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By Wm. J. McCracken

HISTORICAL REVIEW OF BOARD OPERATIONS The Ontario Worker's Compensation Board was established in law enacted by the legislature of the Province of Ontario in 1915. It was designed to pay insurance benefits to injured workers, and at the same time to protect employers from legal suit. It was based upon an enquiry system rather than an adversary system such as that used in the courts process. Initially, the system was designed to pay compensation benefits and subsequently, to pay for the cost of medical treatment and pensions for disability and disease resultant from the effects of traumatic injury. In 1947, the Act was changed to include industrial or occupational generated diseases, not specifically related to traumatology. Such occupational diseases were therefore accepted and benefits paid subsequent to that date. As will be discussed in several minutes, even today the vast preponderance of compensation claims with the Ontario Board continues to be related to the effects of trauma. HISTORICAL REVIEW OF EXPOSURE TO RADON GAS DECAY PRODUCTS In some areas of Ontario, especially in Northern Ontario, there is a natural leaching of radon gas from the underlying rock formation. This constitutes very low levels of radon gas decay product radiation exposure to those persons coming in contact and inhaling these substances. This paper however is designed to discuss the occupational generated types of radon gas exposures. For many years dating back to the 1930's, partially refined ores were being shipped from Northern Canada to a refinery located at Port Hope, Ontario, still in operation and currently operated by Eldorado Nuclear Limited of Canada. Initially, the purpose for the operation was extraction of radium to be sold on world markets for medical treatment purposes. With the advent of World War II, this market collapsed. Subsequent to World War II, the availability of other sources of radiation for medical radio-therapy generally replaced the requirements for radium. During World War II, a new market opened up for the Port Hope refinery however as work into nuclear chain reactions and the development of the atomic bomb identified the need for uranium and enriched uranium. During the period of operations where radium was being extracted at the Port Hope refinery, it is now known that an identifiable radon gas hazard did exist. This hazard disappeared when the production line for extraction of radium ceased operations. In 1954, uranium mining operations opened up in Ontario at two locations, Bancroft and Elliot Lake. At the peak of operations, 16 mines were operational and 11,000 workers were employed in these mining operations. A high level of mining activity continued over a 10 year interval with the Bancroft Mines closing permanently in 1964 following a 10 year life of operation. The other mines in Elliot Lake closed about the same time with the exception of two uranium mine operations which have continued to be operational up to the present time. By 1965, due to a dramatic drop in world demand for uranium, the total work force had fallen to 1/10 of the peak work force, and approximately 1,300 workers remained in employment. It is of interest to note that one significant difference in the work environment between Elliot Lake and Bancroft was the high silica content of the Elliot Lake ore. This gave rise to a number of cases of silicosis developing in relatively short intervals of time in the Elliot Lake miner population. No cases of silicosis were identified from the Bancroft operations. Based upon the experience in investigating and evaluating actual cases of lung cancer in the uranium miners over the years, the medical staff at the Ontario Board also developed the impression that radiation levels were much higher in the Bancroft operations, especially in the earlier years of operation, than at Elliot Lake. This resulted in accumulation of higher levels of Working Level Months (WLM), usually over a shorter exposure interval in many of the cases. This aspect will be further evaluated in this presentation. Subsequent to 1965, the work force remained quite static in numbers until 1975. At that time, there began to develop an increase in the work force, and this increase is continuing at a moderate rate up to the present. INITIAL METHOD OF HANDLING LUNG CANCER CLAIMS The first lung cancer claims in Ontario from uranium mining operations were accepted on the perceived cause-effect relationship. This relationship was based upon the data from the Colorado observations and the Czechoslovakia data. Initially, a series of regression equations on mortality were developed and used to estimate the effect of exposure to low cumulative doses of radon daughters as it might relate to the frequency of occurrence of lung cancer at any particular cumulative exposure level. A probability of cancer being radiation induced as against it being caused from other factors was calculated. This method was discontinued subsequent to 1972 due to problems encountered in carrying out this complex evaluation. Thereafter, each case was dealt with on an individual basis, being based upon whether or not the tumour was of the oat cell type, a cumulative exposure in excess of 120 WLM; latency periods in excess of 10 years, commencement of mining prior to

Jan 1, 1981

By J. W. Jang

The wear behaviors of grinding media with different chemical compositions and microstructures in wet grinding of a quartz-pyrrhotite mixture was investigated under various environments. The marked ball wear tests showed that the wear rate could be minimized by a suitable combination of martensite and ferrite within the same ball. A ball composed solely of hard martensite exhibited increased wear under an oxygen environment, while balls composed of both soft ferrite together with hard martensite exhibited decreased wear due presumably to the difference in their passivation? behaviors. The galvanic interactions between the two phases as well as among these phases and an electroactive mineral were investigated through electrochemical measurements under abrasive conditions in a slurry. The rates of formation of passive films controlled the corrosion currents of the high chromium steel media. Corrosive wear and its effect on flotation are discussed using the passivation kinetics model thereby developed.

