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By R. C. Kirkman

Your interesting and well-written paper brings up the issue of the purpose or use of a mineral property appraisal, specifically the value of these evaluations when the quality and quantity of the mineral resource to be exploited have not been thoroughly investigated. I am not a practicing professional in regards to valuation of limestone investment opportunities. I simply wish to make a few generic observations about the appraisal process in order to stimulate meaningful thinking by appraisal professionals whose reports may be reviewed by professional investors and "laypersons" alike. It is my contention that the analysis given in your paper is very thorough, considering the data that was made available to you. But I would not like to be the person responsible for investment decision making without having a feel for the risk involved. Therefore, the reviewer must classify this type of analysis as relatively "subjective" because they have not received an indication of the inherent risk involved in exploiting a mineral resource. This risk can best be quantified by using probabilistic methods in conjunction with the quality and quantity of the mineral to be produced. Most appraisals passively ignore this problem. Specifically, you have accurately identified the various costs and values entering the economic appraisal process. These can usually be quantified with a great deal more certainty than the capability of the resource to provide 25 years of quality material. Also, in Cases A (West Virginia)and B (Alabama), you are fortunate enough to have previous operations data. What if there had never been production from these properties? How does a potential mineral property investor evaluate the resource capability to meet various projected financial outcomes without spending a good deal of up-front capital to drill and delineate enough quality resources for a 20-plus year operation. These questions pose a substantial risk to any potential mineral investor. Only probabilistic methods should be used to quantify this type of problem, short of drilling and metallurgical testing. I might be considered a "layperson," not being familiar

Jan 1, 1991

By Douglas R. Cook

The Ertsberg East discovery was announced by Freeport in 1979 with proven geological reserves of 45 Mt (49.6 million tons) with a grade of 2.65% copper and recoverable gold values. This recently discovered deposit is one of a cluster of skarn-type copper occurrences found within the limits of Freeport Indonesia's mining con- cession at an elevation of about 3600 m (12,000 ft) in the extremely rugged Jaya Wijaya Mountains of Irian Jaya, the Indonesian half of the island of New Guinea. The Ertsberg East lies about 1.5 km (0.9 miles) east of the Ertsberg ore body, which is also known by the Indonesian name "Gunung Bijih" or "ore mountain." For general background, the original Ertsberg ore body was first discovered by the Dutch geologist Jacques Dozy in 1938 while on a mountain climbing expedition. It was mapped and sampled by Forbes Wilson and Del Flint after a two-week harrowing trek into the mountains from the coastal swamps. A drilling program in 1967- 1968 established that the Ertsberg was a viable mining venture, containing approximately 33 Mt (36.4 million tons) of skarn-type copper ore. The deposit averaged 2.5% copper, with significant values in gold and silver. The first shipment of copper concentrates was made in December 1972, following a capital investment of $200 million. Following the routine production of ore from this deposit, it became necessary to consider additional sources of ore to extend the life of the mining operation. An exploration program was thus started with the initial objective of finding shallow or surface ores which could be blended with the material being mined at Ertsberg. Project geologist Frank Nelson was transferred from Freeport Exploration Co. to the Freeport Indonesia organization at the Ertsberg mine. However, the exploration work was the functional responsibility of Freeport Exploration. Del Flint had overall responsibility for the program from the Reno headquarters office and provided liaison with management in the USA and Indonesia.

Jan 1, 1991

By J. Wes Blakely

In the mid-1950s the importation of foreign oil began to encroach on traditional coal markets such as railroad steam engine fuel, home heating, electrical generating plants, and others. Many of the smaller coal companies with substantial coal reserves subsequently found that they could no longer continue to mine economically and merged in various ways with larger established coal mining concerns. As the seats of top management and authority were removed from the mines to large metro centers such as New York, Pittsburgh, and Chicago, the more personal relationship between mine workers and management that had existed prior to these mergers began to be eroded and replaced by " remote control " mine management. The disappearance of this face-to-face contact of upper management with mine employees tended to create barriers blocking simple effective communications between coal mine management and labor. In examining the human relations aspects of the coal mining industry, many elements, some concerned with work and others not, have depicted the industry as almost devoid of good applied human relations and communications practices, and, in truth, little has been done to alleviate this situation. In past years many coal producing companies, with the exception of one of the largest, failed to adequately define the term "communications" as it applied to their companies, whether internally or externally. Internally many still place heavy reliance on written communications such as company-wide newspapers to do the job of communicating with their employees which they feel ought to be done. One company puts an inordinate belief in its company "newspaper" as the primary instrument to stimulate its work force into increasing productivity. Primarily aimed at the workers, who greatly outnumber the salaried administrative people, this publication mainly documents the fortunes of individual mine workers who are felt to have achieved something with their lives, e.g., a mine worker who had managed to put four children through universities, the philosophy behind this being to inspire other workers to achieve worth- while goals.

