various methods that can reduce preg robbing during processing of carbonaceous gold ores. However, some are not applicable to carbonaceous ores that have significant amounts of gold locked in iron sulfide minerals. Several methods exist for oxidizing iron sulfide minerals and rendering refractory precious metal ore amenable to conventional recovery processes. These include various embodiments of roasting (Jha and Kramer, 1984; Brown, 1984; Arriagada and Os-seo-Asare, 1984; Smith, 1986), pressure oxidation (Muir et al., 1984; Beattie et al., 1985; Thompson, 1986; Weir et al., 1986; Argall, 1986), and chemical oxidation (Scheiner, 1971; Jackson, 1983; Van Weert et al., 1986). Each process has been successfully used for refractory precious metal ores. In many instances, however, their use is precluded by capital and operating costs relative to the value of the ore. An alternative to these methods is microbial pretreatment - a biological process whereby iron sulfide minerals are degraded and the liberated precious metal values can be recovered by conventional technologies. Bioleaching refractory precious metal ores is a relatively new concept compared to roasting and chemical oxidation. It is rapidly becoming established as a viable pretreatment alternative. Numerous studies have been conducted in the laboratory and on a pilot scale (Yudina et al., 1979; Lawrence and Bruynesteyn, 1983; Pol 'kin et al., 1985; Marchant, 1985; Livesey-Goldblatt, 1985; Gibbs et al., 1986). Comparative tests have demonstrated that bioleaching can be equal to or better than roasting or pressure oxidation in terms of product recovery and process economics (Fridman and Savari, 1983; Livesy-Goldblatt et al., 1983; Karavaiko et al., 1985; Gilbert et al., 1986; Hackl et al., 1986). Bioleaching's main drawback is that pretreatment requires days rather than hours. This can lead to excessive operating costs for difficult-to-treat ores. In all of the tests reported thus far, the leaching microorganism used has been thiobacillus ferro-

Jan 4, 1988

By P. T. Perdue, F. F. Haywood

INTRODUCTION Monitoring devices used for detecting and measuring airborne radon in mines are limited due to the rugged environment in which they must operate. Personnel monitors now available are either passive or require filters and an air pump. Area air monitors usually have bulky batteries or require electric power service to the instrument. Thus, there are few instruments presently in use in mines capable of discriminating between the various radon isotopes and their daughters. Occasionally it is desirable to determine which radon daughters are the principal dose contributors in mine air. In some areas, such as the Elliot Lake mine in Canada, there exist sizable quantities of 220Rn in the mine air due to 232Th in the ore (Cote and Townsend, 1981). Since there has been recent interest in radon daughters in homes and factories, as well as caves and mines, methods and instrumentation are being developed to address these cases. Some of these methods can be employed in uranium mines. Although there have been very few determinations of the three radon isotopes (222Rn, 220Rn, and 219Rn) and their daughters made in uranium mines, these isotopes have been measured in process buildings and in private homes under circumstances that suggest these isotopes could also be present in uranium mine air (Perdue, Leggett, and Haywood, 1980; Yarborough, 1980) METHODS Cryogenic air sampling was developed about 15 years ago (Rasmussen, 1972) to concentrate various airborne chemicals for later fractional distillation and measurement of various organics using gas chromatography. A version of this technique has been developed to measure tritium, krypton, and other radioactive gases in nuclear power station buildings (Anon, 1980). In this method, fractional distillation is used in conjunction with various detectors to identify the radionuclides present. The cryogenic method to be discussed here was developed at Oak Ridge National Laboratory (ORNL). It uses the liquid air accumulated in a sampling device for a direct analysis of all gamma-ray emitters present. The air sampler (Fig. 1) could be employed in a mine to liquefy air, using the difference between the point of liquefaction of air (~80 deg. Kelvin) and the boiling point of liquid nitrogen (77 deg. Kelvin) to liquefy an air sample. By using a proper orifice at the entry to the sampling vessel, air of any dust concentration and relative humidity [ ] may be sampled. The energy for sampling is provided by the slow evaporation of the liquid nitrogen; no other power source is required. This type of sampler could be used to augment the usual filter and pump, or passive system presently employed to estimate personnel dose from radon daughters. It would also allow measurement of any combination of daughters from the three radon isotopes. Upon completion of sampling, the sampler is closed and removed from the mine. This sampler, with its liquid air contents maintained by the slow evaporation of liquid nitrogen, is taken to the place of measurement. There, the sampler is opened, the accumulation of water and ice in the neck is chiseled down into the liquid air, and the sample is transferred quickly to a precooled counting dewar (Fig. 2). This counting dewar is then placed upon a gamma radiation detector; we recommend a high purity germanium detector. A determination of all the radon daughters can be made in 1 to 4 hours after the sample is taken.

