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By Y. J. Wang

Optimal sizing of a ventilation shaft is a classical subject in mine ventilation economics.. The shaft size is optimized by minimizing the sum of shaft-construction and ventilation-power costs, or owning and operating costs. The length and air quantity of the shaft are held constant, and the location of the shaft, affecting its depth, and construction and ventilation power costs, is not considered (Hartman, 1961; Lambrechts, 1974; Wang et al., 1978). This paper proposes an approach by which the location and size for a new pair of shafts (intake and return shafts) are optimized for a coal-mine ventilation network. Specifically, the problem is to determine the location and size for a new pair of shafts, and the distribution of air between the new and old shafts so that the combined cost of shaft construction and ventilation power is minimized. The shaft length or depth corresponds to the height of overburden at any shaft location. To simplify the problem, the shaft locations will be limited to the line along the main or submain entries, and the connection entries at the shaft bottom neglected. In order to calculate the ventilation-power cost, the air power of the system should be determined. This in turn requires the calculations of the air quantity and pressure of each fan in the network, satisfying Kirchhoff's laws and Atkinson equation. After obtaining the air power, the cost calculation is straightforward. In the following, we shall assume that the resistances of all airways, and the quantity for each mining section, which is controlled by adjusting the air regulator, are all known. The resistances for the shafts are determined from the estimated friction factors, and shaft dimensions and lengths. Furthermore, it will be assumed that the fans can be selected to meet the duties called for by any network solutions.

Jan 1, 1982

By Nora M. Hutt, Kevin A. Morin

Geochemical static tests are a basic component of geochemical assessments and predictions for acidic drainage, and include acid-base accounting and total-metal contents. A key database for such information is the International Static Database (ISD), which currently contains more than 45,000 static-test analyses from over 200 minesites. Such large databases are useful in identifying general trends and, just as important, the percentage of exceptions. General trends and significant exceptions can be seen in the ISD by comparing one parameter against another. This paper shows that bulk neutralization potential (bulk NP) obtained from acid titration is often a good indicator of neutralizing capacity above values of 10 t CaCO3/1000 t. Other types of calculated neutralizing capacity are then compared to bulk NP to determine if they, in fact, reliably represent actual neutralizing capacity and if they can substitute for the more time-consuming and expensive bulk NP. For example, bulk NP correlates better with Carbonate NP (from carbonate analyses) and Total-carbon NP (from total-carbon analyses) than with (Ca+Mg) CaNP (from solid-phase analyses of calcium and magnesium). This means that, in general, (Ca+Mg) CaNP is a less accurate substitute for bulk NP than the other types, and thus requires intensive site-specific justification if a substitution is attempted. This also means that measurement of the frequently less-expensive parameters of total carbon and carbonate can often substitute, within site-specific error bars, for the bulk-NP titration. However, the ISD also shows that non-carbonate minerals make the primary contribution to neutralization in up to 20% of samples, so Carbonate CaNP and Total-carbon CaNP do not always reveal the full neutralizing capacity of a sample.

Jan 1, 2000

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

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

Jan 1, 2000

By Subhash Sethi

The WIPP (Waste Isolation Pilot Plant) Project near Carlsbad, New Mexico. is a research and development project sponsored by the U.S. Department of Energy, designed to demonstrate the safe disposal of radioactive waste. A mine (repository) is being developed 2150 feet below the surface in a thick salt bed, which will serve as the disposal medium. The underground ventilation system for the WIPP project was originally designed based on a fixed project scope. The design criteria and ventilation requirements were developed for the performance of various activities as then envisioned towards the achievement of this goal. However, in light of new information and actual site-specific experience at WIPP leading to a clearer definition of the scope-related programs and activities, it was realized that the existing ventilation system may need to be modified. A re-assessment of the ventilation requirements--based on the current status of the facility and a clearer definition of the scope of the experimental programs, waste storage methodology, mine developmental activities. and current equipment list--indicates a two- fold increase in underground airflows from 210,000 cubic feet per minute (cfm) to 425,000 cfm. Various options to provide the increased airflow to the underground facility were considered and analyzed. Based on the techno-economic analysis, safety and environ- mental considerations, and other conditions unique to a facility like WIPP, it was decided to construct a new shaft to serve as the primary fresh air intake, along with the installation of additional fans in parallel with the existing set up. Low cost and the ease with which the new design could be implemented and integrated into the existing ventilation system with minimum impact on current activities and projected schedules were the guiding factors in designing the shaft and related ventilation modifications.

