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By D. S. Ramey

In-situ leaching of the oxidized copper mineral resource at the San Manuel mine has proven to be a technically and economically viable method of copper extraction. Although this leaching method has been modeled to predict future production capabilities, the expected magnitude of the ultimate recovery from any portion of the present mine is not fully understood. A closely controlled test has been designed, constructed, and is now operating to determine the rate and magnitude of copper recovery. The test, initiated in June 1992, is scheduled to operate for approximately 18 months. This paper presents an overview of the test design and current results.

Jan 1, 1993

Ion Exchange is finding a greater role in copper processing, and hydrometallurgy in general, both as a technique for removing impurities and as a means of adding value through the recovery of secondary metals. Both the gold and uranium industries use resin adsorption as a primary production technique. While hydrometallurgical processes predominantly rely on solvent extraction as the mainstay of copper recovery there are a number of projects, in particular polymetallic deposits, which are considering the use of resin recovery as an alternative or supplemental production technique. A number of applications have been developed and proven at commercial scale for the removal/control of impurities (iron, antimony and bismuth) from copper electrolytes. These are well documented ion exchange applications. Increasing metal values and scarcity of certain metals are focusing companies to look at secondary metals recovery rather than losing these often highly valuable metals to tails residues. Bleed streams are being seen as valuable opportunities to recover contained metals; iron, copper, nickel, cobalt, bismuth, antimony and rhenium are all metals being targeted for removal or recovery from process solutions, electrolytes and bleed streams. An overview of currently applied techniques as well as some developmental technologies for impurity removal and secondary metals recovery is presented to provide insight into the advantages and flexibility of ion exchange applications in metals processing.

Jan 1, 2007

By KNona Liddell

Many metal recovery processes Involve heterogeneous reactions In porous media accompanied by fluid flow and mass transfer. Modeling such systems Is difficult because of the number of rate processes Involved and is complicated even further by effects due to the structure of the porous medium. Data on even bulk properties of the medium are often severely limited or unavailable. For many such systems, a cell model may provide, an attractive alternative to models based on transport ,mechanisms in a differential volume. A cell model requires less input data and is computationally efficient. Identical residence time distributions can be obtained with models of both types, and both allow conversion and concentration profiles and histories to be calculated. The formulation and properties of cell models for metals recovery are described in detail.

Jan 1, 1996

By L. O. Bond

High costs--and high risks--are normal in coal exploration and mine development. But risks and ultimate costs are minimized by foreknowledge of prospective coal seams and adjacent formations. Extensive drill hole programs are often used to obtain needed information. Cores and cuttings from these holes are used to determine coal location, thickness, and quality; they are also used to study physical and geological characteristics of surrounding beds. Valuable as these core and cuttings studies are, their effectiveness is often diminished by loss of recovery or by incomplete analysis. Needed is an effective and economical method for augmenting these studies. Well logs provide information of the type needed for efficient cool prospecting. Furthermore, obtaining these logs in drill holes, and evaluating the data they provide, is both fast and economical. To investigate the usefulness of well logging methods in coal explorations, the authors worked with Mr. Jon Jenkins of Truax Truer Coal Company. Logs were run in a number of drill holes in various areas of the Illinois Basin; these areas provided variety both in coal quality and geologic conditions. The data were studied extensively using computer methods similar to those now used in the evaluation of oil and gas wells. (2, 3) This study was used to produce a computer-generated analysis designed especially for mining problems.

Jan 1, 1969

By Gary P. Sames

This report discusses some practical applications of a geotechnical investigation conducted by the U.S. Bureau of Mines that can help mine operators meet revised Federal regulations in their roof control plans. The investigation, designed to characterize roof conditions near the coalbed outcrop in drift coal mines in eastern Kentucky, revealed that weathered stress-relief joints near outcrop are crucial ground control factors in the region. The joints' origin and character were determined through underground mapping of many joints in coal mine roof and detailed observations and measurement of joint trends and physical characteristics in widely separated strip mine highwalls and roadcuts. This resulted in an understanding of stress-relief-joint patterns and the effect of various rock types on the intensity of weathering in the joints. That information is used in this report to show how stress-relief joints contribute to roof failure and how, through improved roof support and mine planning, safer roof support plans can be developed.