Jan 1, 1981

By James M. Williams, Michael E. Zellars, Arthur J. Roth

INTRODUCTION Our present estimates of world P205 consumption for all uses show an average annual growth rate of approximately 1.9% per year. The estimated 1985 consumption of 41.5 million tonnes of P205 is predicted to increase to 57.1 million tonnes of P205 by the year 2005, which represents a growth from 157.7 million tonnes in 1985 to 217 million tonnes of phosphate rock in 2005. The present situation of oversupply and extremely weak phosphate product prices is, however, expected to continue in the short term and a normal supply/demand balance may not be achieved until after 1990. The growth in P205 demand, in conjunction with the higher cost of replacing low cost mines as they are depleted, will place economic pressure on world farm economies through the increased cost of phosphate fertilizers. The availability of rock supply is not the challenge; rather, it is the economics of new rock supply. This economic challenge of providing new rock supply is by no means limited to the United States. Even in centrally controlled economies, the cost of rock from domestic mines of marginal quality and economic viability, coupled with burdensome infra¬structure requirements, may result in tonnage additions at very high cost. Although government subsidies to offset higher production cost is a factor in some countries, world production in the next 20 years will be directed, if not governed, by economic factors. Established infrastructure, good quality reserves, and existing operations will provide a competitive advantage for new prod¬uction. It is our opinion that the current world supply structure will remain relatively stable through 2005, in that most of the traditional suppliers will retain supply roles. U.S. rock production, although it is not expected to grow beyond the present level, will still be a major contributor to world supply. Although dramatic changes within mining, fertilizer production and farm economics may appear disruptive, the current supply scenario will, in general, continue with the major issue being production economics. WORLD DEMAND Demand for P205 by region in the world has been estimated using projections of population, crops and yield, arable land, economic and other data as a basis for future demand projec¬tions by Blue-Johnson and Associates, our col¬laborators on a recent multi-client study. Non-fertilizer P205 demands were estimated in the same general manner. These estimates of future demand are based on an assumption of reasonable world political and economic stabil¬ity, debt repayment requirements maintained at workable levels, and resolution of problems in the agricultural economy to a point where normal supply/demand forces can interact freely to provide a basis for continued P205 supply. The results of these studies indicated an average growth rate of 1.9% per year through the year 2005. This growth could be as low as 1.5% per year, or as high as 2.5% per year. It is our opinion that the estimated rate of 1.9% per year is most probable. This moderate growth rate considers both the major economic constraints facing the world agricultural industry today and the general problem of affordability of fertilizers and other crop inputs. WORLD SUPPLY There are adequately defined world reserves of phosphate rock today to provide supply well into the future, even at much higher levels of demand than presently projected. Using the U.S. Bureau of Mines estimates, which are based on stringent criteria for definition of reserves, the present world reserves base is approximately 35 billion tonnes. Identified resources raise the potential for production by orders of magnitude. Potential resources based on geological trends, and including offshore deposits, would again result in orders of magnitude increases in rock resources. Morocco leads the world in reserves base with approxi¬mately 21 billion tonnes. The U.S. is second, followed by the Republic of South Africa and Russia. These four countries control about 87%