Jan 1, 1987

By Maurice C. Fuerstenau

INTRODUCTION The term, chelate, from the Greek for 'claw,' has its origin in the pincer-like representation of more than one atom of a single ligand bonded to the same metal atom, forming a ring-like structure. Depending on how many attachments exist on the sphere of coordination around the metal ion, ligands are referred to as bidentate, tridentate, tetradentate, pentadentate and hexadentate According to Calvin and Martell (1952) and Somasundaran and Nagaraj (1984), chelating agents must contain suitable functional groups, and their location within the ligand molecule must enable ring formation that is sterically possible. Typical structures that meet these requirements have been given by Fuerstenau and Herrera-Urbina (1988) and are listed in Figure 1. As noted, the functional groups of these reagents contain electron donor atoms such as oxygen, nitrogen and sulfur. Basic groups, such as amino (-NH,), are those that contain an atom with a lone pair of electrons. Acid groups are those, such as carboxylic (-COOH) which lose a hydrogen ion and coordinate with the metal atom. Examples of each of these types chelating reagents are given in Tables 1-4. The specificity of chelating agents for various metal ions has long been known and was utilized early in the search for new flotation collectors. In 1927, Vivian used ammonium nitrosophenylhydroxylamine as a collector for cassiterite. Holman (1929) tested dimethylglyoxime as a collector for nickel oxide. He also suggested taurin as a collector for oxidized lead ores. DeWitt and von Batchelder (1939) suggested the use of oximes as flotation collectors. Gutzeit (1946) conducted an extensive investigation on the use of chelating agents as collectors and as depressants for gangue minerals. The depressants are those reagents that form soluble chelates with metal ions that could function as

Jan 1, 2005

By Malcolm Reeves

The heavy oil reservoirs of Alberta and Saskatchewan include sequences of unconeolidated silts and fine sands which respond in an unusual manner when subjected to the high temperature and pore-fluid-pressure associated with steam injection and other thermal enhanced recovery techniques. Apparent transient connections are suddenly established between producers and injectore and there is an associated enhancement in oil (and sand) production. These phenomena have been attributed to local fluidization and piping processes which are known to occur in situations where high pore-fluid-pressures prevent the mobilization of intergranular friction. These breakdown processes may be associated with the penetration of tongues of the heated reservoir fluids with suspended sand into the creep sensitive intact reservoir since the viscosity ratios are sufficiently high for viscous fingering instability of the fluid-fluid interface. A program of laboratory tests has been carried out to measure the strength characteristic of saturated oil sands at a variety of temperatures and fluid pressures. These tests suggest, that in common with the coarser Athabasca sands, the Waseca sands have very little cohesive strength and show the characteristic dilational failure response of dense sands. Some special tests, devised to determine the relative ease with which samples can be fluidized, confirm the expected susceptibility of the fine- sand silts. Both piping and fluidization can be readily induced at in-situ reservoir temperatures and pressures. Analysis of stresses around wellbores suggests that fluid-pressure gradients may be sufficient to initiate piping and that backward erosion along streamlines may link producers and injectors. Such pipes may act as well “adits” gathering oil from the heated reservoir and contributing to enhanced production rates. Mechanical breakdown of the reservoir by drilling, injection and production processes provides a consistent possible explanation for the sand-cut response of production wells.