Jan 1, 1989

By J. C. Yingling, J. W. Leonard

Introduction The absence of coal-face particle management heavily penalizes the transportation of coal from initial loading to final consumption. The penalties include dust problems, significantly reduced mine-loading-cycle productivity, mine-belt spillage, excessively high coal-preparation costs, chute blockages and dangerous pulverizer blockages at the final point of utilization. Fine particles commonly cause environmental and economic problems. It is well known that these fines can cause safety and environmental dust problems. But it is not well understood that these fines can also swell broken coal to a point where 5% to 15% more time and capacity must be used to deliver the same tonnage. In this paper, methods and rewards for reducing and/or managing fines at the mine face are discussed. Computer-based loading-cycle model productivity estimates, viewed from a new perspective, are made on the basis of material volume rather than on the long-established, and frequently misleading, basis of tonnage. It is typically the volume of broken material being transported that defines the capacity of a given transportation system, while the corresponding tonnages are merely a reflection of the specific material densities. Published evidence suggests that the swelling of broken coal can be decreased very significantly using small quantities of certain nonfrothing chemicals, which are added to mine-face spray water, and by employing improved mine-face breakage practices. In a future paper, the effects on transportation productivity beyond the coal mine will be discussed. The precursor to the work presented in this paper, involving the bulk density improvement for broken coal and the subsequent production gains for underground coal mines, was earlier presented in Leonard and Newman (1989). In the past, this topic has been studied and practiced only in byproduct coking in the steel industry. However, a potential exists for an increase in coal-industry productivity by improving the bulk density of coal to yield a subsequent reduction in delivered cost. This can occur with breakage, handling and treatment methods resulting in the loading of greater quantities of coal in fixed volumetric capacity haulage units such as mine cars, shuttle cars and scoops. Laboratory-based experiments to achieve an increase in productivity by increasing coal bulk density were discussed in Leonard, Paradkar and Groppo (1992). Chemical techniques using small quantities of commercially available reagents (surfactants) resulted in about a 13% to 15 % increase in bulk density, which was thought to produce a proportional increase in the productivity of a mine, together with a subsequent reduction in cost. The idea is to mix the reagents with the water that is used to spray coal during mining. In this paper, the impact of bulk density improvements on production rates is presented. Increases in production ranging from 60% to 88% of the bulk density increases are projected. This analysis was performed for atypical continuous-miner section. In the following sections, discussion and results of the analysis are presented. Discussion An analysis was performed to ascertain the impact of bulk density improvements on face-production rates for a typical continuous-miner section. Figure 1 illustrates the section layout and cut sequence. This layout and sequence is identical to the case described in King and Suboleski (1991). As can be seen, the section uses five entries and 12.2-m cuts that are taken by a remotely controlled continuous miner. The seam height is 1.5 m and two shuttle cars (5.7 t nominal capacity) are employed for haulage from the miner to the section feeder, which, throughout the cut sequence, is positioned as illustrated in Fig. 1. The simulation model was coded in the SIMAN simulation language. The major impacts of increased bulk density improvements on such a production system are as follows: •Shuttle-car payloads, in terms of the mass of coal transported per haul cycle, are increased proportionally to the increase in bulk density that results from the application of surfactant. •Shuttle-car discharge times should remain largely unchanged, because they are determined by the volume of material that is discharged, rather than the mass, and this volume does not change.

Jan 1, 1996

By Michael Michaylov

INTRODUCTUCTICN The coal seams at Bobov dol basin are highly prone to spontaneous combustion in Bulgaria. The risks of endogenous fires (SPONCOM) exist mainly in gob area. During the last 15 years three different technologies for preventing and fighting SPONCOM have been applied in underground mines of this basin: pulp from hard material and water, foam and inerting with nitrogen. On the basis of experimental application of these technologies as well as theoretical and experimental research a new technology - Foam Pulp Technology (FPT) was worked out. It is very effective in difficult to reach places in a gob. BACKGROUND AND THE NEED OF FPT Application of the three methods mentioned above showed advantages as well as some disadvantages, allowing determination of their effective application scope depending on geological and working conditions. Their advantages can be summarized as follows: - Foaming: high viscosity at low density, which allows forming a thick layer; good surfactant wetting properties; excellent effective cooling during evaporation in high temperature zones. - Pulping with mixture of solid particles (e.g., fly ash) and water from surface installation: easy to apply in a mine; low operational and energy costs due to gravity transport from the surface to the gob: good ecological results because the waste material is returned back underground. - Inerting with nitrogen: volumetric action which doesn't restrict the thickness of treated coal pile; easy to apply in a mine; controllable nitrogen delivery depending on fire hazard conditions. Main disadvantages of the same three methods are: - Foaming: rapid disintegration due to intrusion into heterogeneous and porous media and high costs due to the need of frequent large volume foam flows (short foam's life); - Pulping: inapplicable for coal piles with high thickness which is of great importance in the cases of seams with ununiform hypsometry (rapid and frequent changes in a seam configuration) and/or great volumes of coal or ore piles particularly heated to a high temperature; strong dependence on gob floor declination; water draining of the working face area which impedes the movement of supports and chain conveyor's. - Inerting: relatively high costs hardly depending on the rate of face advance (fig.1) in comparison with the other methods; it is difficult to ensure sufficient oxygen at the exit end of the face. This problem is crucial in faces with a high level of methane emissions; accelerating of the oxidation and spontaneous heating of the pile after turning off the nitrogen supply. Inerting should be applied again at higher capacity of nitrogen supply (fig.2) and what is more important, it shouldn't be

Jan 1, 1995

By Charles E. Barker

Striking similarities exist between the sediment features and facies formed in some modern heliothermal lakes and the characteristics of some ancient borate depositional environments. Heliothermal lakes are density-stratified with shallow submerged margins surrounding areally restricted deep pool(s) containing a dense brine overlain by a much less dense brine. The reflective brine interface allows solar energy to be trapped in the dense brine which may warn to over 900C. carbonate precipitated from the dense brine is the typical sediment produced in warm deep pool. The strong density stratification reduces turbulence in the lake so clastic sediment is deposited under most of the lake as finely-stratified mudstone-limestone couplets (varves). This depositional process results in clastic lake sediment interfingering with and enclosing a thick evaporite pod(s). Because deep pools can exist at the same time or through geologic time, the evaporite pod(s) may occur sporadically throughout the lacustrine beds.