Jan 1, 1981

By R. K. Singhal, R. D. Lama

Underground coal mining systems have kept pace with developments that have occurred in other types of mining. A diversified group of machines are now available with high horsepower motors, built-in microprocessor technology, and numerous options to satisfy miners' needs in various geological environments. The results will be a greater degree of mechanization and recovery of coal from inclined seams, thins seams, and seams mined in lifts from shallow as well as deeper deposits. This article is based on a general suvey carried out by the authors to detemine the current status of mechaniztion and systems development in underground coal mining. It indicates that the next two decades will see increasing use of longwall mining, which has already reached a high degree of sophistication. Automation, remote control, and robotics are the key words of the future. Data transmission from underground to surface and over long distances will be almost routine. Many components of the high technology industry that are just being considered by the mining industry regarding equipment health and perform¬ance monitoring, quality control, application of expert systems, and computer networking will become a routine part of underground mining technology. Within the next 21 years, "manless" mining, first attempted about 25 years ago, will become a reality. References to any trade name are meant to aid understanding of the text and do not imply endorsement by the authors or their users. Current coal face developments Developments in underground mining equipment, its application, and associated technology have made phenomenal progress compared to those in other types of mining. Room-and-pillar systems, on one hand, and longwall mining, on the other, have sophisticated equipment at their disposal. Both of these systems have become highly productive, with output per manshift about 50% higher than that achieved a decade ago. Continuous miners continue to be the mainstay of room-and-pillar coal mining. Numerous developments, how-

Jan 1, 1989

By George Aul, Richard Ray

Short-hole horizontal degasification systems have been utilized in Island Creek's Virginia Division mines since 1986. These systems remove methane from the coal seam in advance of longwall mining. Without an in-seam methane drainage program, methane liberation rates from the longwall faces may exceed 0.47 cu m/s (1,000 cfrn), requiring 47 cu m/s (100,000 cfm) of air for adequate dilution capability. Optimization of the degas systems has enabled methane liberation levels and the associated ventilation requirements on the longwall faces to be reduced by 50 to 80 percent. Since the inception of the first degas system at the Virginia Pocahontas No. 6 Mine, Island Creek has developed an overall methane drainage program designed to maximize gas production and minimize ventilation requirements for methane dilution. Gas production from the degas systems is affected by four main design criteria: (1) hole location, (2) hole length and spacing, (3) collection system capacity, and (4) time available for drainage. Drilling equipment and materials for the gas collection system are selected to meet the necessary criteria while minimizing manpower requirements. Benefits gained from reducing methane liberation rates from the longwall faces are fourfold: (1) improved safety due to lower methane concentrations in the face area, (2) fewer production delays caused by excessive methane concentrations, (3) lower fan operating costs resulting from reduced air requirements, and (4) enhanced worker comfort resulting from lower air velocities and less dust generation.

Jan 1, 1991

By L. B. Anderson, G. A. Carmody

We wish to comment on the paper by G.W. Heunisch (MINING ENGINEERING, January 1987), which raised several important points regarding treatment of acidic waste waters. This paper confirms, in part, tests carried out by Allyn Corp. Allyn is the manufacturer of the proprietary product Alka-Ment, one of the materials that Heunisch tested. The first point is the extreme sensitivity of the sedimentation rate on the nature of the treatment material used. This apparently results from differences in nucleation of the metal hydroxides, and is consistently higher for Alka-Ment. This decreased sludge volume is also observed when Alka-Ment is combined with flocculants, such as organic polymers (see Table 1). Our second point is to emphasize that Heunisch's work shows how dependent the sludge volume is on pH. It is common practice to overtreat AMD by raising pH to 9 or 10. This practice, in addition to being wasteful of material, actually decreases the density of the resulting sludge, probably by creating charged particles that do not readily coalesce. When treating with Alka-Ment, we recommend that the treatment be monitored by testing both pH and dissolved iron. For both tests, it is important to allow adequate time to elapse between treatment and measurement, so that dissolution of alkali, aerobic oxidation of iron, and precipitation of the metal hydroxides are nearly complete. Our tests have indicated that from 30 minutes to one hour is commonly required. Only in this way can a minimum pH be attained, with adequate treatment, denser sludge, and lower treatment rates achieved. Thirdly, Heunisch correctly recognizes that ASTM titration of Alka-Ment does not reflect the true treatment value available in a unit of Alka-Ment. Our work at Allyn indicates this is a most important consideration when evaluating various materials for the treatment of acidic waste waters. While we were unaware of Mr. Heunisch's work, we welcome comparisons between Alka-Ment and any competitive products readily available for field use. Allyn has hundreds of hours of laboratory and field data we are ready to share with interested parties treating acidic waste water.

