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By Andrew B. Cecala, Robert J. Timko, Christopher J. Pritchard

As mine venti1ation systems become more complex, it is increasingly difficult to supply fresh air to the working face. A high percentage of the fresh air entering a mine never reaches the working face because it short circuits through stoppings and doors, exiting the mine without being utilized. When ventilation limits are reached, the only option presently available to most U.S. mines is to sink a new shaft - an extremely expensive undertaking that many companies cannot afford. Reusing or recirculating a percentage of ventilating airflow may make it possible for some mines to remain operational while maintaining a safe environment for workers. Such a study was performed at a trona mine in Granger. Wyoming. A vane axial fan was used to recirculate approximately 9.4 m /s (20,000 cfm), 25% of the total ventilating airflow, back to the working face area of a three-entry continuous miner section. This study involved two evaluations: the first when the face was 305m (1,000 ft) from the recirculation fan, the second when the face was 610m (2,000 ft) from the fan. The evaluation consisted of monitoring the mine atmosphere for respirable dust, methane (CH4). carbon monoxide (CO), and nitric oxide (NO), at various locations in the intake and return entries. Several shifts were monitored, with and without the recirculation system in operation. With the recirculation system in operation. there was an increase in measured respirable dust concentrations in the intake along with a slight increase in the mine temperature (1.67°C). There were also some minor changes in CO and NO levels in the intake entry, but this seemed to depend on the proximity of the instruments to the turn-around point for the diesel haulage vehicles, than to whether the recirculation system was or was not in use.

Jan 1, 1989

By Thomas J. Neil, O&apos

Most analysts find that progress has been evolutionary in any technologic field. To be sure, a few discrete breakthroughs can usually be identified, but viewed in its entirety, the curve of technologic progress over time tends to be relatively smooth with few discontinuities. Even in those cases where a major scientific discovery is achieved, the rate of technologic progress is limited by the often surprisingly slow rate of commercializing the concept into usable products. Whereas discovery and invention tend to be discrete events, technical innovation is often a painfully slow process. In spite of overall gradual change, technologic progress in most fields, when viewed in retrospect, tends to reveal certain relatively homogeneous periods where little change is evident, separated by episodes of accelerated progress. Thus, open-pit mining over the past century has changed from predominately train haulage to truck haulage and is now experiencing a major increase in belt haulage. Ore reserve estimation is another field that fits the same pattern, and the history of mine evaluation lends itself to a similar analysis. Evaluating mine investment opportunities is a fascinating aspect of mining. The unique set of risks inherent to every mining project is particularly intriguing. Geologic risks are most obvious, since faulty ore grade and tonnage estimates are among the leading causes of mine failures. Wide swine's in mineral commodity prices are usually cited as another major risk in mining. Because any ore deposit is a finite resource and the miner has limited discretion in selecting a mining sequence, mines are generally at the mercy of notoriously fickle commodity markets. There is a tendency for many mine investment analysts to feel that mineral market uncertainties are unprecedented in severity. Of course, prices in other markets rise and fall also-in some cases as precipitously as with minerals. However, two other mining characteristics tend to compound this marketing risk. First, the preproduction development period for a new mine is characteristically long. Under extremely favorable conditions, a significant mine may be developed in as little as two years, but the norm is four to six years or more. Second, mining is among the most capital intensive industries, consistently ranking near the top among all industrial sectors in assets per employee, and near the bottom in annual sales dollars per dollar of assets. Thus, mine investors typically have large sums invested for a long time before the outcome of the venture is determined by notoriously unpredictable markets. This is not a business for the faint of heart, and the importance of sound investment decision making is obvious. The high risk and technical complexity of mining operations have combined to require greater attention to financial evaluation by mining engineers than by most engineers in other specialties. Nearly all early mining engineering textbooks dealt with problems in mine valuation in a detailed manner; and, today, the sophistication of capital investment analysis is high in the mineral industries. Traditional Mine Valuation Methods Although some evaluation process has no doubt accompanied nearly every mine development throughout history, a formalized quantitative approach first emerged in the last half of the 19th century with the establishment of academic programs in mining engineering. The philosophy and approach to mine valuation that evolved at this time were, with periodic minor refinements, considered the standard for nearly a century. Only after 1960 did major changes begin to occur. The traditional approach to mine valuation reflected the investment community's historic attitudes toward risk. Mines were generally considered too risky for debt financing, and virtually every new mine was financed from internally generated funds or new equity subscriptions. The investment evaluation methods that evolved focused on mining's higher risk relative to other