Jan 1, 1986

By Mark Board

INTRODUCTION As the world's supply of near-surface metalliferous ore bodies is depleted, more emphasis must be placed upon mechanized bulk systems of underground mining. In particular, block caving is becoming increasingly important as a method of mining low to moderate grade deposits at depths of somewhat less than 1219 m (4000 ft). 00 ft). The block caving method requires that the ore body be completely undercut, thereby removing its means of support and inducing failure which progresses from the undercut to the ground surface. As the span of the under- cut increases and the weight of the overlying crushed ore body becomes greater, large stresses are induced on slusher and haulage drifts. Stability of these drifts is critical in block caving operations and often proves to be an extremely difficult task (Kendorski, 1976; McWilliams, 1975). Slusher drift maintenance can be costly, often exceeding that of the original support costs. This chapter analyzes the ground support problems of an Arizona block caving mine. The mine presently uses monolithic concrete liners as a support method. Failure of the liners occurs as the undercut approaches, thus initiating a maintenance problem which will continue until the ore body is depleted in that block or panel. To examine this problem, a simple analysis of the stresses induced by the advancing undercut was per- formed. Utilizing a design method developed by Dixon ( 1973, 1974), the calculated stresses were used to deter- mine the minimum thickness required for varying section shapes of unreinforced concrete liners. It is found that the thicknesses required for stable linings are not economic. Recently developed empirical methods of support de- sign are utilized (Barton, 1974; Bieniawski, 1973) and compared to the analytical method. A final discussion of alternative methods of support is given with recommendations for the use of passive or "yielding" support methods. LAKESHORE MINE (GENERAL GEOLOGY) The Lakeshore copper property is located in Pinal County, AZ, approximately 45 km (28 miles) south of Casa Grande. The Lakeshore deposit is a porphyry cop- per similar to others in Arizona. The ore body consists of disseminated chalcopyrite and fracture fillings in quartz monzonite porphyry (QMP), Cretaceous volcanic (KVS), Precambrian diabase, and Mescal limestone. The ore body has been divided into three distinct zones of mineralization which have been termed (South, 1972) the oxide zone, the sulfide zone, and the tactite zone. The oxide zone occurs in the upper portions of the ore body and consists primarily of leached and oxidized quartz monzonite porphyry and Cretaceous sedimentary and volcanic rocks. The sulfide zone underlies the oxide zone and includes all quartz monzonite porphyry and Cretaceous rocks which have not been oxidized. The thickness of the sulfide zone is variable and ranges from about 61 to 122 m (200 to 400 ft). The tactite zone occurs in a series of Paleozoic metasediments consisting of tactite, quartzite, diabase, hornfels, and dikes of quartz monzonite porphyry. The tactite zone is highly variable in thickness and underlies the Cretaceous volcanic and sediments throughout the mine. Together, the tactite and sulfide zones are known as the thick or tall sulfides. The greatest support stability problems occur in mining of the tactite zone and are therefore the subject of the remainder of this chapter. MINING METHOD The Lakeshore mine is a block caving mine utilizing a panel caving method. The present mine consists of two

Jan 1, 1982

By Cyril W. Draffin

Introduction Americans have long been proud of their ability to recognize a need, define it, and then develop workable solutions. The U.S. underground coal conversion (UCC) program is a good example. In 1974, the need to utilize coal resources that could not be mined was recognized, and Congress gave the Bureau of Mines $1.7 million to gain a better understanding of the UCC technology and its potential applications. Since 1977 when the U.S. Dept. of Energy (DOE) was formed, workable technologies have been tested, the market for underground coal conversion has been recognized, and extensive industrial interest has developed. Perspective The U.S. is blessed with extensive coal resources. Coal supplies about 19% of our national energy needs, yet recently promulgated environmental standards and increasing mining costs have limited direct coal use. UCC potentially can use coal that is economically or technically infeasible to mine because the coal is too thick, deep, dirty, high-ash, and wet; has an excessive angle of dip; or has unpredictable and poor overburden characteristics that make mining unsafe. Coal can be gasified underground or perhaps liquefied below the earth’s surface. Underground coal gasification (UCG) is the prime element of the underground coal conversion program. It enables coal to be converted to gas underground. UCC Coal Resource Recent data show a total U.S. coal resource base of about 6.4 trillion tons. Of this, 1.8 trillion tons might be accessed by UCC, including some minable coal. A commercial UCC process could quadruple roughly the 450 billion tons of U.S. coal reserves, which presently are recoverable by mining. The U.S. coal resources are divided into five sectors. The two most eastern sectors contain relatively thin, swelling bituminous coals that are difficult to gasify. Of the two, the bituminous coal in the Eastern Province is the most difficult because it is deeper than , in the Interior Province and the terrain is more rugged. The sector along the Rocky Mountains contains huge amounts of thick low-rank coals that are the most promising for early development by UCC. The coal has good continuity and the terrain is suitable. Low-rank coals such as subbituminous and lignite are highly reactive and tend to shrink and disaggregate when heated. Most of the coal in the Pacific