Jan 1, 1992

By R. J. Reuss

Mina Maria is a high-grade, underground copper mine located in the Cananea district of Sonora, Mexico. The ore body is a massive sulfide breccia located in mountainous terrain. A major highway and an electric transmission line pass directly over the near-surface ore body. The mine is operated by Minera Maria SA de CV, a joint venture company that is 51 % owned by Empresas Frisco SA de CV and 49% by Cominco Resources International Ltd. Empresas Frisco is the operating partner. Production began in October 1990. The mining method used is undercut and fill. A continuous miner and a single-boom, hydraulic jumbo mine the ore. Backfill is a weakly cemented sand-gravel mixture that is pneumatically placed. The ore is treated by conventional crushing, milling and flotation. The Maria deposit is located about 40 km (25 miles) south of the Sonora-Arizona border and 11km (7 miles) northwest the city of Cananea and the Cananea Mine. The mine portal, at an elevation of 1780m (5840 ft), is on the east side of an important drainage divide. The mill site is situated about 800 m (2625 ft) northwest of the mine on the northwest side of the drainage divide where the drainage is more favorable. Its elevation is 1860 m (6100 ft) (Fig. 1).

Jan 1, 1992

By C. T. Bischoff

Introduction Air conditioning is a term which has come into common use only during our generation. By definition, it means the positive and accurate control of temperature, humidity, movement, purity, and component constituents of the atmosphere within an enclosure. Claiming such results for Noranda's ventilation system, or for that of any other large mine, at present would be an idle boast. However, we are working toward that end with considerable success. Modern underground ventilation at Noranda dates back about six years. The later development of air conditioning is in its fourth season, but still in the experimental stage. Various ingenious methods have been devised to cool mines which, because of climatic or geological conditions or depth, are excessively warm. A few installations have been made to warm intake air during the winter months where this is necessary. However, the method currently in use at Noranda is unique in that it accomplishes both ends by one process. This method utilizes the freezing and melting of ice to equalize year-round temperatures. Mining Methods While a discussion of mining methods at Noranda is not proposed in this paper, a brief summary may be in order. The Horne mine of Noranda Mines, Limited, has been developed by six shafts, four of which are now active. The most recent, an internal shaft, extends to 6,000 feet below surface, but all of the ore mined to date has come from above the 3,000-foot level. Figure 1 shows a model of principal ore bodies. The area mined is relatively small in plan, considering its vertical extent and the tonnage per vertical foot. Most of the ore lies within an ellipse about 600 by 1,000 feet. While this has certain advantages, the mining of such massive orebodies has presented many problems.

Jan 1, 1947

By L. C. Lang

Effective and profitable mining begins with good blasting results. While the importance of sufficient energy to break the rock is recognized, more attention should be paid to the time factor in blasting. Fragmentation, looseness and the shape of the muckpile can be improved and energy levels lowered through the use of proper delays. The long delay is also the most important parameter in establishing stable walls along the final limits of the pit. The time factor in blasting is explained through the rock acceleration process and its practical application is demonstrated by case histories.