Jan 11, 1982

By Clyde M. Pearce

This paper describes the facilities at Pointe Noire, Quebec that Wabush Mines utilizes for the receipt, distribution, stockpiling and shipping of iron ore concentrate and pellets. The main topics to be covered will be: 1. - General description 2. - Concentrate receiving facilities 3. - Conveying circuits 4. - Stockpiling 5. - Vessel loading GENERAL DESCRIPTION The map of the eastern Quebec - western Labrador area as shown in Appendix #1 shows the location of Wabush Mines properties. You will note the Scully Mine located at Wabush Lake, Labrador, 235 miles north of the town of Sept-Iles, Quebec. Our neighbors, the Iron Ore Company of Canada have their properties further north at Shefferville, near by at Carol Lake and at Sept-Iles. To the Southwest, our other neighbors, Quebec Cartier Mining, have their iron ore processing facilities and their terminus on the St-Lawrence at Port Cartier. Concentrates are prepared at the scully Mine by mining, crushing, grinding, concentrating, drying and high tension separation for shipment by rail to Pointe Noire, Quebec. Pointe Noire is situated across the bay from the town of Sept-Iles, Quebec on the North Shore of the St-Lawrence River. The concentrate consists of 66% iron, 2 ¼ percent silica and practically no moisture, permitting the use of specially constructed railroad cars and handling facilities to convert the concentrate into pellets, the final product.

Jan 1, 1967

By Lauren M. Roberts, Ralph R. Sacrison

Pipeline or borehole free-fall transfer of material has long been done in support of underground mining. This paper addresses some characteristics and design elements in selecting systems suitable to the transferred media and the mine requirements. Though aspects of overall systems are discussed, principal attention is directed to the slickline/dropline/transfer line and the discharge diverter/rockbox/energy dissipator/boot. These terms are common to these systems, often used interchangeably, and will be used as appropriate to the material, equipment and construction. Following several design sections, fabrication and construction, and system operating and maintenance procedures are discussed.

Feb 1, 2025

By A. J. Kroha, N. Wesis

Most copper mines produce both ore and low-grade "leach" rock or acid waters that contain recoverable copper. The sulfide ores pre¬dominate, and a portion that is too low grade for milling to produce concentrates for smelting, but has to be mined and trucked away anyhow, may be leached successfully with acid in dumps. Most of this leach material consists of sulfides and silicates or carbonates, and if the gangue is such that it consumes a high quantity of acid, this factor may rule out a leach operation. There are also valuable deposits that contain mostly acid-soluble copper, or occasional sulfide ores from which a sulfide concentrate can be roasted and acid-leached to produce a copper-bearing solution. Finally, there are milling ores in which the lesser part of the copper is acid-soluble and can be precipitated with iron or synthetic inorganic precipitants that produce metallic copper or copper sulfides that will float with the sulfides. Ordinarily, ores that contain copper associated with the sulfur ion, such as in the minerals chalcopyrite, chalcocite, bornite (and others), are milled to produce a 25-30% Cu concentrate for smelting, while a lesser amount of acid-soluble copper may be converted from solution to cement copper on iron scrap. A fast-growing percentage of such copper, however, is removed from solution with exchange resins or organic compounds in organic carriers such as kerosene (solvent extraction), then eluted with strong acid and subjected to electrolytic precipitation either in marketable form or as anodes that can be refined further. From the point of view of conventional copper smelting, copper flotation concentrates and cement copper are of chief interest in this chapter. Table I is a condensed open schedule for concentrates that generally run between 25 and 35% copper, and much less frequently as low as 12-15% or as high as 65-75% copper, the former being due to intimate relationship with pyrite (like the former United Verde Extension), and the latter representing such ores as the Bolivian Coro¬coro ore in which the copper is in the form of chalcocite in sandstone. These extremes are no longer common. When they occur, a special purchase schedule has to be negotiated. Included in Table 1, copper precipitates (cement copper) generally run from 70-85%a copper, and the same basic purchase schedule is used as with flotation concentrates. Sulfide Flotation Concentrates The sulfide copper concentrate produced in the mill as a flotation froth, with water then added for transportation of the heavy mineral particles from the flotation cells to thickeners, may run 60-80% water by weight, and the removal of water down to 25-50% by weight by means of thickeners, followed by further dewatering by continuous vacuum filters to 7-18% moisture by weight (depending on size of solids by screen analysis and also by content of clay) is a critical operation. Mill operators would like to produce a filter cake with 7-9% moisture content, but even with the help of steam on the filter this desirable condition is seldom realized when the concentrate is as fine as 80% -325 mesh. More commonly, the final concentrate is reground in pro- to produce best copper recovery and grade of concentrate (or molybdenite separation). In those cases, increasingly frequent, the filter product may not be a cake at all, but a mud that is hard to handle-even requiring a thermal dryer. Greater difficulty of form¬ing a manageable cake often comes from the copper-molybdenum separation by flotation, because the alkaline sulfides and hydrosulfides, or cyanide, or other similar reagents used for the separation, may leave the now relatively molybdenite-free copper concentrate even more difficult to filter. Handling a wet filter cake is difficult enough when its destination is only a short distance away-a matter of yards rather than miles. In those cases the filter cake may be thermally dried near the point of production, using rotary or multiple hearth, or fluidized-bed dryers. Alternatively, the concentrate may be pumped or carried in slurry form to the smelter and filtered there, or it may be spray-dried and compacted. For transportation to a smelter just a few miles to a few thousand miles away by ship or railroad other factors may be important, such as: in shipping by sea, avoidance of spontaneous combustion; in shipping by rail, losses by leakage if too wet or by wind and sun if too dry. It is the responsibility of the millman-usually the mill superinten¬dent-to make sure that his concentrates are in satisfactory condition when they leave the mill so that they meet these requirements: 1) They must have been accurately sampled and dry-weighed, the latter meaning that a moisture determination and gross weight must have been taken. 2) They must be dried sufficiently when necessary to prepare them for safe transportation. 3) They must arrive at the smelter with reasonable likelihood that they can be check-weighed and sampled fairly and equitably,