Jan 1, 1980

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

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

Jan 1, 1990

By James R. Boulton

Fire detection using Carbon Monoxide sensors is based on the assumption that a rise in the level of Carbon Monoxide in the atmosphere over a fixed, predetermined ambient level is indicative of a fire. The method is simple, reliable, and effective in those mines where the only source of CO is from undesired combustion. Ambience for this purpose is defined as that level of the gas that is present on a day to day basis whose source is not a fire. In a diesel powered mine the level of CO can and does vary so widely from hour to hour and minute to minute that the selection of a predetermined, fixed level represents a compromise between sensitivity and reliability. The compromise is a best of both and worst of both situations and the computer system often resembles the little boy who cried wolf instead of a state of the art early fire monitoring and detection system. Cyprus Coal personnel have developed method of determining the ambient level of CO in the atmosphere on a real time basis and using that dynamic ambient level to determine whether or not a fire exists in the mine. The method has been given the name of Wave Trend Crossing.

Jan 1, 1991

By J. M. Dean, R. L. Grayson

Introduction With the continued downward pressure on coal selling prices, mine management must make its operations as efficient as possible. The problem is the same one throughout the mineral industries, streamlining operations to maintain competitive min¬ing costs (Born, 1991). There has been a great deal of research into better using production manpower. But the story is different for operations support personnel. Traditionally, planning man¬power requirements are done annually, with modifications to the operating plan occurring generally quarterly. On the other hand, the demands on manpower are often dynamic, being responsive to changing productivity levels, mine conditions (including water influx, bad roof, poor floor) and unanticipated project requirements (responses to machine breakdowns or a new section setup requirement because of a drop in coal quality in another area). Such changes may impact the manpower structure significantly. At times they may cause serious labor relations problems concurrently. Realignments and the generation of instability in the workforce because of active job switching (bidding) are a result. Most responses by coal operators are reactive rather than proactive, a situation that could be mollified by more accu¬rate and detailed planning of predictable manpower require¬ments. A manpower planning tool is needed that takes into consideration such things as changing productivity levels and conditions as manhour requirements for different work areas are planned, rather than blindly allocated through traditional staffing modes. Mine Management Support System The Mine Management Support System (MMSS) is an integrated decision support system that was developed using Sun Microsystems hardware according to the contract be¬tween the US Bureau of Mines and West Virginia University. The software used in the development of MMSS included X¬ windows (Scheitler and Gettys, 1986) for the user interfaces, LASER (Raman, 1985; Hayhurst, 1990) for the expert sys¬tem and graphics components, and the C language for general programming requirements. Sun DOS was used to allow the use of IBM-compatible applications software on the Sun networks. Sun OS is the operating system. It can be viewed as an enhanced version of UNIX. One of the components of MMSS that was developed to address the many specific objectives outlined in Grayson et al. (1990) is the work scheduling system. MMSS accom¬plishes work scheduling by using a mathematical (goal) programming approach to minimize the amount of switching of workers from their regular jobs to perform other jobs (Grayson and Nutter, 1991). Manpower support program The main objective of MANSUPP is to plan a detailed manpower structure for a mine by incorporating a production schedule, condition-modified work standards and manage¬ment preferences to predict required manhours for various support activities, for example, belt moves. This is accom¬plished by gathering and processing planning information the same way a mine planner would. The program is menu¬driven with the selection of a particular menu item locating the user in the appropriate area within the spreadsheet to review/add information. All of the interactions between productivity, section conditions and management prefer¬ences and their effect on the amount of support work required must be considered. Information regarding the productivity of a particular section, the geometry of the section, work standards for individual activities and section conditions must all be gath¬ered for input into the program. In the initial stages of development, it was decided that the program be developed in an environment that mine planners were familiar with in order to accommodate the use of the program. Lotus 1-2-3 or a compatible (Quattro Pro) met this environmental need. Other reasons included ease of calculation due to built-in functions, powerful command language and macro capabili¬ties. Each of the major components of the program defined are summarized. Production schedule This area allows the user to input a production schedule for the entire mine over 12 periods. Space has been provided for five sections, advance or retreat, and two longwall sections. There are several calculations occurring in this area that do not apply to the determination of manpower, but serve as an aid in production planning. Section description The section description area of the spreadsheet is shown in Table 1. The first three items are calculated in the production schedule area of the worksheet and then transferred to this location. The remaining information is input by the user to be used in later calculations.