Jan 1, 1988

By Paul C. Schmall, Arthur B. Corwin, Hugh S. Lacy, Derek Maishman

For reasons of hydraulic design and constructability, water supply tunnel systems are often located relatively deep, in competent rock. Where the rock lies below a thick cover of water bearing soils, artificial ground freezing offers a low risk, cost-effective method of constructing access shafts. Several examples, in Kansas City, MO and in Brooklyn and Queens, NY, illustrate the use of freezing to provide structural support and ground water exclusion. In the case histories presented, the shafts range from 6m to 12m (20 to 40 ft) in diameter and soils vary from clays to boulder beds with depths in excess of 73m (240 ft) to competent bedrock. Where tunnels for infrastructure are designed to be driven in rock, vertical access may encounter a considerable thickness of overburden soils with groundwater. In such cases, shaft sinking problems are analogous to those confronted by mining engineers over 100 years ago. At that time, newly developed refrigeration machines were set to freeze the ground water and stabilize running sands. Although modern technology has refined our efforts, the basic features of the ground freezing to be described here are the same as those used in the last century. Vertical holes drilled on a circle surrounding the shaft site are equipped with leak-free steelpipes through which a refrigerated brine is circulated. In this manner, a frozen cylinder, seated in rock, is formed when the earth freezes between adjacent pipes. Shaft excavations take place inside a self-supporting, ground-water excluding structure. For mines, no better method of sinking production shafts through deepwater-bearing ground has yet been established. Indeed, major deposits of coal, potash, and salt would have remained inaccessible to this day were it not for the continued use of artificial ground freezing. If mine shaft freezing became relatively commonplace, early applications to civil engineering were few. Mussche and Waddington (1946) mention only five projects between 1892 and 1935; three of these relate to subway or vehicular tunnels in Paris, Antwerp, and (the most extensive) at five locations of the Moscow Metropolitan Railway development. In the United States, the method was used to spectacular effect in 1936, freezing about 3,000 cubic yards of silt to stop a slide

Jan 1, 1999

By John C. Nichols

The Ouachita, Ozark-St. Francis, and Mark Twain National Forests are located within the Ozark and the Ouachita Mountains regions of southern Missouri, central and northern Arkansas, and eastern Oklahoma- an area generally referred to as the Ozark-Ouachita or Interior Highlands. These national forests share many similarities in geography, biotic resources, and areas of shared public, market, and user interests. The three national forests have joined together to evaluate the natural resources within the Highlands and to study the impacts of the forests on local and regional economies and on the people who use these national forests and their resources. Other federal, state, and local government agencies and universities have joined with the Forest Service in this project, broadening the assessment area beyond the national forest boundaries to include 107 counties in Arkansas, Missouri, and Oklahoma. The diverse, resource-related data for the as- sessment area address social, economic, terrestrial, and aquatic issues and conditions throughout the Highlands. The assessment report currently is being developed by the Forest Service and will include an accoMpanying CD-ROM. The Ozark-Ouachita Highlands assessment has provided a forum for minerals- and geology-related resources, industries, social significance, and potential to be profiled and discussed along with other critical assessment resources. The state geological agencies in Arkansas, Missouri, and Oklahoma were active contributors to this effort. Their reports and other key data bases developed and used for the minerals sections of the published assessment report will be available either as a supplemental CD-ROM or on the Internet. Updated information on the progress of the assessment report can be found at the Ouachita National Forest Website, at <http://www.fs.fed.us/oonf/ouachita.htm>. This will also include a matrix displaying the minerals in the assessment area, their occurrence by assessment-area counties, and a potential ranking, as well as information on interesting geologic features. The state and federal reports, resources, and data bases, have helped answer key questions concerning mineral resources and related issues in the Highlands:

Jan 1, 1999

By Morris Loshinsky, George Ziegler

The New York City Transit System is celebrating its Diamond Jubilee this year. In 1904, 9.3 route miles of New York&apos;s first subway was placed in operation in lower Manhattan. Since that date the system has grown to 230 route miles and 458 stations (See Exhibit #1). The age of the system immediately provides a clue to the nature of our construction activity. While we must always plan for the future and expand the system in the direction of the future needs in New York, a substantial amount of time and money must be channeled towards the modernization and maintenance of the existing system which provides vital transportation for 3-1/2 million passengers a day. This paper however will be directed to our New Route&apos;s Construction and, specifically, to three rock tunnel and cut and cover sections presently under construction. Currently, New York is engaged in the construction of about 9 New Route miles of structure. Briefly, the 3 Sections we will discuss are as follows: Route 131-A Section 5A is a 4 track structure in Manhattan and includes the new Lexington Avenue Station. This 1240 foot long section is located between Park Avenue and 3rd Avenue and is a combined cut and cover and rock tunnel section under 63rd Street. Route 131-A Section 5B is an intricate 2423 foot long 4 track rock tunnel section extending from 3rd Avenue to the FDR Drive under 63rd Street in Manhattan.