Jan 1, 1985

By D. S. Choi

The paper presents the result of analysis of subsidence data collected from coal mines operating in Northern West Virginia and Southwestern Pennsylvania. A particular emphasis la placed on the measurement of the angle of draw for the protection of surface structures.

Jan 1, 1986

By Christian F. Baiz

Industrial minerals can have similar financing requirements to those needs of metaliferrous and coal mining projects. Although, one may see many similarities in the fundamentals of the various risks that must be evaluated in lending to any specific project, one of the outstanding differences is the understanding of the many different markets that these industrial minerals represent. There are in excess of 30 principal industrial minerals ranging in terms of annual production from the largest quantity such as sand and gravel, to possibly the smallest quantity of production of a legitimate industrial mineral which may be defined as industrial and rough gem diamond. If we assume that it is necessary to finance independent projects of ore bodies of industrial minerals for development, the financing techniques are similar to the fundamentals that we would use for financing coal projects or metaliferrous deposits. Therefore, let's begin by redefining project financing in the traditional banking or private sector area, and then in the later development of this paper, we will apply these techniques to industrial minerals in general and cite it with several specific examples.

Jan 1, 1982

By Deborah Shields

Aggregate is an essential input to every economy, so governments need to ensure a sustainable supply of these resources through planning. There is also a need to shift from a linear economy to a more circular one, in which wastes are recycled rather than landfilled, so as to increase eco-efficiency. The purpose of the Sustainable Aggregates Planning in South East Europe (SNAP-SEE, SEE/D/0167/2.4/X) project is to develop a Toolbox for Aggregates Planning to assist governments and stakeholders in national, regional, primary and secondary, aggregates planning in southeast European countries. The Toolbox comprises documents on stakeholder consultation, data and analysis methodologies, a vision of best practices in aggregates planning and how to build an aggregates plan. The Toolbox will also support ongoing efforts to enhance minerals policy in the European Union (EU) as called for by the European Innovation Partnership on Raw Materials. Industrialized and developing nations see long-term and continued economic growth as essential to the well-being