Jan 1, 1981

By Earl Cook

A high-energy economy based on fossil fuels is in fundamental conflict with an ideal of clean air, clean water, undisturbed scenery, and maintenance of fragile ecosystems. There is an inevitable trade-off between the material standard of living of a modern society and the conservation of its environment. During the past four years, some of the costs of environmental protection have been internalized in the energy-flow system of the United States. The cost of bituminous coal and fuel oil usable under new air-quality standards has doubled or more. The cost of mining coal underground has risen because of a 20-30 percent reduction in productivity resulting from new environmental health and safety legislation; the economic impact of proposed surface-mine regulation may be equally severe. The cost of producing electric power has been rising at a 12 percent annual rate because of environmental costs, and would be climbing still more steeply were it not for the continued forced subsidy of natural-gas consumers by natural-gas producers. The efficiency of the automobile, already a monument to the triumph of self- indulgence over economy, is being further decreased by addition of anti-pollution devices. Part of the recent sharp rise in agricultural costs stems from the fact that five energy units of fossil fuel go into U.S. agriculture for every energy unit of raw food that leaves the farm; further large injections of fossil-fuel energy occur in the transport, processing, and refrigeration of farm products.

Jan 1, 1974

By G. F. Zhang

Yunnan mineral is a kind of containing Ta-Nb rare earth deposit, the most potential economic valuable element in it is Sc, and raw mineral has 17.0 g/t Sc. Take the character of Sc into account, the key problem in producing scandium chloride molten salt by adaptive technology is how to man-ufacturer anhydrous scandium chloride molten salt. During the course above, the key problem is that preventing Scandium chloride from being hy-drolyzed into ScOCl. The ScOCl content of molten salt not only affect purifying scandium but al-so do harm to the production of Mg-Al alloy. In order to restrain the creation of ScOCl, add ammo-nium chloride into molten salt. As long as ScOCl is created, it would be chloridized. Thus, during the course of dehydrating anhydrous chloride molten salt this method prevents scandium from hydrolyzing. Accordingly, the key phase of this producing course is eliminate ammonium chloride and the surplus crystal water, and the most crucial of it is restrain scandium chloride from being hydrolyzed into ScOCl. This research analyzed the theory that ammonium chloride is trend to be hydrolyzed into ammo-nium chloride, and propose the research scheme of restraining this process from creating ScOCl. The new process of restraining the generating of ScOCl and dehydration for crystal water rare earth chloride is adopted: Inert gas temperature increases rapidly to a certain degree to eliminate ammonium chloride and crystal water. The experimental research were done such as : the temperature of eliminating ammonium chloride and crystal water, amount of ammonium chloride, time, temperature growth speed, influence of inert gas flow toward to scandium percent hydrolysis, whereby, we conclude the optimal process condition in this period: Temperature: 400°C, proportion of oxidized scandium and ammonium chloride is 1:1.5(Proportion of oxidized scandium and additional salt is 1:5.4), time is 120minute, temperature growth speed is 10?/min inert gas flow is 5L/min.

Jan 1, 2014

By O. C. Blade

This report on the properties of motor fuels sold through service stations in the United States was made in accordance with a cooperative agreement between the American Petroleum Institute and the Bureau of Mines of the United States Department of the Interior. By agreement with the American Petroleum Institute, identification of the items is confidential. .It presents analytical data for 4, 457 samples, representing the products of 92 companies. The samples were collected by companies during June, July, and August 1956, As in previous surveys, the gasolines covered by this survey include those from both large and small suppliers. The data were obtained by laboratories of various refiners, motor manufacturers, and chemical companies and submitted to the Bureau of Mines for compilation. A list of the National Motor Gasoline Survey reports published during the past 10 years is on page 5.

Jan 1, 1957

By C Fountain

Australian mining and engineering companies have increasingly been investing in the mining industry in Brazil. It is a country of enormous mineral and agricultural wealth, with a well-established mining sector. Some Brazilian mining companies and personnel have been journeying in the opposite direction. There are significant differences in the approaches to engineering between the Australian and Brazilian industries that can cause confusion for companies operating in the other country. These range from differing expectations during the various stages of feasibility studies to style of plant and contracting strategies during the construction phase of projects. Some of the differences are driven by tax regimes, labour costs, or different settlement patterns. Others are the result of different histories and pressures on the industries in the two nations.This paper describes some of the differences in approach found in the two countries and some of the reasons for them.