Jan 1, 1979

By Robert A. Jankowski, Fred N. Kissell

Recent allegations about tampering of coal mine dust samples have led to suggestions that the coal industry needs a respirable dust compliance sampling method that is less subject to tampering. The sampler designs under consideration include fixed location monitors that could be mounted on a mining machine or located at a stationary location in the mine, and samplers that are worn, like the existing samplers. Fixed-point samplers have a design advantage, since samplers worn on the body will have size and weight limitations. On the other hand, in some cases there are respirable dust concentration gradients, which will dictate careful placement of a fixed-point sampler inlet to determine the exposure of workers. This study examined dust concentration-gradient data obtained by the Bureau of Mines and others over the last decade. As an example of spatial gradients at continuous miners, a fixed-point monitor located in front of the operators cab shows a higher dust concentration than a sampler located on the operator and reduced exposure when the sampling location is behind the operator cab. It may be necessary to use a sampling tube with a machine mounted monitor, to sample in the operators breathing zone in the cab to provide a more representative sample of the miners exposure. For face workers on longwalls or on continuous miner sections using remote control, the study shows significant dust level gradients and gradient variations between typical operator positions and mining machine locations. In addition, if fixed-point monitoring is used to monitor an area, vs the exposure of specific workers, dust control technologies currently employed to ensure reduced dust exposure such as the diversion of dust laden air away from the workers, or through the use of remote control of mining machinery may not indicate the same dust reduction obtained with current sampling techniques.

Jan 1, 1993

By Kanaan Hanna, Mir Heydari, Dipack Sengupta, Gordon B. French

This article, first in two, discusses the safety, technology, and economic impact of eliminating safety fuse from metal and nonmetal mines. It is based on operation and manufacturer experience over the past five to 10 years. If safety fuse is eliminated, blasting accidents will decrease. There should be only limited technical impact. The economic impact will be minor for large operations, but will be significant enough for small operations that some will likely close rather than change. Most mines would probably change over, with the market absorbing the extra cost. Elimination of safety fuse would obviously have an impact on the two safety fuse manufacturers. One manufacturer produces an alternative system and would be unaffected. The other, however, will be affected if it does not change its production strategy. From the standpoint of employment and production, abolishing safety fuse will have a very minor effect on the mining industry as a whole. This article discusses the current use of initiation systems, incidence of accidents, and fuse consumption. Next month, part two describes the economic impact of eliminating safety fuse on the mining industry and manufacturers, and presents an economic analysis of the initiation systems. Introduction The objective of this study was to determine the safety, technology, and economics of eliminating safety fuse from metal and non-metal mines. There were five major tasks: • Studying initiation-related accidents from the Health and Safety Accident Center (HSAC) data file; • Visiting mines, collecting relevant data, and interviewing mine operators; • Establishing the trend in safety fuse consumption and type of users; • Corresponding with fuse manufacturers, domestic and foreign; and • Developing an idealized blasting cycle to evaluate the possible economic impact of safety fuse elimination on large and small operations. More than 450 explosives-related accident reports have been screened and analyzed by investigating HSAC files. These reports have been studied and classified according to the four initiation systems, size and type of mines, and time and frequency of accidents from 1973 through 1978. Twenty-four mining operations were visited to obtain supporting data for this study. Mine personnel were interviewed to determine attitudes about safety fuse and alternative detonating systems. This article decribes available data. It also attempts to provide information that can be used to answer such questions as: • Is the use of safety fuse (an antiquated detonation system) a major cause of mining accidents? • Are alternative systems available that could result in fewer in-juries to miners?