By W. C. Streib

A wide variety of materials have satisfactory to superior insulating properties and can be used to control the flow of heat, provide energy savings, improve personal comfort and contribute to personal safety. Such materials range from completely organic products--both of natural origin and synthetics--to completely inorganic. In many cases, thermal protective systems are actually combinations of both. Obviously, those insulating materials which are principally inorganic have the inherent advantages of low to minimal combustibility as well as utility at vastly elevated temperatures. Both of these properties are highly desirable in providing not only thermal energy conservation, but also personal safety in residential and commercial applications. As the thermal insulation needs of man and industry have developed, there has been an ever-increasing demand on the minerals industry to provide the resources required to fill these needs. On a global scale, industrial minerals have become the fundamental base upon which today's thermal insulation industry has been built. The use of minerals to control heat flow range from the most unsophisticated applications to those requiring the ultimate in scientific technology. In many cases the minerals are themselves converted in- to the major insulating media; for example, fiber glass, rock wool or refractory fiber. In other cases, minerals are used as the chemical reactants to form the binder matrix which in turn provides the product strength required for adequate handling and installation; for example, calcium silicate bonded products wherein the binder is the reaction product of lime and silica. In still other cases, low density mineral products are added to a high density binder System to provide lower weight and better insulating properties; for example, the use of expanded perlite in concrete.

Jan 1, 1976

By J. P. Ramsey

This is a description of the efforts of a research project intended to demonstrate methods of alleviating pollution from coal mine refuse piles and slurry lagoons. These are mainly derelict barren sites. Some are burning and some are vulnerable to wind erosion which causes air pollution. Some occupy valuable land near metropolitan areas. They are an integral part of the landscape of the Appalachian and Midwestern Coal Fields. They are important in the light of growing civic awareness of the great general problem of waste disposal and environmental improvement. In the V-shaped valleys of the Pocahontas Region, they continue to be piled higher and higher. In flatter topography, they tend to form broad, flat ridges. Some contain more metal forms than others and some have a minimum of metal, but weather to a fine talc, which chokes streams with a fine particle suspension. Most are burning or have burned, but burning or weathering, most all are sources of pollution to the atmosphere and to adjacent waterways. This demonstration project is being carried out at an abandoned mine near the town of Du Quoin in southern Illinois. It is jointly sponsored by the Truax-Traer Division of Consolidation Coal Company and by the Federal Water Pollution Control Administration of the Department of the Interior. The site area is a refuse pile and slurry lagoon, each of fifty acres. It is hoped that basic data from this study will be generally applicable to other sites and will provide a basis for rational and efficient engineering design of future reclamation projects. Most refuse piles have extremely long term capabilities to produce highly potent acidogenic oxidation products typifying the more serious aspects of acid mine drainage contamination. Numerous sites remain where the abandoned mine no longer discharges acid waters, but its associated refuse pile continues to generate significant quantities of acid and other oxidation products even after many years.

Jan 1, 1969

By William Mayrsohn

INTRODUCTION The net smelter return (NSR) calculations made on various smelter-refiner alternatives, is usually done by taking the gross metal value less metal deductions, smelter-refiner charges and transportation-insurance costs. An important further step in the calculations is the financial charges imputed by refining time, transportation delays, to the time gold is returned (if tolled) or paid for if bought by the refiner. Metal that is priced in advance of payment also requires a forward premium that the mine may or may not be getting. This is a hidden cost that is often omitted from the calculations. Segmentation of each of the above costs are vital to see where improvements can be made through negotiations, better choice of metal pricing or transportation alternatives. BASIC NSR CALCULATIONS SMELTER-REFINER CHARGES Payable Metals - Treatment Refining Charges: Since these can vary from one smelter to another, the effect of a $25 per ounce price increase/decrease should be calculated and kept in mind. This effect is compared to differences in refining/treatment charge and a total smelter/refiner cost on a cost per ounce of dore, gold, or a dry ton of concentrate. Transportation - Insurance: Research has to be done on transportation alternatives perhaps with some assistance from the smelters, railroads or trucking companies. For dore, the U.S. postal service is viable and inexpensive. For railroads, deal with the regional office ie., B&N El Paso, if production warrants regular bulk shipments. Spot rates tend to be higher. A small mine producing concentrates usually goes to the nearest smelter due to extremely high transportation costs. Here is where a refiner can help. Insurance-Quotes for rates on dore/ concentrates are easily obtainable, but the mine has to shop for the best rates. Your insurance broker is the best source of this information. Differences between Mine and Smelter/Refiners Metal losses can be caused by the following: - Concentrate dry weight differences caused by moisture determination or faulty scales (often at the mine). - Sample bias by poor preparation practices on either side ie., grinding of the sample removing gold; or a poorly mixed concentrate. Dore bullion has no such difficulties.