Aug 8, 2011

By Ritika Kundu, Moussa Wone, Amanda M. Morgan, Tatiana K. Proctor

The Piney Branch CSO 049 Storage Tunnel (PBST) will be the last major project for the DC Clean Rivers Program. The project includes construction of a 2,200-foot-long, 18-foot-internal-diameter storage tunnel; two connections to combined sewer infrastructure; and two shafts. Major challenges include excavation underneath an existing live outfall that must remain in service; excavation of the tunnel through several areas of low rock cover and mixed-face conditions; and negotiations with a variety of federal, regional, and local stakeholders to address project impacts. This paper describes the overall project scope; geotechnical conditions along the alignment; design and construction challenges; and construction procurement.

Jun 23, 2024

By J. Cech

Discussion is presented of a decision-making support method for the spacing of ore passes along a long vein orebody and the division of the orebody into convenient stopes for future mining operations. The method was tested on a Slovak siderite deposit and also on some hypothetical cases with the objective of obtaining more general results

Jun 18, 1905

By J. Szekely, G. P. Martins

A formulation is given for describing the rate of expansion of spherical bubbles rising in liquids the freeboard of which is evacuated. The computer solution of the resultant differential equations has shown that, for low freeboard pressures (less than about 5 mm Hg), on approaching the free surface the bubbles expand much less rapidly than predictable from equilibrium considerations. In other words, in this region the pressure inside the bubbles will be significantly larger than the static pressure in the liquid corresponding to the position of the bubble. These theoretical predictions were confirmed by experimental work, using two-dimensional air bubbles rising in mercury. The important consequence of these findings is that the expansion of gas bubbles in vacuum degassing operations will be a great deal less than expected from hydrostatic considerations. This would lead to a significant reduction in the available interfacial area and may explain the apparent poor efficiency of many vacuum degassing units. VACUUM degassing as a treatment for liquid steel has gained widespread popularity in recent years; the number of known installations exceeds several hundred at the present time.' Although much information is available on both the thermodynamics of the system and the overall performance of various industrial units, much less is known about the fundamental aspects of the process kinetics.2-4 The basic physical situation common to virtually all vacuum degassing operations is the interaction between gas bubbles (swarms of bubbles) and the surrounding molten metal, held in a container, the freeborad of which is at a low absolute pressure. Once formed (or introduced from an external source) the bubbles will ascend, due to the buoyancy forces, and, during this ascent, a significant increase in their volume will occur. This progressive increase in the bubble volume is due to two factors: a) the continuous reduction in the static pressure acting on the bubble during its rise; and b) mass transfer into the bubble from the surrounding molten steel. In a recent paper Richardson and Bradshaw developed equations5 for describing mass transfer into gas bubbles from molten metals at reduced pressures. However, in deriving these expressions it was assumed that the pressure inside the bubble was identical to the static pressure in the adjacent liquid. In other words, the volume of the bubble was considered to be in equilibrium with the pressure of the fluid adjacent to it. This assumption, thus their analysis presented, was thought to be reasonably accurate for most of the bubble's ascent; however, it was unlikely to be valid in the immediate vicinity of the free surface, held at a low pressure. It was pointed out in the discussion6 that the region close to the surface may be of considerable importance as both the driving force and the interfacial area available for mass transfer are at their highest value here. The ' 'anomalous" behavior of gas bubbles when approaching a free surface at low pressures was recently confirmed in a preliminary investigation by Szekely and Martins. ?1 Here high-speed motion photography was used to study air bubbles rising in a column of n-tetradecane with a freeboard pressure of 0.1 mm Hg. It was found that significant distortion of the bubbles occurred on approaching the free surface; furthermore, the expansion observed was much less than what one could expect from hydrostatic considerations, i.e., factor a previously discussed. It follows from the foregoing that a detailed study of these phenomena would be justified both from fundamental considerations and because of their potential relevance to technology. The purpose of the paper is to present a more realistic formulation for the expansion of a gas bubble approaching a low-pressure region, together with a comparison of the theoretical prediction with experimental measurement. An inert bubble will be considered in the first instance; it is thought that the understanding of the fluid mechanics is an essential first step toward the realistic formulation of the mass transfer process. This latter problem will be the subject of a subsequent publication. FORMULATION The Physical Model. Consider a spherical bubble, of initial radius Ro, rising in a fluid, having a density pL. Initially let the bubble be at a distance H from the free surface, and at a pressure Pgo, as illustrated in Fig. 1. Pgo, the initial pressure in the bubble, is given by the following equation: pgo = Po = Ptp +pLgH [ la] where Po is the pressure in the liquid corresponding to the initial position of the bubble and Ptp is the pressure at the free surface. The fluid pressure at