Jan 1, 1996

By D. L. Johnson

A number of small but rich copper sulfide (chalcocite) ore deposits are located on Michigan&apos;s Keweenaw Peninsula. These deposits are considered too small to warrant the capital investment required for conventional mining. However, they are being considered for in-situ solution mining as a way to revitalize the area&apos;s once prosperous copper extraction industry. Geology, mineralogy and geohydrology of the deposits have been investigated to determine their suitability for leaching; Three conclusions were drawn from the work. First, the sulfide mineralization is found primarily in fractures (65%) and, therefore, is accessible to injected leach solutions. Second, the ores are shallow, occurring from surface exposure to about 200 meters depth; injection and recovery wells would not have to be excessively deep. Third, the rock was found to have low permeability, in the neighborhood of thirty millidarcys or less. However, preliminary hydraulic fracturing experiments indicate that this will be an effective method for permeability enhancement. In one experiment a permeability increase of 100 times was achieved using a very low injection pressure at shallow depth. Additionally, fracture aperture spacings measured from drill core were correlated with fair success to field permeabilities.

Jan 1, 1989

By Zbigniew Hladysz

INTRODUCTION In the Polish coal mining industry, the Upper Silesian coal basin has an important position. It contributes over 98% of the total coal production of Poland which reached 193 mt in the year 1979. In view of the significant urbanized area in this region as well as the industrial potential, it is important to protect the surface by stowing the mined out underground workings. Approximately 18% of the total production is mined from longwall faces with stowing. However, caved longwall faces are most popular. In the upper Silesian coalfield, approximately 337 highly mechanized longwall faces are in operation. Out of these 337 faces, 301 longwall faces are caved. One of the main requirements for effective and safe coal extraction is a suitable method for strata control. We can divide the complex process of strata control into four steps: 1. Strata investigations 2. Selection of supports for roadways 3. Selection of supports for longwall 4. Rock bursts and other hazards protection STRATA INVESTIGATIONS Estimation of the structure and physico-mechanical properties of the rock mass surrounding the workings is an important constituent of strata control. These factors definitely influence the rock stability, support load, working conditions as well as the mining technology. Complex methods of strata investigations involve a set of indicators enabling the estimation of the stability of the roof rock, strength of the coal seam and load bearing capacity of the floor. The basis for the evaluation of roof stability for a coal seam is the classification according to the resistance to bed separation (1). The bed separation resistance is equal to the strength of the rock in tension perpendicular to the bedding plane. In order to properly classify a rock, it is indispensable to make continuous profiles of the bed separation resistance from the immediate roof up to a distance of three times the thickness of the coal seam. The profiles are developed by utilizing mine roof investigations with the hydraulic borehole penetrometer. (Fig. 1) The penetration resistance is determined by the critical pressure required to cause crushing of the rock by the penetrometer indentor (2). The classification of roof rocks and their ability to cave spontaneously during longwall mining is shown in [Fig. 2]. The measurement of the critical pressure also allows for the determination of the compressive strength of the rock, the index of workability and roof fissuration.

Jan 1, 1982

By R. S. Scheetz

The search for a more effective means of coping with methane liberation has resulted in the expanded use of seam degasification. Its advantages are measured not only in reducing the hazards due to methane, but also in producing a marketable commodity: .the gas, itself. However, the success of a methane drainage system is the culmination of many complex factors, the most important being the coal seam permeability. The two most popular drainage techniques used mining today are the vertically simulated borehole method and the in-seam horizontal borehole method. The objective of this paper is to determine the permeability range in which the selection of the degasification technique is based primarily on equipment and economic analysis. A discussion of the equipment involved will be incorporated into examples evaluated by the Penn State coal methane simulation model.