Jan 1, 1988

By Betty L. Gibbs, Jean A. Arcamone

Introduction Microcomputers are becoming more convenient to use at all levels in the business world, including the mining industry. The state of the mining industry requires that we become more productive and innovative in our analysis of deposits and operations. With the help of computers, examination of options to increase productivity can be performed quickly and inexpensively. Effective computer use to design underground workings combines the calculating power and speed of a computer with the technical expertise and judgment of an engineer. Computer programs can perform in minutes calculations that might take several months for an engineer. But the engineer must be on guard when using computers for calculations. It is easy to assume the results produced by a computer program are always correct. However, the computer is only a sophisticated calculator. It cannot recognize logical errors or transposition of numbers. Also, some intuitive or obvious decisions made by an engineer are difficult or impossible to program effectively. In this article, selected commercial programs are discussed. These programs can provide engineers with tools to evaluate engineering designs for underground mines. The programs perform much of the tedious calculation. The engineer, though, must still make decisions and conceptualize designs. The software mentioned in this article runs on microcomputers and minicomputers. The mention of particular software, however, does not imply endorsement. Mining projects After a decision has been made to proceed with an underground mining project, more detailed evaluation of the mining method and extraction sequence begins. If a mine is already in operation, schedules and plans are needed to more clearly define future production. Existing operations can benefit from identification of bottlenecks in material movement and produce extraction options to make the mine more efficient. Detailed studies for a new mine are based on the technical engineering design. These studies provide specific details to determine equipment requirements. Detailed studies done during a project are similar to short-term planning and control carried out daily by the mining engineers. The engineering design phase includes: • Study of the infrastructures - entries, shaft, access from surface to underground; • Definition of the mine layout, hydrogeology, geotechnical considerations, transport, and ventilation; • Choice of machines and equipment; • Mine scheduling, including waste extraction and production sequencing; and • Economics, including labor and supplies. Further detailed studies include: • Blasting patterns; • Detailed geotechnical calculations; • Secondary ventilation; and • Production cycles along with coordination of the workings. Objectives The primary objective of underground mine planning is moving material out of the mine at the least cost. Development of the production schedule considers location and volume of ore units, transport distance, equipment and manpower interactions, and grade or quality of material. Two methods can be used to determine the scheduled material removal: optimization or trial and error. The microcomputer programs discussed here do not cover optimization for underground mine planning. For several reasons, optimization is often a "black box" approach. Although optimization is possible with linear programming techniques, few mining operations can be described effectively with linear equations. An interactive trial-and-error approach is desirable because engineering judgment can be inserted as a plan develops. Because of the large number of factors in an underground operation, an optimization technique seldom covers all possibilities. Few of the programs available consider all of the possible parameters. The programs use an interactive sequencing or a simulation approach and are either event-oriented or deposit-oriented. The most desirable programs offer an integrated approach. They consider a broad range of operating conditions including equipment, labor, costs, material movement, timing, and mine layout, among other factors. The two types of models generally used in designing underground workings include: • The interactive model, which requires direct, incremental input from the user. Outlines of the section to be mined may show on the screen. The engineer causes the model to develop piece-by-piece. The program may offer feedback on parameters such as tons and grade, or other values, as the section is "mined." • The simulation model, which is usually an automatic technique. Parameters related to the process are entered. The program then calculates the mining sequence over a designated time period or until the economic material is exhausted. A simulation model imitates step-by-step performance of equipment or other processes.

Jan 3, 1988

Introduction Surface mining is a business, and the objective of most businesses is to make as much money as possible within certain responsible constraints. Businesses normally measure their potential to make money by estimating the discounted cash flow rate of return on new projects or the net present value for ongoing projects. These estimators of profitability are calculated using the following well-known equation: [ ] where i is the year, n is the total number of years, A, is the after tax net cash generation in year I, r is the discount rate and NPV is the net present value. In its simplest form, when the value of r makes the NPV equal to zero, the discounted cash flow-rate of return (DCF-ROR) is defined. The maximum economically recoverable reserve is identified at the point where the NPV or DCF-ROR is at a maximum. Depending on the measure used (DCF-ROR or discount rate) for calculating the NPV, the definition of recoverable reserve will be different for exactly the same deposit. Maximization of the measure of profitability and the economically recoverable reserves should not be separated from the mining method and sequence of exploiting the reserve. The surface mining methods may involve multiple-pass development of the resource, mining to the ultimate limit in one pass or a combination of the two methods with backfill. Multiple-pass development When planning a multiple-pass mining operation, it is important to determine the width of pit expansions that are appropriate for a specific situation. Production personnel typically want wide, open work areas where operations can be set up to run at peak efficiency. While this is understandable, the trade-off between highly efficient production operations and an overinvestment in stripping should be carefully weighed. As long as safety is not compromised, a few cents per bank cubic yard increase in operating cost due to lower productivity by working in more confined spaces may well be offset by a decrease in the overall short-term stripping requirements. A mini study on productivity and cost should help quantify this cost before deciding on minimum phasing requirements. The bigger challenge associated with establishing smaller work areas is securing operations and management buy-in to marginally lower productivity when reporting results. When cost reduction is the important management goal and everyone is focusing on unit costs, one sometimes loses sight of the goal of total cost management and, even more importantly, maximum profitability. The first pit should be defined by the first grouping of recoverable mineral that has the highest after-tax unit value and forms a feasible mining unit. The assumed mining of the first pit should be followed by the assumed mining of the second most profitable practical pit or pit expansion. If this process of grouping and sequencing is followed until the profitability approaches zero, the general sequence required to extract the maximum economically recoverable reserve is identified. The curve in Fig. 1 is constructed for a rate of production, R, and a quality management strategy, Q, by calculating the NPV for a series of sequences, with each sequence assuming that the operations are terminated on successively lower profit pits. These values are plotted to form the curve in Fig. 1. If the operation ceases on a pit that has too high of a profit, opportunity is lost, and the NPV is lower than the maximum. The NPV curve peaks and starts to decline before the zero-profit pit is assumed mined. This occurs because the money invested in advanced stripping would have made more money invested at the discount rate than by mining the next pit expansion. Therefore, the maximum economically recoverable reserve is defined by the pit limit corresponding to the maximum NPV. There may be an opportunity to refine the definition of the "ultimate pit" around the maximum-value point. The precise definition of the ultimate pit early in the mine life may or may not be important today, depending on how far off in the future these limits will be reached. By the time an operation must decide on its last increment of economically recoverable reserve. much of the