Jan 1, 1970

By Bruce Pinkston

This report presents a description of the Prototype Demonstration Vehicle developed under U.S. Bureau of Mines Contract No. JO 366003, "Improving Man Transit Safety". This vehicle was constructed to assess the feasibility of a set of equipment related guidelines which were drafted in Phase I of the same contract. State-of-the-Art improvements were embodied in such areas as collision survivability, occupant comfort, and multiple mode (on and off rail) operation. The electric battery-powered vehicle carries a twelve man section crew at speeds up to 12 mph. The description of the demonstration vehicle is presented in Volume I. Safety Guidelines are presented in Volume II.

Jan 1, 1981

By Lutz Markworth, Evren Ören

"The worldwide demand of ores and the depletion of the high-purity deposits make it necessary to recover ores in increasingly finer size fractions in the feed material. The flotation technology of more than 40 years old plants was not able to recover the ultra fine material. Those slurry tailings ponds now have the chance to be re-processed with PNEUFLOT® technology.The bubble sizes of different flotation technologies have a huge influence of the results especially regarding the recovery of the ultra fine particle sizes.PNEUFLOT® technology of HUMBOLDT WEDAG Coal & Minerals Technology GmbH in Germany was applied to ultra fine Phosphate, Copper and Platinum ores. The results will be presented in this paper.Conventional technology like agitator flotation has its limits in the flotation of fines due to turbulences in the cell. The position of the agitator influences the recovery of the fines. A deep impeller generates a higher froth zone. This helps to catch the medium fines. Due to the static head of the slurry inside the cell, compressed air is required for such cell designs.Column flotation is well known for a good selectivity without having in the first cell a good recovery. This leads to invest in several cleaning stages to reach an acceptable recovery. Further more, compressed air is required to generate the air bubbles to the flotation process, which are much bigger than with PNEUFLOT® technology applied.INTRODUCTIONHigh metal prices are actually the basis to invest in poor deposits or deposits with very fine intergrown ores which have to be milled down to ultra fine powders less than 60 micron. The liberation grade has to be determined by a mineralogical analysis. Polish sections have to be prepared and interpreted by using of a microscope. Figure 1 and figure 2 are showing polish sections of a platinum-nickel ore from Petchenga Nickel at Zapolyarni Mine in Russia. The head grades for rich ore are for copper between 0.05 – 0.15%, for nickel between 3 – 7% and for platinum group minerals approx. 0.15 oz/t. With an efficient flotation technology it can be upgraded to 20 oz/t."