Jan 1, 1984

By G. Piegols, J. A. Lombardi

Substantial volumes of cyanide are consumed in precious metals mining due to the purposeful destruction of base metal cyanide complexes and free cyanide to meet surface discharge regulations. This destruction of cyanide resource runs counter to the prevailing social ethic of resource conservation wherein cyanide recovery-and-reuse is considered a more socially responsible cyanide use pattern than systematic cyanide destruction. HW Process Technologies, Inc. (HWPT) has developed membrane based methods for the recovery of cyanide from barren solution as a means to reduce cyanide waste. The HWPT Engineered Membrane Separation (EMS™) system uses state-of-the-art spiral wound membrane elements that are engineered specifically for the treatment of high pH cyanide process solutions. The membranes fully reject WAD CN, SAD CN and most minor metals, and partially reject SCN, Free CN and precious metal cyanide from a high clarity barren solution feed to a small volume EMS™ “concentrate” stream. The EMS™ “concentrate” stream is returned to the process. The large percentage of the EMS™ process feed exits the system as cyanide lean EMS™ “permeate.” The EMS™ permeate, after residual cyanide destruction, polishing, and pH adjustment, is, thereafter, typically suitable for discharge. Separately, the HWPT EMS™-PLS method of cyanide recycling is an internal plant cyanide recovery step that removes base metal cyanides from a pregnant leach solution to a small volume EMSTM concentrate stream. The concentrate stream is then typically processed further for the recovery of metals and recovery-recycling of cyanide. The method makes it economically practical to recover gold from Merrill-Crowe or car-bon plant “fouling” solutions, e.g., those with greater than 400 ppm Cu, as might result from the use of the EMS™ method wherein EMS™ concentrate is constantly thrown back into the mine leach solution pool as part of the free cyanide recycle process, or as might result from the single-pass leach treatment of high copper content ores or concentrates. The EMS™ cyanide recovery method for barren solution treatment is shown to be “economic” for the treatment of clear solutions at the approximately 200 ppm total cyanide level (CNT), and for slurries at the approximately 400 ppm total cyanide level, where economic is defined as a 2 year EMS™ plant payback. The EMS™ plant payback calculation includes the beneficial monetary impacts of gold and free cyanide recycling and the avoided operating cost of cyanide destruction.• The cyanide recovery plant payback period is reduced further when the avoided capital costs for cyanide destruction (EMS™ permeate “clear solution CN destruction polish” only required) and/or metals precipitation are considered.

Jan 1, 2004

By H. Knorre, A. Griffiths, J. Loroesch

Introduction Hydrogen peroxide is known to the mining industry as a powerful oxidant, able to attack and to destroy cyanide, the most important leaching reagent (Knorre and Griffiths, 1984; Griffiths, 1988; Griffiths, 1988). CN- + H202 => OCN- + H2O The most recent development, however, is the use of H202 as an oxidant in the cyanidation of precious metals. This application may appear to be a contradiction to the above mentioned detoxification process. This can, indeed, be the case if certain conditions are not considered carefully. Degussa has developed a new process to enable the application of H2O2, in this field without having any detrimental effect on the cyanide. The incentive to seek an oxygen supplier like H2O2 instead of air was the fact that one of the most important conditions for the fast dissolution of gold is the presence of physically dissolved oxygen in the pulp (Kydryk and Kellog, 1954; Warstell, 1987). 2 Au + 4CN- + 1 /2 O2 + H2O => 2 [Au(CN)2]- + 2 OH Complete gold recovery within a reasonable leaching time may only be achieved if the pulp is saturated with dissolved oxygen. Gold losses in leach operations mainly occur due to the following reasons: • In some cases, the saturation of oxygen in the pulp cannot be achieved by aeration due to the difficulty to transfer oxygen from the gaseous into the liquid phase. This happens, for example, in pulps with a high viscosity. In such cases, even violent aeration only results in a low concentration of dissolved oxygen in the pulp. • In any case, the starting phase of leaching requires the highest amount of oxygen. Therefore, it is especially important to raise the dissolved oxygen level to saturation in the starting phase. Peroxide assisted leach process (PAL) The intention of the process is to introduce oxygen in a liquid form. As a result of intensive research work, it could be shown that hydrogen peroxide can be used as a source of oxygen in the leaching of gold ores. To avoid side reactions, the following conditions must be applied: • Hydrogen peroxide must be diluted before it is added to the pulp. • The dosage of H2O2, must be properly regulated based on the oxygen level of the pulp. • An important part of the continuous 02-control is a special oxygen electrode, which resists the extremely abrasive action of the pulp.

Jan 1, 1989

By Thomas P. Kwiatkowski, Len Cobb, Christopher D. Orlandi, Judith H. Jones

The Inland System supplies raw water to the industrial and domestic users in the City of Birmingham, Alabama. The Inland System, which is nearly 60 years old, consists of an approximately 57.9m (190 feet) high earthen dam across the Blackburn Fork of the Black Warrior River, a 39.6m (130 feet) high reinforced concrete intake structure, a 1524mm (60inch) diameter Dresser coupled steel transmission pipeline, and various distribution lines extending into Birmingham. The Inland System was owned, operated and maintained by the Industrial Water Board (IWB) of the City of Birmingham, an agency of the city government, from 1958 until 1994. In 1994, the City of Birmingham consolidated all water and distribution functions under the Birmingham Water Works Board (BWWB). In 1988, the IWB determined that the Inland System had reached a point where its reliability could not be assured without an in-depth evaluation and renovation. The IWB engaged Christy/Cobb, Inc. (CCI) of Birmingham as its consultant for the renovation work. CCI evaluated the Inland System and submitted a report to the IWB in April, 1992. The report contained several recommendations for capital improvements to the system including increasing the emergency spillway capacity, reconstructing the fuse plug, relocating public access to the lake away from the dam area, constructing a new intake structure, and constructing a tunnel through the dam abutment to deliver the water from the proposed intake structure to the existing 1524mm (60 inch) diameter supply line. The new intake structure and tunnel would provide for a future maximum demand flow rate as well as provide bypass and level control capabilities. CCI was authorized in 1992 to proceed with the engineering design for the renovation of the Inland System with emphasis on the spillway and intake structure. PROJECT DESCRIPTION The major component of the renovation of the Inland System was the new intake structure and tunnel through the dam abutment. CCI engaged Jenny Engineering Corporation (JEC) of Springfield, New Jersey as a consultant for the tunnel design. JEC was responsible for the design of the tunnel, the connecting shaft between the intake structure and tunnel, and the steel pipe final lining, while CCI was responsible for the intake structure and the connection to the existing supply pipeline as well as overall project management. CCI and JEC evaluated the project site in order to select the location for the new intake structure and set the alignment for the tunnel. A factor in the evaluation process was that the level of the lake could not be lowered to construct the new intake structure since the supply of water from the Inland System had to maintained. Therefore, a location was required which was