Jan 1, 1987

By A. S. Winters

INTRODUCTION Homestake Mining Co. owns and operates the Bull¬dog Mountain mine near Creede, CO. Lead-silver con¬centrates are produced from ores extracted from narrow veins which dip from 1.04 to 1.39 rad (60 to 80°). An¬nual mill throughput is approximately 90 700 t (100,000 st). Ore shoots average 2 m (7 ft) in width and the ore is often weak, muddy, and poorly consolidated. Walls can be either strong or highly fractured along the strike of the veins. Cut-and-fill methods were initially selected to mine the deposit and considerable overhand mining occurred during the initial stages of the mine. Because of un¬stable ground conditions, mine management recognized that other methods would have to be employed. Undercut-and-fill mining was adapted to extract the loose ore from narrow veins and is presently the primary method used. Approximately 80% of the annual pro¬duction is extracted by this system. Loading ramps are positioned along the strike of the vein on 90-m (300-ft) intervals to allow for 45-m (150-ft) long cuts in each direction from the extraction raise. This interval was selected for efficient slusher operation and stope cycling. Loading ramps rather than chutes were selected for ore movement into 4.5-t (5-st) Granby-type cars because of extremely sticky ore conditions. Mined-out areas are filled with deslimed mill sands. Sand is pumped from the mill into two mine storage dams where the water is decanted. Each dam has a 544-t (600-st) capacity. Sand is pumped into the mine at a 30 to 40% slurry through 2195 m (7200 ft) of 51-mm (2-in.) thick wall pipe. When called for, sand is reslurried within the dams to a density of 55 to 65% and pumped to the particular mine opening requiring fill. Dry cement is added to the slurry prior to pumping. The capacity of the mine fill system is 20 t/h (22 stph). DEVELOPMENT Vertical development below the main adit level oc¬curs on 61-m (200-ft) intervals from an internal winze. Main haulage drifts or laterals are 2.7 x 2.7 m (9 x 9 ft) in section and driven parallel to the vein structures. Ground conditions generally dictate that better progress can be realized in the footwall than in the hanging wall. A distance of approximately 21 m (70 ft) is maintained between vein and drift openings to accommodate a smooth 18-m (60-ft) radius curve for crosscuts which are driven on 91-m (300-ft) intervals. Approximately 18 m (60 ft) of tail room is driven past the vein inter¬section to accommodate train loading. Ore and waste are hauled using diesel or electric driven locomotives operating over 18-kg (40-Ib) rail [762-mm (30-in.) gage]. Loading ramps are established above each crosscut¬ vein intersection. Scram drifts are then driven in the vein to determine the width and grade of the ore. If a prior decision to mine the block had been reached, then the scram is driven only far enough [about 12 m (40 ft)] to establish a drawpoint and scraper tail room. In most cases, the drawpoint is a 2.1 x 2.1-m (7 x 7-ft) cutout driven about 2.4 m (8 ft) into the footwall. On the level above, a raise bore station is established in the footwall to one side of the crosscut-vein inter¬section. The station is of sufficient size to accommo¬date a Robbins 41R raise boring machine. Support for the station is accomplished with 1.2 and 1.8-m (4 and 6-ft) roof bolts and wire screening. Once completed, a 152-mm (6-in.) concrete slab is poured and the raise drill set in position to center the 1.5-m (5-ft) diam hole approximately 1.8 m (6 ft) from the vein footwall. A 229-mm (9-in.) pilot hole is then drilled to the scram cutout below and reamed back to full diameter (see Fig. 1). STOPING Initial mining in a typical undercut-and-fill stope be¬gins from the top of the borehole. The initial cut can be either immediately beneath the track level or the old scram drift depending upon the particular situation. Cuts vary but are generally 2.4 to 3 m (8 to 10 ft) high. Support is accomplished by square-set timber, overhead caps, or roof bolts and screening depending upon ground conditions. Once the initial cut is removed, underhand timber and screen is installed and cemented sand fill placed. The stope is then ready for undercut mining. Prepara¬tion work begins with sinking 4.5 to 6 m (15 to 20 ft) on the borehole plus cutting a slot to the vein. As sinking progresses, hanging sets are installed to provide man¬way and service facilities. Hanging sets 1.7 m (5 ft 8 in.) long and 1.6 x 2.2 m (5 ft 4 in. x 7 ft 4 in.) in section are installed with steel hangers similar to shaft sets. Framed 203 x 203-mm (8 x 8-in.) timbers are made up in advance and shipped to the stope during sinking operations. When manway and service facilities are completed to the floor of cut to be mined, generally 3.65 to 4.3 m (12 to 14 ft) below the previous sand fill, the ore is removed to the hanging wall plus a round each way along the vein. Levelers or long stringers are then installed over the raise opening to support the work deck, slusher, and grizzly opening. All waste and ore from the sinking process is fed by gravity to the raise (see Fig. 2). Once the stope is equipped, mining pro¬gresses outward in each direction from the grizzly open¬ing. Stope equipment consists of 907 kg (2000 lb) pull air tuggers, timber skips, 11-kW (15-hp) ventilation fans, 15 to 22-kW (20 to 30-hp) electric slushers, 762 to 1270-mm (30 to 50-in.) scrapers, 67-mm (25/a-in. ) jackleg drills, as well as other drilling and blasting supplies. In narrow stopes [generally less than 2 m (7 ft) wide], no support is required for the underhand caps of the previous fill. Where wide sections in the vein occur, segment braces are installed beneath the caps to insure that the sand support cannot slip away from the hanging wall. Sections as great as 6 m (20 ft) wide and 4.5 m