Jan 1, 2009

By Leonard B. Gulbransen, John R. Lewis, W. Martin Fassell, J. Hugh Hamilton

T. E. Leontis (The Dow Chemical Co., Midland, Mich.)—This paper is of particular interest to me because of my own work with F. N. Rhines on the oxidation of magnesium and magnesium alloys a few years ago. The authors are to be complimented on their development of an accurate and reproducible technique for measuring ignition temperatures and on their comprehensive study of the many variables that affect the ignition temperature of magnesium. It is indeed gratifying to see that they have obtained a good correlation between ignition temperatures and the oxidation rates reported by us. The correlation is valid not only with composition within one alloy system but also between alloy systems; that is, alloying elements which effect the greatest increase in oxidation rate also produce the greatest decrease in ignition temperature. There are a few points upon which I would like to comment. In attempting to correlate ignition-temperature data, one must be sure that the same definition of this quantity is used by all investigators. It does not appear to me that such is the case in the authors' comparison of their data with the theoretically calculated values of Eyring and Zwolinski. The equation derived by these investigators defines the ignition temperature, To, as the temperature at the gadoxide interface, whereas the present authors use the metal temperature as the criterion for ignition. The contradiction in the effect of oxide-scale thickness on ignition temperature between the predictions of the Eyring-Zwolinski equation and the observations reported in this paper indicate that some variable has not been taken into consideration. Could that be the geometry and size of the specimen? There is a marked difference in the type of specimen used in this investigation and that used in our work which formed the basis of Eyring and Zwol-inski's theoretical treatment. Another factor which plays an important role in ignition is the vapor pressure or the rate of vaporization. Combustion can safely be assumed to take place in the vapor phase by the reaction between vaporized magnesium and oxygen. Thus, a more accurate theoretical analysis may be made on the basis of the rate of vaporization which may be the controlling rate of the process. The effect of a large number of alloying elements on the ignition temperature has been reported in this paper, but beryllium was not included. Practical experience dictates that beryllium markedly decreases the burning tendency of magnesium. I was wondering if the authors plan to study the effect of beryllium in their future work. The authors predict that concentrations of sulphur dioxide in the furnace atmosphere greater than 5.8 pct would be expected to increase the ignition temperature to values still higher than those they measured. I would like to mention that large concentrations of sulphur dioxide markedly increase the rate of combustion of magnesium once ignition has started. Although it has been shown in the paper that the ignition temperature of magnesium in oxygen increases with increasing sulphur dioxide content up to about 1 to 2 pct, in practice relatively low-melting commercial cast alloys (AZ63A and AZ92A) are being continuously heat treated at temperatures just below the melting point in air containing 0.5 to 0.75 pct SO*. In regard to the change in color of the oxide scale observed on magnesium and magnesium alloys just prior to ignition, I would like to mention that in our work alloying elements were found to color the usually white magnesium oxide even though ignition did not occur. For example, the oxide formed on Mg-A1 alloys was gray, increasing in intensity with aluminum content in the alloy. Finally, I might suggest that the authors indicate their source of the value of 0.8 g per cc for the density of MgO as it is formed on magnesium upon oxidation at elevated temperatures. W. M. Fassell, Jr. (authors' reply)—The comments by Dr. Leontis are very excellent ones and I will attempt to answer them in order. First, the problem of ignition of magnesium is a rather difficult one since many factors are involved. Concerning the comparison of the To in the Eyring-Zwolinski equation, eq 4, with the experimentally determined values, it will be noted that the calculated and experimental values of the ignition temperature in Table I are not self-consistent. In the case of the 1.78 pct A1-Mg alloy the calculated value is 49°C below the experimental value; for the 3.81 pct A1-Mg alloy, 122°C below the experimental value; for Mg with 5x10-' cm film, 19°C above the experimental value; for Mg with 2x10-I cm film, 28 °C below the experimental value. Thus, if it were merely a matter of difference of location of temperature measurement the calculated ignition temperature would always be below the experimental value, the difference being due to the thermal gradient through the oxide film. The possibility of a thermal gradient in the magnesium metal must be considered. From Carslaw and Jaeger,'Y t can be shown that the maximum temperature gradient that could exist between the oxide-metal interface and the center of the sample is of the order of O.Ol°C. The geometry and size of the specimen could certainly have some effect on the ignition temperature. The equation for ignition that has been proposed in reference 14 is of the following type containing terms to account for this and other factors: M dT AHv(T) =Cp--------------\-J(.T—TB) + ZAHl-M A dx where AH is the heat of reaction, v(T) is the velocity of the reaction at temperature, Cp is the heat capacity of sample, M is the mass of sample, A is the area of sample, t is time, J is the total coefficient of heat transfer outward from the reaction zone, TR is the temperature of the bath or furnace, and AH,, is the heat associated with any phase change involved. Prior to the instant of ignition, the vapor pressure of magnesium is of no special significance. After ignition, neither eq 4 nor the above equation is applicable. The actual combination of magnesium cannot safely be assumed to take place in the vapor phase. While experimental data is lacking to support a hypothesis that ignition does or does not occur in the vapor phase, some observation on the pressure ignition experiments may be of interest. At high oxygen pressures, once ignition has occurred, the reaction of magnesium with oxygen approaches near explosive violence, the entire sample being consumed in probably less than 1 sec. At atmospheric pressure it usually requires 15 to 20 sec. Thus it appears that the oxygen concentration becomes the rate determining factor. Further, if burning magnesium is observed through darkened glass (Lincoln Super-visibility Shade No. 12) the magnesium sample is very much hotter than the "smoke" and the outline of the sample is retained perfectly. No "flame" is visible above the metal. No work was done on Mg-Be alloys. We do, however, intend to study this problem in the near future.