Jan 1, 2000

By S. Polegeg

The basic principle of this separation method was already defined in the middle of the nineteenth century. However, a change over for the treatment of minerals has followed from the construction of proper machines i.e. airjigs and airtables as late as in the twenties and thirties of this century. The main application of this technology was subsequently found in coal processing and reached quite a great dimension in the two following decades. The number of applications of this method in mineral dressing plants decreased rapidly and has almost vanished in the fifties in favour of wet concentration procedures. It has been for a long period as long as up to the eighties that no more consideration was taken for applying dry gravity separation in the field of mineral processing also due to the successful introduction of flotation. Due to the fact that labour cost was increasing remarkably within the past period after the end of World War II also the economic conditions in the agricultural sector became more though. As a result there came up a great demand for harvesting machines. Since agricultural products can only be processed by dry methods in a prevailing number, highly developed machines were introduced for the dry processing of food and feed stuff.

Jan 1, 1992

By William F. Riggs

The basic theme of this symposium and panel Is Rotation Pads: Are They Optimized? There Is a. reason for phrasing the title In the form of a question. There Is not only the technical competency which we must address; there Is the operating philosophy that must be evaluated on the part of both the customer and the supplier. Customers desire reagents which are trouble-free and capable of providing that extra amount of selectivity or recovery. When they receive ft, after the supplier has provided several years of Internal research, one of the first concerns/complaints Is the price of the product. This has a tendency to rapidly reduce a supplier's support level In the future. Suppliers are equally guilty from another perspective. When they approach a customer to Introduce a product, they often attempt to market by offering only a price Incentive. They then wonder why a customer doesn't respond Immediately to the incentive. They are often oblivious to the fact that the reagent cost is such a minor aspect of the operating budget, and the customer has many more pressing problems on a day-to-day basis In comparison to the reagent cost. We need to establish the understanding that reagent cost Is an Inconsequential cost of operation, and yet has such a disproportionately high Impact on the success of the entire operation. This understanding Is required by both the customer and the supplier. We say to each other,' why are we discussing this since this has been obvious for some time?' The reason is relatively simple in that we talk about it, acknowledge it, and yet we do not adhere to it. The supplier provides a product along with test data containing statistics, analysis, recovery, grade and cost calculations while most of the time ignoring the operating technique which must be applicable In the plant In order to optimize the product. He expects the reagent to be substituted In the plant for the existing reagent and ft works or does not work after trying several variables. The operating management Is equally guilty, In order to best explain this to both the customer and the supplier, ft becomes necessary to review the basic purpose of the major reagents utilized In flotation. A collector is basically to Impact selective, maximum water repellency on the surface of a particular mineral, The frother has the purpose of providing a chemically stabilized membrane on the surface of the bubble at the air-water interphase. This, then, provides a host environment for the attachment of the collector-coated mineral to a bubble. The depressant functions In the reverse of the collector and must demonstrate the same or greater degree of selectivity than expected of a collector. The key area which has been Ignored Is the rate by which these reactions occur and Interrelate. This has a very specific effect on the operating technique and the compatibility of the chemistry, equipment, and the operator himself. Researchers, suppliers, and customers provide reams of data to demonstrate how their products or design produce, for example, higher kinetics, more selectivity, or more recovery. The Information is often true. After all, we are all learned men and laboratory and actual plant data do not lie. However, we must remember the theme of this symposium and panel: Flotation Plants: Are They Optimized? and Optimizing of Flotation Reagents? The direct, honest comment to the two titles is very simple. OF COURSE THEY ARE NOT The plants, equipment, and reagents had better not be optimized or else we are in trouble. The Issue of this panel discussion is to approach this subject from a slightly different or perhaps mainly Ignored aspects of optimizing reagents in flotation. When we have reagents which provide higher kinetics, more selectivity, and better recovery, how do we use them? Since each reagent has a different physical characteristics of froth, rate of recovery, volume effect on the compatibility of equipment, and many more aspects too numerous to mention, the question which has been severely Ignored Is, 'What degree of study and cooperation by both the supplier and the operating management has been conducted In order to prepare the operator for maximizing the performance of a reagent In relation to the rest of the system?" Prior to testing a new reagent, how much time Is spent to bring the actual operator(s) Into the program to make them feel part of the program? How much time is spent explaining to the operator on the float floor how to possibly take advantage of a reagent with faster kinetics or one which Is Inherently more selective? What