Jan 1, 1997

By Paul L. Russell, Charles O. Hook

While the oil shale resources of the US are becoming increasingly known, few people in this country are aware of the extensive worldwide occurrence of this resource. Two recent international conferences* confirmed the growing worldwide interest in oil shales, both as future and current energy sources. The 1981 conference noted the difficulty in characterizing worldwide oil shale resources because of lack of common terminology in reporting both assay value and resource potential. Assay value is variously reported as: •US gallons per short ton; •Imperial gallons per metric ton; •Liters per metric ton; •Liters per cubic meter; •Weight percent oil; and •Volume percent oil. Oil shale resources are variously reported in barrels, short tons of oil shale, metric tons of oil shale, and cubic meters of shale oil. Worldwide petroleum shortages sparked interest in alternate energy resources including oil shales. While petroleum supplies appear to have stabilized, at least for a while, high costs are affecting the balance of payment index of many countries, resulting in a continued search for local energy resources. Kerogen is the organic constituent of oil shale and may be the most abundant form of hydrocarbon on earth, possibly exceeding coal occurrences and far more abundant than naturally-occurring petroleum resources. Oil shale, by one name or another, is widespread and is known to exist in more than 38 countries (Table 1). Oil shale industries have existed in at least 14 countries at one time or another. Many predated the Drake oil discovery in Pennsylvania in 1859, and most were unable to compete with natural oil; these operations either converted to its use or closed. At the present time shale oil is being produced on a commercial scale only in the USSR and in China. Only very rough estimates of grade and quantity are available for most of the worldwide shale occurrences since only a few deposits have been thoroughly explored. The Green River Formation of Colorado, Utah, and Wyoming is an exception. For many years this oil shale deposit has been described as the largest single concentration of hydrocarbon material in the world. With increasing world interest, larger deposits may yet be found. In 1965, the US Geological Survey published an in-place world oil shale estimate. This study, while made some 16 years ago, is still a reasonable order-of-magnitude estimate. Although exploration and evaluation are underway and a few oil shale deposits have been appraised, information on which to evaluate most worldwide oil shale occurrences has improved little since 1965. Table 2 presents an order-ofmagnitude estimate of in-place world oil shale resources. Total US resources and Green River Formation

Jan 1, 1982

By P. B. Marchant

Bioleach testing and piloting was conducted on refractory gold sulphide concentrate from Austin Gold Venture in Nevada during 1988. The objective of the test programme was to determine the technical and commercial viability of bioleaching as a concentrate pretreatment step prior to gold extraction by cyanidation as an alternative to shipping concentrate to the smelter. Baseline CIL cyanidation typically resulted in < 20% gold extraction from the Austin concentrate. Bioleaching enhanced gold extraction up to 95% using conventional CIL techniques. This paper summarizes the bioleach test programme and results, including pilot plant studies onsite, and describes aspects of the commercial design for a possible retrofit of the Austin flotation tailings leach circuit to accommodate 20 tpd and 40 tpd capacity concentrate bioleach facilities.