Jan 1, 1982

By Ernest H. Gilmour

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

Jan 9, 1985

By Michael Drozd

Introduction Flotation is a recovery process that utilizes natural or chemically-introduced hydrophobicity of minerals to separate ore from gangue by attaching the ore particles to bubbles. Flotation has been known since 1906, when it was patented by E. L. Sulman, H. G. K. Pickard and John Ballot. Flotation's first documented plant use was in 1911, when it was used on gravity tailings from a zinc plant. Cyanide leaching was first patented in 1887. Most early precious metals flotation was linked to lead or copper concentrates. Most precious metal flotation is linked to sulfide recovery and utilizes such reagents as xanthates, dithiophosphates and thiocarbanilides. But the involvement of flotation in precious metals recovery does not have to be limited to active flotation of the gold and silver bearing minerals. Flotation can be used to remove interfering minerals, such as carbon, or cyanicides, such as copper, arsenic bearing minerals or other cyanide complexing minerals. Cyanide Leaching The use of cyanide to recover precious metals is the backbone of the industry. The various recovery processes that use cyanide are: 1. Leaching-Counter Current Decantation(CCD) Merrill-Crowe 2. Carbon-in-pulp (CIP) 3. Carbon-in-leach (CIL) 4. Vat leaching 5. Heap leaching These processes are selected depending on ore grade, ore leaching characteristics, reserve size and capital limitations. Ore grade is used as the first criteria to determine the type of cyanide plant required. However, the reserve size can determine whether a ground ore alternative is used. The ore leaching characteristics may indicate that the higher recovery, associated with a process that employs grinding, is the only economic recovery alternative. There are some extremely large leaching plants with grades that would be non-economic for some existing heap leach plants. Many factors, such as strip ratios, metal recovery and reserve size, determine what the ultimate plant design will be. The exact processing scheme that is used, is a multifaceted decision matrix, which gives the process metallurgist some job security. Precious Metal Recovery Using Flotation Flotation recovery can conceivably be used on any ore that has been ground. Many operations utilize gravity separation on the ore to remove native gold from the ore prior to sending it to cyanide leaching. On certain ores, where the gold is in visible nuggets or particles, this processing alternative is advantageous because the native gold cyanide recovery is

Jan 1, 1993

By Aarne Visnapuu

Unmodified pneumatic rock drills may have noise levels of 115 to 130 dB at the operator's position. Experimental modifications have been made to a standard drill to suppress the noise from the three main sources: the air exhaust noise, the noise radiated from the drill, and the resonance noise of the steel drill rods. A prototype model developed at the Bureau of Mines, Rolla Metallurgy Research Center, operates at a noise level below 100 dB. A close-fitting case and muffler unit, consisting of a metallic skeleton filled with viscoelastic absorber and having a durable cover, provided both muffling and noise absorption. Damping alloys were used at strategic points in the drill. Drill steel resonance noise was reduced by a constrained-layer treatment on the outside of the drill rod. The singular and combined effects of the treatments are presented in detail.