Jan 1, 1993

By L. M. Cenegy

Introduction The formation of scale is common in gold ore processing plants. A working definition of such scale is: "A hard, tightly adhering deposit formed in place by the precipitation of calcium-bearing mineral compounds from water." The most common scale found in gold processing plants is calcium carbonate (CaCO3) although calcium sulfate (CaSO4) is occasionally a problem. The problem of scale in carbon strip circuits Carbon strip circuits present a particularly challenging scale inhibition problem. Many of the factors that tend to promote scaling are present in these systems. For example, carbon strip circuits are run at high temperatures and high pH levels and require long batch cycle times. Tortuous paths are encountered by the strip solution circulating through heat exchangers, pipes and elbows, causing transient pressure drops throughout the system. Severe conditions such as these cause the water to possess a high scaling potential and may lead to the breakdown of some of common types of scale inhibitors. Formation of scale Natural waters contain dissolved minerals and gases that promote scaling. One of the contained gases, carbon dioxide (CO2), is an important factor in scale deposition. Carbon dioxide dissolves in water to yield carbonic acid according to the following reaction: [CO2 + H2O - H2CO3 (carbonic acid)] Carbonic acid rapidly dissociates to yield bicarbonate and carbonate ions depending on the pH of the system: [H2CO3 -s H+ + HCO-3 (pH range 6-10) HCO3 -a H+ + COO (pH range > 9)] In gold cyanidation processes, lime is added to raise the pH of the mill solution to 10 or above to promote cyanidation of the gold and to eliminate the possibility of hydrogen cyanide (HCN) formation. Lime contributes both hydroxyl ions (OH--) and calcium ions (Ca++) to the solution. The normal scale forming reactions in mill water systems were described by Beasley (1973) as: [HCO-3 + OH -s C03 + H2O Ca++ + C03 -a CaCO3] Factor affecting scale deposition According to Linke's (1958) solubility tables, the solubility of calcium carbonate in carbon-dioxide free water is only 13 mg/dm3 (13 mg/L) at 25° C (77° F). Scale deposition occurs when the concentration of calcium and carbonate ions in the solution exceeds the solubility of calcium carbonate. This occurs, in most cases, when the water has undergone some chemical or physical change. Some typical changes that might occur that lead to scale deposition are discussed below. Temperature An increase in temperature can greatly increase the tendency to form calcium carbonate scale. Studies by Miller (1952) have indicated that calcium carbonate is less soluble at higher temperatures. Hence, high temperatures promote scaling. For this reason, scale is often noticed on heat exchangers, autoclaves and other high temperature surfaces. Pressure A decrease in pressure may lead to an increase in the tendency to form calcium carbonate scale. Fulford's (1967) work has shown that scale is more soluble at high pressures. Hence, scale is likely to be prevalent in pump suction lines, elbows, baffles and other areas that produce turbulence and subsequent pressure drops. Change in pH Calcium carbonate scale is pH dependent and tends to form more readily at high pH. Any event that causes an increase in pH, such as the introduction of lime or caustic, would tend to initiate scale formation. Since most gold recovery operations are run at a pH above 10, the tendency to form calcium carbonate scale is always great. The use of scale inhibitors to prevent calcium carbonate scale Many products are commercially available to inhibit the formation of calcium carbonate scale. Some of the more effective types include phosphonates, as well as acrylate and maleate-based polymers and copolymers. Many of these products function by means of threshold

Jan 1, 1992

By A. M. Edwards

In many ways a government could be regarded as a macro business enterprise. Its shareholders are its peoples, the productivity of the nation depends on them, and, indeed, the profits or losses of a nation's balance sheet can be linked to the surplus or deficit in its balance of payments. International loans need to be repaid, and in most developing nations the noose is tied so tightly that one wonders if such debts will ever be repaid. In their endeavour to meet such obligations, their currencies may be devalued [sometimes in frightening proportions], in the hope of boosting the export potential. However, this is a vicious circle, since exports will not be achievable unless sufficient technical expertise and infrastructure exist to address the opportunities which may have been created through devaluation. Many of the poorer countries have exportable resources of minerals, which are supplied on the international market. This may be the good news, but, unhappily, steps were seldom taken to enhance the value of the mineral products through further beneficiation and the production of high-value sophisticated products. That is where the vicious circle comes in: inadequate infrastructure is created, competent and well-trained manpower is not generated, and the result has inevitably been that the GDP slips backwards, and the nation becomes progressively impoverished. To exacerbate the problem, it is unfortunate that the more impoverished a nation becomes, the greater is its birthrate, and hence the greater the need to create wealth at the very time that the nation is becoming poorer and poorer. Within this scenario, one may well reflect that this is exactly what nations on the African Continent are experiencing, namely exploding populations, although most of the poorest nations in the world are found on this continent. Yet they often possess a superior mineral production potential. Africa still suffers from what may be called the colonial influence, in that the developed world has regarded this continent as a source of raw materials to feed its own factories. The colonial powers gave only secondary attention to developing the nation's peoples and its infrastructure, such as railways, roads, harbours, and other services. These problems all naturally fall within the ambit of government interest, and it is therefore obvious that the state should play a dominant role in addressing them. The reasons for governmental involvement in mineral-related research and development differ considerably, depending on the state of industrial development of that country. Thus, in the highly industrialised world, government's interest will lie mostly in the fields of strategic minerals, which may not be economically exploitable. An example would be the exploitation of American low-grade chromites during the Second World War, or minerals which need to be imported from unfriendly or unreliable sources, such that disruption would not be in the national interest. Colonial powers like the United Kingdom and France established R&D facilities to assist, inter alia, in the exploitation of mineral resources in their colonies. The approach of developing countries would differ from that of the industrialised nations, in that their primary quest would be to promote, support and encourage the development of infrastructure, and the establishment of major industries in order to create and provide employment opportunities, as well as to support small operators who cannot afford to undertake their own research and development. On the debit side, it is a sad fact that, however good their intentions may be, underdeveloped countries are invariably unable to provide the requisite financial support to sustain their own research-and-development laboratories. In a high-tech world, the cost of sophisticated equipment is steep and cannot be afforded. Again, the proverbial vicious circle comes into play. The competent researchers find themselves without rewarding employment and would therefore shun such institutions. Recent trends in the developed world have been to privatise research-and-development institutions, the argument being that they should draw closer to the industry they serve, and that the provision of taxpayers' money for strategic purposes is no longer as important as it may have been once. Complete state withdrawal has had its negative effect in that long-range, high-calibre research has tended to be replaced by short-term problem-solving projects. Industries engaged in the minerals business differ somewhat from major ones, like motor, computer or pharmaceutical industries, which are large enough to support their own internal research and development, and whose work is highly confidential, so that any leakage could have disastrous consequences. Furthermore, private companies are answerable to their shareholders and costly long-range, 'blue-sky' research would be difficult to justify. Since mineral research has become of lesser importance in the sophisticated First World, this trend has persisted. On the positive side, R&D institutions which are totally financed by government, suffer because they tend to forget the reason for their existence. They have often become a form of auxiliary to an academic institution where higher degrees are collected by scientific boffins who, in their quest for academic excellence, tend to disregard the benefits that industry may ultimately gain from this research endeavour. The trend in recent years - and South Africa is no exception - has been for the state to progressively reduce its subsidy, whilst at the same time encouraging research organisations [for their ongoing survival] to interact closely with industry, and to conduct work directly for them on a contract basis. Organisations such as CSIRO in Australia, CANMET in Canada, the USBM in the United States and Mintek in South Africa all generate income from the private sector. However, the critical fact is that the State should never withdraw