Jan 1, 1989

By Ellis Brien, O&apos

The fine coal fraction of a coal stream is that fraction which contains particles essentially in the -28 mesh size range. This fraction will normally first appear, in a coal preparation process, as a result of classification to segregate raw coal into various size components for beneficiation. More specifically, it could appear in desliming screen underfiow, classifying cyclone overflow, and in other process circuit components. It&apos;s recognized that the top size of this fraction has been determined more by secondary reasons, such as desliming screen area for the heavy medium cyclone circuit, the dewatering equipment for the fine clean coal, or even the stability of the refuse thickener circuit. Benefits, both economic and environmental, gained by beneficiating the fine coal fraction of a raw coal stream are increasing in number, until today, fine coal beneficiation is a necessity in most coal operations requiring any degree of preparation. This is due to some of the following reasons: • In the past it was frequently more economical to discard -28 mesh raw coal sizes than to recover the coal values from them. Today, mining costs have escalated to the point where all coal values leaving the mine should report to the finished product if possible. • Today&apos;s mining methods are producing run-of-mine coals containing higher percentages of -28 mesh sizes. This heavily weighs the economic balance in favor of beneficiating this fraction. As much as 25% of the run-of-mine coal from some mines is in the fine coal fraction. Beneficiation is recommended so that clean coal contained in the fraction can be recovered to reduce the loss of coal values. • Existing and pending environmental regulations mandate that solid fuels now consumed in power station boilers contain almost no pyritic sulfur. This form of sulfur appears in raw coal in the fine coal fraction, but can be removed by coal beneficiation. Environmental considerations dictate that every year, more of the fine coal produced for steam generating purposes in the US be beneficiated. To deliberately discard any portion of the fine coal can only add to this cost, not only in the sense that fewer tons will be recovered but, more importantly, additional tonnage will be diverted to the most capital- and labor-intensive section of the entire preparation plant-the fine refuse dewatering section. It is apparent to preparation engineers that dewatering and ultimate disposal of the fine refuse derived from this beneficiation step is certainly the most difficult technical problem presented in the entire plant design. Therefore, any beneficiation procedure which increases the fine coal recovery, or in any manner decreases the tonnage of fine refuse to be dewatered, is going to be cost effective. Since the -28 mesh fraction contains both coal and refuse particles in their finest particle sizes and thus presents the greatest degree of liberation, one would immediately infer that the highest degree of beneficiation would be achieved. This is not the case-the opposite is more likely true. Conventional coal cleaning devices based on specific gravity (such as

Jan 8, 1980

By V. Rajaram

V. Rajaram&apos;s paper does not accurately reflect the status of Eastern oil shale development as it stands today. Today, there are numerous firms and research organizations pursuing development of the Eastern oil shale resource. Details of the current projects can be found in the symposium proceedings for the 1985 Eastern Oil Shale Symposium held in Lexington, KY, in November 1985. Additionally, the comparison of the Paraho and HYTORT processes is invalid because the HYTORT process design used it not an accurate representation of the HYTORT process. I will briefly address this point. It is disconcerting to see the author reference hydroretorting work conducted at the Institute of Gas Technology (IGT) (circa 1978) in the introduction, which indicates hydroretorting (the HYTORT process concept) yields of up to 250% of the Fischer Assay oil yield, and then turn around and base the remainder of the paper on the 1981 "Synthetic Fuels From Eastern Oil Shale Report," which depicts the HYTORT process as yielding only 115% of Fischer Assay. Indeed, work conducted at IGT to 1981 indicates commercial oil yields for the HYTORT process would be in the range of 200% to 210% of Fischer Assay. A more thorough literature search by the author would have yielded a 1981 publication by Schora et al., entitled "Update on the HYTORT Process" presented at the Symposium on Synthetic Fuels From Oil Shale in Nashville, TN, October 1981. This paper thoroughly discusses the shortcomings of the referenced study with respect to the HYTORT process. For the interested reader, I can also recommend a June 1984 report entitled "An Economic Comparison of Five Process Concepts for Using Eastern Oil Shale," published by Los Alomos National Laboratory (LA-10106-MS, Parkinson et al., NTIS No. DE84-015981). This study applies a realistic HYTORT design (circa 1982) and indicates that HYTORT is the most econom¬ical method of processing Eastern oil shales (about $20/bbl less than Paraho). Additionally, completion of the HYTORT feasibility study in June 1983 by Phillips Petroleum Co., Bechtel Group Inc., HYCRUDE Corp., and IGT has resulted in development of designs that will make the use of HYTORT on Eastern oil shales even more economic. HYCRUDE Corp., a second-tier subsidiary of IGT, is involved in commercialization and licensing of the HYTORT process and is active in efforts to develop the Eastern oil shale resource. As such, I am disappointed to see outdated material that has the potential for adversely affecting development of Eastern oil shale resources published in MINING ENGINEERING. I can only hope that this response will serve to mitigate any negative opinions that may have been formed. Today, activity in commercializing Eastern oil shales dominates the oil shale scene in the US and indicates a realization of the potential for these shales. Reply by V. Rajaram I appreciate the updates on the HYTORT process provided by R.C. Rex, Jr. I hope that Hycrude&apos;s work will lead to the funding of a pilot plant to demonstrate that yields of greater than 200% of Fischer Assay can be achieved in a commercial plant using eastern US oil shales.