Jan 1, 1989

By Thomas O’Neil

Part 2 of this article continues with a discussion of features that distinguish contemporary investment analysis in mining from earlier methods. Part 1 traced the evolution of evaluation methodology from the traditional and discounted cash flow methods to the present. Quantitative investment models of the past 20 years played an important role in mine evaluation, but a new integrated approach to mine investment decision-making is now developing because of growing economic, technical, and political uncertainties. One of those uncertainties -inflation-was discussed in Part 1. Increased Risk At a time when mineral production technology and analytical evaluation tools have never been more sophisticated, it may seem anomalous to claim that mining ventures are riskier. Whereas technical risk has, indeed, been notably reduced, business risk the chance of a major loss due to market factors-has increased. This has been caused by two factors: the rapidly rising capital requirements of new projects and increasingly unpredictable future economic conditions. These difficulties are compounded by the long lead times of large, new mining projects. Rising Capital Requirements A $100 million investment is no longer considered to be a large mining project in most cases. For a large metal mining project, the total cost of the concentrator alone can exceed $100 million. Amoco Minerals&apos; Thompson Creek molybdenum mine being developed in Idaho will cost over $350 million, for a relatively modest sized (23-kt/d or 25,000-stpd) ore operation with no extraordinary features. For a major grass roots copper operation producing blis¬ter or refined copper, the capital requirements are billions of dollars, not millions. Project costs are higher for two reasons. One, the projects themselves are larger; and two, the capital cost per unit of production is rising rapidly. Projects are larger because lower grade ores, yielding smaller margins, must be mined in greater volume-particularly in view of the high fixed costs of modern technology. When analysts contemplate any capital project, an important consideration is the minimum economic size of the facility. This is the smallest size that can be built and still be competitive on a total unit production cost basis. In the conventional copper smelting business, for example, a production rate of at least 91 kt/a (100,000 stpy) of blister copper is about the absolute minimum for reducing unit fixed costs to competitive levels. Clearly, many variables affect the minimum economic size of a mining facility, but with few exceptions that size continues to grow. Unit capital costs of new mining projects also are continuing to rise. The accompanying table shows that the capital costs in primary copper production have risen faster than the cost of living. Of particular note is the capital charge of over $1.70/kg (77411b) required for a new, fully integrated facility in 1981. Thus, the mid-1982 copper price would not even cover the cost of capital for a new copper producing complex without considering the direct operating costs for such a facility. The breath-taking costs of new mining projects are caused by factors such as increasingly complex technology, the rising social costs of creating attractive living and working conditions for employ¬ees, and the expenditures re¬quired for environmental protec¬tion. For whatever reasons, how¬ever, unit capital costs in mining continue to exhibit staggering growth rates. Future Economic Conditions Risk is not only a function of the size of the potential loss-an im¬portant measure of which is the capital intensity of mining dis¬cussed above-but also of the like¬lihood of incurring the loss. It is here that the ever increasing lead times of mining projects and the glaring limitations of economic forecasting are so important. In comparison to nearly all other in¬vestments, mining has more capi¬tal at stake for a longer period of time before commercial uncer¬tainty is resolved by distant eco¬nomic conditions that are subject to high forecasting errors. A sobering example of the pre¬ceding points occurred with the Lakeshore copper mine in south¬ern Arizona that Hecla Mining Co. undertook to finance and manage in 1969. The project was a joint venture of Hecla and the El Paso Natural Gas Co. and included two mines-one producing oxide ore and a second in the sulfide zone¬and a complex processing facility producing both electrowon cop¬per cathodes and copper precipi¬tates. The project originally was

Jan 12, 1982

By R. R. Beebe

When an economic mineral deposit is first discovered, it is almost immediately defined in terms of tonnage and grade. As exploration and assessment continue, the initial definition is ordinarily qualified by such terms as "proven," "probable" and so on. Even at this stage, the fundamental philosophies of the discoverers may manifest themselves in conservative or optimistic views of the deposit&apos;s potential. It is unusual, however, for the philosophies to be expressed in terms of metallurgical performance criteria at such an early stage unless the ore is very unusual in its chemistry or mineralogy. Even before drilling is completed, the first mining plans appear, with or without the benefit of metallurgical test data. This is the point at which management philosophies can have some of their greatest impacts. In ideal cases, where large, uniform ore bodies can be developed and mined according to assay cutoffs, the conventional philosophy of maximizing tonnage while minimizing cost can be followed without much metallurgical difficulty. Unfortunately, such cases are rather uncommon. Even large, low-grade

Jan 1, 1991

By Sam H. Patterson

This paper is concerned primarily with the kaolin deposits extending northwestward from Macon, Ga., which is in the western part of the Irwinton district (fig. 1). These deposits are in a belt of kaolin districts extending along the inner edge of the Coastal Plain from the Aiken district, S. C., southwestward to the vicinity of Macon, a distance of 150 miles (250 km). Large kaolin deposits associated with bauxite also occur in an extension of this belt southwestward from Macon as far as the Eufaula district, Ala., a distance of about 130 miles (208 km). The belt splits away from the inner edge of the Coastal Plain in a southwesterly direction from Macon; the Eufaula district is approximately 50 miles (80 km) south of the contact of Coastal Plain sedimentary rocks with much older crystalline rocks. Structure The strata of Cretaceous and Tertiary age containing the kaolin deposits all dip gently seaward irom the landward edge of the sedimentary rocks. The dip of the Cretaceous strata is about 25 feet per mile (4.9 m per km). The dip of Tertiary strata is about 15 feet per mile (2.1 m per km). In places, the Tertiary strata overlap the older sedimentary beds and rest directly on the crystalline rocks. Faults having a few feet of displacement have been observed in several mines, but no major faulting is known in the area.