Jan 1, 1973

By James B. Meredith

The European Economic Community (The Common Market) is now shifting from I a coal oriented to an oil oriented economy. The proportion of energy supplied by oil has increased from only 13% in 1950 to 38% in 1962, and will approach 50% by 1970. In the past five years Common Market petroleum demand has in- creased 15% per year compared with only 3% per year in the United States. This rapid shift to oil can be explained by looking at oil's competition, First; there is simply not enough other energy - European coal, natural gas and hydro-electric power - to meet the demand, Secondly, oil?s primary competing energy in the Common Market is coal, and Community coal is high-priced - it costs more than two and a half times as much as U. S, coal at the pithead. The third reason for the shift to oil is consumer preference for the ease of handling, convenience and cleanliness of liquid fuels. As these reasons suggest, much of the shift to oil and the resulting growth in demand would have occurred whether or not the Common Market organization had existed. On the other hand, the Common Market has stimulated the use of oil by causing a higher rate of energy consumption, and oil has been the beneficiary of most of this growth.

Jan 1, 1963

By Michael P. Evans

Traditionally in coal processing jigs have been used to clean the coarse coal fraction, the +3/8" or +1/4" material. Today with the advent of modern fine coal jigs in the United States, its place was moved from the coarse to the inter- mediate coal circuit, the 3/8" x 100 mesh. The most economical coarse coal circuit to build and operate is a heavy media bath circuit. The following table summarizes the cost per raw coal ton processed for a coarse coal jig circuit, and a coarse coal heavy media bath circuit. The flowsheets used to compare these two circuits are shown in Figures 1 and 2.

Jan 1, 1983

By M. Deniz Turan, HASAN NIZAMOGLU, MÜCEVHER TURAN, RAMAZAN ORHAN

Tailings, or flotation waste, formed during copper production may be considered in the class of valuable wastes because of the amount of copper they contain. With the decreasing grade values of copper ores, the utilization of wastes like tailings that carry certain amounts of copper as secondary sources has become a necessity in both the economic and environmental senses. This study was carried out with the aim of using ammonium derivatives for the recovery of copper from tailings waste produced during copper production and utilizing their complex formation characteristics. In the conditions of six hours of leaching at 30 °C in the presence of 22.5 percent ammonium hydroxide and 100 g/L ammonium persulfate, 91.47 percent of the copper was transferred into the solution, while iron was not transferred, indicating that copper can be selectively recovered from tailings.

By R. Kalra

The current paper looks at evaluating the merits of large geared mills versus gearless driven mills by evaluating the following factors: The capability of currently available systems in terms of power and mill diameter Efficiency in terms of overall electrical power consumed vs mechanical output power Capital cost Running costs Cooling requirements in terms of air versus water cooling Installation and delivery time World?s largest geared drive ? 2x8500kW ? Wushan, China (operating) Among the many reasons that geared drives on large SAG mills are not favoured, is due to not having an efficient variable speed drive option (60 to 80% critical speed), which is required for process reasons. As technological advances have occurred quite rapidly in recent years, this paper uses the latest data in applying these comparisons and attempts to dispel some of the myths regarding Slip Energy Recovery (SER) and variable frequency drive technology on twin pinion drive systems. Early attempts to provide load sharing on twin pinion drives used primitive electronics which have since been developed to eliminate the problems encountered in earlier attempts. In particular, torque pulsations between the twin motors have been overcome using a higher number of devices (pulses) and better electronics. With successful operation of 2 x 6.5MW (Phu Kham, Laos) and three 2 x 7MW units under manufacture (Pascu Lama) and a further 40 projects worldwide, energy efficient Hyper Synchronous SER drives are now mature and ready for use on 2 X 10MW dual drive mills. Speed variation required is in the range of +/- 15% of synchronous motor speeds. VVVF drive packages from ABB and Siemens are also available for 2 x 10MW in low and high speed options. Dual hydraulic inching drives provide a much safer and convenient operation with the added advantage of locked charge detection.

Feb 27, 2013

By B. Huang, H. Sevim, R. Honaker

The use of medium-to-high sulfur coals has been negatively impacted by the 1990 Clean Air Act Amendment (CAAA). In this article, a method is demonstrated that can render these coals competitive and yet compliant with Phase II of CAAA. The method uses strategic mining and optimized preparation plant operation to minimize production costs, and it utilizes advanced coal-cleaning technology and an SO2 allowance to produce a compliant coal. The method is illustrated through an example from southern Illinois. For an annual production of 3.2 Mt (3.5 million st) and with a $150/st SO2 allowance cost, it was possible to produce a competitive coal for $0.84/million Btu. The production and allowance costs were $0.608/million Btu and $0.232/million Btu, respectively.

Jan 1, 2002