Jan 1, 1987

By Joseph Alexis

Humphreys Engineering Co. has been associated with mining and milling since the turn of the century, but its greatest contribution to the industry was development of the Humphreys spiral concentrator in 1943. Few devices in the history of mineral beneficiation have been so uncomplicated. The spiral concentrator has no moving parts and requires no power except that needed to deliver slurry to the feed box, a requirement common to all concentrating devices. Although Humphreys Engineering Co. produces several models of spiral concentrators, the one -commonly used for coal cleaning is Model 24C-6 which has a flatter pitch than most other models. The spiral concentrator consists of a spiral conduit with a modified semicircular cross section. A slurry containing 20-35% solids by weight is fed to the feed box at the top of the spiral and flows downward. The heavier refuse particles concentrate along the inner side of the slurry stream and the lighter coal moves to the outside. The coal spiral has six complete turns-each turn consists of three 120°-sections and each section has a refuse port at the lowest point of the channel. The Humphreys spiral concentrator separates minerals in accordance with their physical properties (primarily specific gravity and particle shape). Aside from pumping slurry to the distributor, no energy is required to produce clean coal since the separating action is controlled by gravity. As slurry flows down the spiral channel, each particle is subjected to a centrifugal force, tangential to the channel. This force piles slurry on the outer rim of the spiral until the flowing stream reaches an equilibrium between outward centrifugal force and downward gravitational force. Once established, the stream position is relatively constant. In any flowing stream, differential velocities are present within that stream (the slowest fluid levels nearest the channel surface and the fastest layers at the center). As the channel curves, the lower slurry levels, retarded by friction, have much less centrifugal force and tend to move sideways toward the inner radius of the channel. As the bottom layers move inward along the bottom of the channel, the upper layers must move outward to maintain equilibrium. This cross sectional rotation separates the minerals and moves those of highest specific gravity toward the inner radius where they drop through refuse ports (see figure on following page). The simplicity of the spiral concentrator is apparent in operation. After tonnage, slurry volume, and percent solids of the feed have been determined-generally with laboratory testing-the operator&apos;s main concern is simple adjustment of the splitters located in the refuse ports. These splitters control the amount of refuse collected and, generally, once properly set they need little attention unless grade conditions drastically change or trash builds up and hinders laminar flow. Spiral concentrators are not overly sensitive to changes in pulp densities, variances in grade, or parti¬cle size and shapes. All these conditions may fluctuate over fairly wide ranges without adversely affecting recovery and product grade. In certain applications, desliming may improve recovery.

Jan 8, 1980

By S. K. Mishra

Introduction The history of the Australian coal industry is almost as old as the history of Australia after new settlement. The first cargo of coal was exported from Newcastle to India in 1799. Coal has been mined in Australia for almost two centuries. But there are indications that the first coal was processed in 1877 in the Newcastle area. This is nearly 50 years after the first coal processing in Europe. Processing on any significant scale began in 1944 at the Newcastle Steel Works. In Queensland, however, coal processing did not begin until the second quarter of this century. The dramatic growth of mining and processing of Australian coal since the mid-60s is primarily due to the expansion of the export market in Japan, Asia, and Europe. Australian coals Black coal occurs in most of the major sedimentary basins of Australia. The better known deposits of commercial importance are in the Sydney, Bowen, Oakland, and Gailiee Basins. Apart from the Collie Basin in western Australia, other coal deposits of the Arckinga Basin in South Australia, the Bonaparte Basin in Northern Territory, and the Tasmania Basin are basically unexploited. The Sydney Basin of New South Wales and Bowen Basin of Queensland remain the major producers of black coal in Australia. Currently exploited or potentially commercial deposits of black coal in Australia occur in strata ranging in age from Permian to Tertiary. Estimated black coal reserves are summarized in Table 1. Coals from different ages have significant differences in their petrographic characteristics. Many of the Permian coals are usually rich in inertinite. Jurassic and Tertiary coals are typically rich in vitrinite content. Typical petrographic and chemical characteristics of some important Australian coking coals are summarized in Tables 2 and 3. The reserves of coking coals are restricted in the main to the Permian coals of the Sydney and Bowen Basins. Coal deposits in the latter are at shallow depth. A high percentage of these are being strip-mined. On the other hand, in the Sydney Basin, the majority of the reserves of good quality coking coals are at significant depth. The mining cost of these deposits is substantially higher than those in the Bowen Basin. It is usually possible to make strong coke with Australian Permian coals of vitrinite content much lower than the other international coals. But there are many major Permian deposits in Australia where low vitrinite content presents problems. As they are classified as marginal coking coals, optimization of their pro¬cessing plays an important role in meeting the market specifications of coking coal. The major reason for processing Australian coal is to reduce the ash content. Compared with British and some American coals, Australian coals are of relatively high ash content. Unlike American coals, more than 99% of the Australian coal deposits have sulfur content less than 1%. Consequently, sulfur control has never been a major problem. The coking quality of coal is improved simply through the ash reduction in the marketable product. However, with the increase in intensive mechanization and multi-stage handling of coal be-

Jan 9, 1985