Jan 1, 1952

By J. A. Herbst

Ball mill grinding circuits can be subject to upsets that arise from variable characteristics of the feed ore. These upsets are transients in critical circuit variables such as ball mill mass holdup, circulating load, sump level, cyclone efficiency and product particle size. They may occur frequently so that the operation is continuously fluctuating or they may be of such magnitude that some of the equipment is overloaded, resulting in shutdown, or an undesirable product particle size. The feed ore characteristics that influence ball mill circuit behavior in this manner include the grindability of the ore, the eventual transport characteristics of the ore in the mills and the feed ore size distribution. Control strategies for grinding circuits using conventional proportional-integral (PI) control have limitations to their effectiveness in controlling against feed ore disturbances. The first of these are the interactive nature of two or more control loops on a single circuit. Under some circumstances they begin to fight each other, increasing control error and settling time. PI control loop gain constants can be tuned to give best response for some set of conditions but once these change, as they would with a change in the feed ore, they are no longer the "best" ones (Shinsky, 1967). Due to these limitations, alternatives to conventional PI control strategies have been investigated for grinding as well as other mineral processing operations. These alternatives, all computer-based, include self tuning regulators, decoupling control and optimal control (Herbst and Bascur, 1983).

Jan 1, 1984

By Russell Frith

Probably the most expensive but least well understood roof falls in underground coal mining, are those that occur ahead of the powered supports in the centre of longwall face. Even small roof cavities result in the dilution of the ROM product and large cavities can stop the face for days if not weeks and the incurred cash flow losses can be catastrophic. Expensive and time-consuming remedial measures (e.g. PUR, cavity filling, removal of stone from the AFC by hand) are required and equipment wear and damage can be significant when the need to blast and convey stone is also considered. Through four major research projects within the Australian coal industry and numerous consulting assignments, the author has spent the best part of fifteen years attempting to identify and defme the various geotechnical and operational factors that determine whether a major roof cavity occurs ahead of the powered supports or a normal goaf fall occurs just behind them, the difference in operational outcome being enormous. The paper presents a condensed summary of the overall technical findings covering such issues as roof fall mechanisms, the measured influence of cover depth, periodic weighting of massive strata units and its quantification, the role of cut-throughs in chain pillars, face width and height, the role of the powered support and the influence (both positive and negative) of mine site operations. The big picture outcome is the view that high production longwall extraction is well-suited to certain geotechnical environments, is marginal in others and under certain conditions is fatally flawed and beyond the control of mine operators, although fortunately the latter category is a relatively rare situation. Various risk-based research findings in this regard will be given as a basis for future discussion.

Jan 1, 2005