Jan 1, 1995

By Floyd C. Bossard

A method of hand-calculating the air-cooling requirements of a conceptualized underground mining operation is presented for the reader's orientation. Separate air heat load calculations were conducted for adiabatic compression, electromechanical equipment, wallrock, broken rock, groundwater, and blasting operations. The air heat sources were calculated for four mining levels under conditions representative of a typical planned mining operation at depth. The total heat gain on each level will approximate the level air-cooling requirements. The results of these hand calculations can be further modified by the use of mine ventilation computer software that refine the heat-source calculations, predict underground ambient air temperatures, and establish the air-cooling requirements of a mine. INTRODUCTION The principal mining method employed is a modified vertical crater retreat (VCR) blasthole operation. Typical scopes range from 20 to 40 feet wide, 75 to 100 feet long, and 100 feet in height. Six and one-half inch blastholes are drilled with an "in the hole" hammer drill. Eight-foot high horizontal rounds are blasted down. Mucking of ore from the undercut to orepass is done with LHD equipment operated from a remote control station. Backfill includes hydraulically placed tailings with cement, and waste rock when available. Stope access is from ramp sub-levels on 50-foot vertical intervals. Crosscuts are ramped down to the first stope cut 25 feet below the sub-level elevation. Then the crosscuts are raised by taking down the back when each stope cut is completed, until an elevation 25 feet above the sub-level is reached. The crosscuts are filled with tailings, and/or waste rock. See Figure 1. Typically, a two-pronged approach to defining the air-cooling requirements is conducted. First, the principal sources that make up the air heat load are individually hand-calculated. Second, projections of mine heat load are calculated by utilizing computer modeling techniques. This paper discusses the first method (hand calculations) of determining the individual components of heat flow into the mine. CALCULATED AIR HEAT LOADS Adiabatic Compression The plans for the mine include delivery of 300,000 cfm of air to ventilate the lower level operation. This is equivalent to approximately 200 cfm/ton of rock produced (300,000 cfm/1500 tpd of ore and waste). As air flows down a shaft, with no heat interchange between the shaft and air and no evaporation of moisture taking place, the air is heated in the same way as if it were compressed in a compressor. Dry air increases in temperature about 5.4°F per 1000 feet. One BTU is added to each pound of air for every 778 feet of decrease in elevation, or is subtracted for the same elevation increase. For dry air, the dry-bulb temperature change is 1/(0.24 x 778) = 0.00535°F/ft., or 1°F/187 ft. elevation. Auto-compression may be masked by the presence of other heating or cooling sources, such as shaft wallrock, groundwater, air and water lines, electrical facilities, etc. The major factors influencing the temperature of the air delivered underground by a shaft are (1) the night time cool air temperature's effect on the rock or lining of the shaft, (2) temperature gradient of ground rock related to depth, and (3) evaporation of moisture within the shaft which increases the latent temperature and decreases the sensible temperature. [Calculation For Adiabatic Heat of Compression] a. Assumptions: 1. Three hundred thousand cfm of fresh air at 3000 level has increased in temperature during the summertime to the point where it has little available cooling power. Air-cooling will be required on 3000 Level and below. 2. Elevation of 3000 Level is approximately +1000 ft. above sea level.

Jan 1, 1993

By J. Burrows

The Chamber of Mines of South Africa is an association of major mining corporations, gold mines, coal mines, diamond mines and other metalliferous mines. It looks after those interests of its members which can best be handled on a co-operative basis. The Chamber's activities include the engagement of labour; the negotiations of all labour matters, including minimum wage rates, with recognised trade unions and associations; the carrying out of advanced technological research; representing the industry in dealings with the government and the administration of miners' training colleges and rescue training stations. Subsidiary companies of the Chamber handle matters such as the refining of gold and other metals; the processing and marketing of uranium and the marketing of Krugerrands. The thirty-two gold mines members of the Chamber of Mines of South Africa produce approximately 694 tons of gold per annum. To achieve this they employ 430000 persons and break about 8400 kilotons of rock per month. The working revenue is approximately 4 225 million US dollars per annum and the annual working profit is approximately 2 050 million US dollars. The electric power consumed by the gold mining industry is 1650 MW, about 20% of the total power produced in South Africa, and the annual costs of this power is approximately 200 million US dollars per annum. It is estimated that about 25% of this power is used for ventilation and refrigeration appliances. Thus the annual power costs for ventilation and refrigeration is approximately 50 million US dollars. The gold bearing reef is a narrow tabular deposit situated at varying depths throughout the gold fields. The weighted mean rock breaking depths and, virgin rock temperatures range from 1 700 - 9 000 ft (520 - 2 740 m) and 68 - 113°F (20 - 45°C) respectively. The maximum virgin rock temperature at which mining currently takes place is 130°F (54.5 C). Obviously at such depths and temperatures and with such a large labour force the provision of an acceptable environment is critical and to achieve this 59 000 000 cfm (28 000 m3/s) of air is downcasted during the main shift and 122500 tons (430 MW) of refrigeration are provided. Legislation in South Africa requires that on metalliferous mines when more than 1000 persons are employed underground during any one shift the manager shall appoint one or more ventilation officers. The legal responsibilities of the ventilation officers are concerned mainly with the measurement and control of dust. In practice, this is not the main part of their work and ventilation personnel perform many other duties such as measuring and controlling the distribution of air, checking the performance of fans and cooling plants, planning air and refrigeration requirements and, more recently, measuring and controlling noise and illumination levels. Ventilation personnel are specialists who provide management with advice on all matters relating to the mine environment. At the present time there are 506 persons engaged as ventilation personnel in gold mines and a further 1744 persons are employed in gold mine ventilation departments in a non-technical capacity, for example as construction workers, filter cleaners, fire patrollers, etc.