Jan 1, 1977

By John M. Hird

Texasgu1f&apos;s Lee Creek Mine in Eastern North Carolina is usin8 a combination Dredge-Dragline method for mining deep phosphate ore at a rate of 8.5 million tons per year. The matrix is thick, uniform in grade, and is processed into agricultural products. As in many other good mineral deposits, there are tough obstacles to overcome in recovering the ore. The deposit is in an area with a land surface of about ten feet above sea level, is close to tidewater, is overlain with 90 to 130 feet of unconsolidated overburden, and is underlain by an artesian aquifer with over 140 feet of natural hydrostatic head at the bottom of the bed, The upper 40 feet of overburden will not support stripping equipment and is unstable when cast as spoil. Barriers are emplaced to prevent flooding from hurricane tides. The ground water problems have been successfully controlled by pumping 60 million gallons per day from the aquifer. Only recently have the severe ground problems been overcome by this unusual Dredge-Dragline combination of mining equipment.

Jan 1, 1978

By Gerrit V. R. Goodman, Charles D. Taylor

A technique was developed to determine the fraction of air recirculated from the scrubber exhaust toward the face and the fraction of gas captured by the scrubber. To determine these two values, a series of tests was required. In the first test, gas was released at the face area and concentrations measured at the scrubber exhaust. In the second test, gas was released directly into the scrubber with concentrations again measured in the scrubber exhaust. The simultaneous solution of two independent flow expressions gave values of recirculation and scrubber capture. A series of tests was conducted with various combinations of scrubber flow, curtain flow, setback distance, and machine location. Statistical analyses then were conducted to determine those factors which influenced the levels of scrubber recirculation and scrubber capture in a deep cut. INTRODUCTION As cutting depths increase, the trend is toward the use of blowing face ventilation systems. With these ventilation systems, the potential exists for recirculation between the scrubber exhaust and the intake air stream. One concern is that methane gas, having been removed from the face area, is subsequently recirculated back into the intake airflow. Under these conditions, methane-laden air could be substituted for fresh intake air, leading to increased methane concentrations at the face. There is currently no standard for recirculation between the scrubber and intake airflows. The implied understanding of the Mine Safety and Health Administration (MSHA) is, and has always been, that any recirculation of ventilation air is unacceptable. To minimize recirculation potential, MSHA suggests that operators maintain scrubber and curtain flows within 0.47 m3/sec (1000 cfm) of each other. Given that recirculation may occur in some circumstances, the question arises as to how much recirculation constitutes a hazard for the face workers. Unfortunately, there is currently insufficient information to answer this. Emphasis must instead be placed on means to quantify the amount of recirculation between the scrubber and face area. Methods must also be examined to control this phenomenon. This paper presents a technique for quantifying recirculation and scrubber capture levels in a blowing face ventilation system. Several ventilation parameters are investigated to determine their effects on these levels. DEVELOPMENT OF RECIRCULATION EXPRESSIONS One way to measure airflow quantities at a point near the face would be to measure the gas flow rate Qgas, and gas concentration c at that location. The airflow Qair, would be simply the ratio of these two values, Qair, = Qgas/c. Although gas concentration values, c, can be obtained easily with existing gas monitoring equipment, it is more difficult to obtain gas flow readings, Qgas, at specific locations. To obtain these values of Qgas, another method was needed. Previous work in this area was conducted by Kissell and Bielicki (1975). However, the calculations in their model assumed a constant gas concentration in the region around the continuous miner. In a deep cut, it is questionable that gas concentrations would remain constant or uniform around the mining machine. The technique presented here does not rely on this assumption. It instead assumes that the gas concentration in the scrubber exhaust remains constant. A quantity of gas released at the cutting face either flows into the scrubber inlets or bypasses the inlets and flows around the off-curtain side (in this example, the right side) of the continuous miner (Figure 1). The fraction of the face gas that flows

Jan 1, 1993