Jan 1, 1980

By D. Bleazard

Barrick Goldstrike Mines, Inc. is actively involved in improving the reliability of mining and processing equipment through the use of failure analysis and condition monitoring activities. Program results include an increase in equipment capacity and an overall reduction in operating costs. The identification of equipment failure modes and the development of associated capital asset management practices can minimize failure consequences. Successful programs have included; alterations to management practices, equipment modifications, and/or the implementation of condition monitoring programs. By controlling or minimizing the effects of incipient equipment failures, corrective actions can be taken earlier in the failure cycle, resulting in an associated reduction to repair costs and equipment down time.

Jan 1, 1997

By Steve O. Green

Introduction The US coal mining industry is in the midst of a serious economic crisis. To overcome increased operating costs, operators must take advantage of technological advancements in mining systems and equipment. A relatively new concept available to underground operators is that of extending the cutting depth for continuous mining machines (continuous miners). This system is referred to as deep-cut mining, and may be described as mining cuts to depths greater than the standard 6 m (20 ft). Productivity is increased through the reduction in continuous miner tram time since the machine stays longer in each working face. To successfully implement a deep-cut mining system, three problems must be overcome: Exposing personnel to unsupported roof, methane removal, and controlling respirable dust. Remote control cutting The solution to the unsupported roof problem is remote control. Recent advances in remote operation, especially in radio remote controls, have resulted in what is generally accepted as a safe and reliable tool. By installing remote control on the continuous miner, the operator and helper can now be stationed in a supported area. This allows for improved visibility while cutting coal. With this in mind, the limiting factor now becomes the position of the haulage equipment operator. North River No. 1 uses Joy 21SC shuttle cars and Jeffrey 4110 diesel ramcars. These cars are pulled against the bumper of the continuous miner. Thirteen meters (43 ft) can be cut with the shuttle car and 14 m (45 ft) with the Ramcar before the operator of the car is inby permanent supports. By loading away from the bumper, about 1.5 m (5 ft) can be added, bringing total cut depths up to 14.5 and 15 m (48 and 50 ft), respectively. Methane removal The second problem involves methane removal, and the line curtain-to-face distance. In other words, how to maintain the line curtain to allow sufficient ventilation to remove methane and not expose workers to unsupported roof. The solution, in most cases, is to use a blowing-type ventilation system. Unlike the exhausting system that, to be effective, must be installed within 3 m (10 ft) of the working face, the blowing system can be effective at greater distances from the face. Numerous studies have documented the advantages of a blowing-type ventilation system over an exhausting system. Furthermore, a blowing system can be significantly improved by in¬creasing the velocity of the air delivered without increasing the volume of air. With this greater ventilation efficiency, it is possible to secure ventilation plan approval from the Mine Safety and Health Administration (MSHA) for a greater curtain set-back distance. This eliminates the need for personnel to travel inby permanent supports to extend the face brattice. The distance from the face MSHA will approve, however, is on a mine-by-mine basis. MSHA will evaluate the particular mine's face methane liberation rates, along with the specific proposals regarding face ventilation volumes and velocities. Dust control The third problem is controlling respirable dust. This is actually a result of the solution to the second problem. In the blowing system, while mining on the slab cut, both the continuous miner and haulage units are located on the dusty side of the line curtain (Fig. 1). In the early 1970s, underground operators virtually abandoned this system, in order to comply with the 2 mg/m3 (0.87 gr per cu ft) federal dust standards. The blowing face curtain did offer increased face ventilation efficiency. With that, greater curtain-to-face distances were also

Jan 8, 1987

By Stanley W. Ivosevic, Mark E. Emerson

Introduction Metals exploration activity is moderately intense in the US despite the 1981-82 recession in the mining industry. Although exploration expenditures are significantly less than the 1981 peak, quality and goal orientation of current exploration are high. Major foreign and domestic mining companies as well as nonmetal companies had been attracted to invest in minerals in the US. Its capital markets, political stability, and open-door policy are attractive for hard mineral explo¬ration, particularly in the West. The approach was to either acquire fully or partially developed ore bodies through merger and acquisition, or to fund large scale exploration programs in search of bulk-tonnage deposits. The acquisition avenue has not proved the panacea envisioned for overcoming lack of substantive mineral development. Nor has the decentralized approach to exploration that was adopted proven cost-effective. The emphasis on gold, to the exclusion of all other commodities, raises questions regarding the direction of exploration in the US for the rest of the 1980s. Overview More than 300 companies are active in metals exploration in the US. These range from established, well-known domestic and foreign corporations to hurriedly incorporated, unknown companies financed by having promoted mineral prospects or occurrences of questionable quality. Junior companies tend to be oriented toward property entrepreneurship, rather than the development of minerals resources. These companies have helped maintain mineral lands from prospect to production by holding blocks of property intact during periods of exploration inactivity. This was done through promotional efforts, even though few properties that attracted major venturers proved to have real merit. Exploration objectives of senior and junior exploration companies currently embrace only a few commodities. Arranged in approximate descending order of priority, these are: precious metals; base metals in monometallic or polymetallic deposits; strategic metals, such as tungsten, tantalum, and tin; unconventional commodities, such as diamonds; ferrous metals; and commodities such as beryl, zeolites, and barite. Prospectors are important in searching for these targets. In the US, where they have been financed by major and junior companies, prospectors have proven a useful and low-cost means of locating gold prospects (e.g. the Alligator Ridge property in Nevada). The tenacity and persistence of independent prospectors have been factors in discovering properties rejected or overlooked by professional geologists (e.g. Lornex mine in British Columbia). This has been realized by the provincial government that grubstakes bona fide prospectors. Exploration interest in commodities has always been cyclic with price. Gold is one of the few metals that has experienced a price increase of such magnitude to warrant attention. The potential for realizing adequate operating margins for the Nevada open-pit "no-see-um" gold deposits is great. The present declining gold price, however, may indicate that this emphasis on gold, to the exclusion of other commodities, is shortsighted. Mining companies have been the last to react to price increases, opening new mines based on unrealistic market expectations. A scenario could develop where plummeting gold prices bring recession to the gold industry, as lately occurred in uranium, molybdenum, and copper. US Exploration Climate The US is attractive for exploration due to its strong mineral po-

Jan 4, 1984

By R. O. Burt

The Tanco pegmatite is located in the Bird River greenstone belt of south-eastern Manitoba approximately 160 km northeast of Winnipeg (Figure 1). The orebody was discovered in 1929 and subsequently attempts were made to exploit it for tin by Jack Nutt Mines Ltd. and later in the middle 1950's by Montgary Oil Ltd., who extensively developed but did not produce from its lithium deposits. In 1966 systematic investigation of the pegmatite as a potential source of tantalum led to the founding of Tantalum Mining Corporation of Canada Ltd. (TANCO), and production of tantalum concentrates from the Bernie Lake mine commenced in 1969. Tanco's currently jointly owned by Hudson Bay Mining and Smelting Co., Cabot Corporation, and the Province of Manitoba.

Jan 1, 1988

By Ralph A. MacMillan

Electronic systems introduced for haulage equipment in recent years probably exceed what mine management is prepared to use. Haulpak trucks alone are equipped with General Electric's Statex III drive system controls, Detroit Diesel's DDEC, Cummins' Centry engine management system and the Haulpak Management System (HMS). The future holds promising developments of even more sophisticated electronics that will lead to unmanned truck operation and control. Trolley-assist will finally come into its own due to the importance of environmental issues, as well as the high potential for increased productivity with dramatically lowerfuel consumption. All of these high technology systems do not mean much unless they add value to mining operations. Some of the benefits that can be expected from these systems are reviewed. In the near future, haul truck control and monitoring systems will be fully linked by telemetry to a mine's dispatch system, as well as to its maintenance management information system. The manufacturer and distributor can also be kept in the communications loop, so that they can better serve customer needs. Portable computers will allow technicians to talk to machine systems to quickly locate a problem, look up service information, check what parts are required and electronically order them.

Jan 1, 1994

By André Vervoort

At the 2017 ICGCM-conference in Morgantown, WV, a case study was presented with measurements of the upward surface movements or uplift after the closure of two neighboring coal mines (Winterslag and Zwartberg, Belgium). One was closed in the sixties, and the other closed in 1988. The study was based on satellite data (Radar-interferometry or InSAR). The analyses showed both a significant variation of uplift values above the mined area and the surrounding zones, and a variation of the uplift rate as a function of time. In other words, the phenomenon of uplift is complex. At that time, it also was clear that more research was needed to better understand this phenomenon, e.g., an identification of the various mechanisms that are relevant for the process of uplift and the quantification of the various impacts. A systematic analysis of all relevant parameters resulted in a framework that forms the basis for the analytical calculations of uplift values. These calculations show that the main impact is due to changes in the hydrostatic pressure. However, the expansion of the softest parts, i.e., the collapsed or goaf material, is not only important, but the expansion of the (stiffer) non-collapsed strata layers due to a change in pore pressure also must be integrated into the model. For the case study, the partial flooding of the underground after closure of the mine that was abandoned in the sixties is integrated into the model. Larger stiffness values also are applied for this mine, in comparison to the values of the more recently closed mine. Overall, a very good correlation is observed between the InSAR-measurements and the calculated values along a north-south transect.

By R. L. White

Classification of Mills Mills may be classified under four categories that relate to the type of action utilized to grind the material: Tumbling. The most important of these to the minerals processing industry are tumbling mills. Grinding media is actuated in a free¬form pattern in a rotating drum causing it to tumble, material being ground passing through the tumbling media. Work is done by a combi¬nation of attrition and impact. Impact. Grinding media is actuated in a distinctive pattern so that it impacts the material being ground. Work is done by impact. Roller. Grinding media is actuated in a distinctive pattern, mate¬rial being ground passing under or between the media. Work is done by compression. Fluid Energy. There is no grinding media as such. Energy is imparted to the material by transfer of kinetic energy from the fluid to the material. Grinding is accomplished by interparticle impact and attrition. Tumbling Mitts The main physical arrangement consists of a lined cylindrical drum mounted on trunnion bearings driven by a ring gear and pinion, in turn connected directly or through a reducer to the electric motor. Variations to this general arrangement include: (1) Drum mounted on rollers instead of trunnions. (2) Bearings on the drum directly, not on trunnions. (3) Utilization of multiple pinions on the gear. (4) Trunnion-type drive arrangement through a gear reducer with¬out the use of a ring gear. (5) Motor rotor mounted directly on the drum with stator around the drum, thus eliminating the gearing. The drum consists of a shell and heads lined with replaceable liners of various types and configurations. Various types of tumbling mills are tabulated; illustrations of some are shown. The types are so designed for specific purposes. Media utilized include steel rods, balls, or rock of the same material as being ground or of special material in the instance where iron contami¬nation is detrimental (see Figs. 9 and 10 and Table 2). Impact Mills Although there may be some overlapping of interpretation of terms, impact mills are considered those in which the media is actuated in a distinctive manner, at relatively high speed, impacting the material to be ground. There is an overlapping also in what may be considered to be crushing equipment rather than grinding equipment. Hammer Mill or Impactor. Material is fed into the equipment; inside hammers of impactors revolve at a high speed, breaking the material on impact with the hammers (see Fig. 11). Material is also thrown against the periphery of the mill, causing impact grinding to occur. Equipment is normally lined with replaceable abrasive resis¬tant metal; hammers are also

Jan 1, 1985