Jan 1, 1986

By L. A. Puchkov

The paper presents the outline of the results in the algorithms and software of the microprocessor-based system for the powerful drugline bucket automatic motion, as well as the technological process control and functional diagnostics of the electrical equipment. The system is, aimed at the increasing both the draglines' working efficiency and the professional, competence of the operators. The excavator-dragline, equipped by the automatized operation, diagnostic and information processing system, possesses considerable technological: flexibility, which can widen therefore its field of application. The tasks of the dragline bucket motion optimal operation, programming of the excavation and overburden dumping processes, as well as the tasks of the control and the analysis of the technological parameters and the diagnostics of electrical equipment conditions, are defined in the paper. In our paper we present the results of these tasks being solved, providing the algorithms of the main mechanisms for the automatic operation for the swinging bucket controlled motion, the algorithms of the statistical analysis of the working efficiency parameters the algorithms of the diagnostics for electrical driving systems. The algorithms of the motion operation are constructed using the predicting dynamic models. The diagnostic algorithms at the normal working conditions form a number of algebraic invariants always equal to Ø with the accuracy up to when the electrical equipment is parametrically adequate. The algorithms of evaluating and analysis of the excavation efficiency make use of the direct measurement and analytical identification. The working condition and quality of the operation and control algorithms are investigated by simulation and natural testing. The results of the mathematical simulation of the bucket motion (with the 20 m3 capacity) of the excavator-dragline with the boom of 90 m in length are given in the paper.

Jan 1, 1992

By E. W. Geist, W. J. Bronkala

PRINCIPLES OF MAGNETIC SEPARATION The application of magnetic separators relies on the different behavior of individual mineral particles under the influence of magnetic forces. When exposed to a magnetic field, all materials are affected in some way. According to whether materials are attracted or repelled by the magnetic field, they are designated paramagnetic or diamagnetic, respectively. Paramagnetic materials are attracted along lines of magnetic force from points of lower magnetic field intensity to points of higher magnetic field intensity. The very strong response of some materials to a magnetic field is sometimes designated ferromagnetic and the particles attracted frequently retain magnetic properties after the applied magnetic field is removed. Diamagnetic materials are repelled along the lines of magnetic force from points of high magnetic intensity to points of lower field intensity. Only limited commercial use of diamagnetic effects has been made. For practical applications, the paramagnetic groupings are: 1. Strongly magnetic (ferromagnetic) 2. Magnetic 3. Weakly Magnetic 4. Non-magnetic The limits of these groupings are not well defined and will vary in some mineral species. Only a few minerals are ferromagnetic (magnetite is an example) and arbitrarily some investigators have selected 10,000 gauss, or 1 Telsa, as the dividing line between low and high intensity magnetic fields. I. S. Wells has developed a table indicating the magnetic intensity required to separate a selected grouping of minerals which is shown in Table 1 on page 2. A similar table has been developed by Carpco Inc. of Jacksonville, Florida in conjunction with the use of its high tension separators. This table, Table 2 on page 3 shows other minerals grouped in the broader categories cited above. Magnetic separation methods are used either to separate valuable minerals from nonmagnetic waste or magnetic impurities or other valuable magnetic minerals from bulk nonmagnetic values. Magnetically susceptible mineral particles occur as either individually discrete particles,

Jan 1, 1986

By Leo J. Miller

Sulphur deposits occur within the upper Salado and Rustler gypsum beds. The formation of a sulphur ore deposit requires large quantities of oil and /or gas, which must find a permeable pathway into the gypsum beds. The Permian beds weathered for 90 million years in an arid environment prior to the advent of a more humid climate at the beginning of the Cretaceous Period. The first Cretaceous rivers enlarged the sinkholes and caverns of the Karstified Permian terrain and deposited coarse conglomerates and sands into the subsurface gypsum beds. The resulting network of caverns and sinkholes, filled with permeable sediments, permitted the free flow of hydrocarbons through gypsum. All sinkholes and caverns infilled with Cretaceous sediments that have penetrated beds of gypsum are associated with sulphur deposits. The Castile gypsum lies directly upon the oil bearing Bell Canyon formation. Sinkholes and caverns filled with Cretaceous sediments occur in the lower Castile. Biogenic calcite and sulphur are adjacent to these sediments. However, faults assume the major role in generating the initial permeability, allowing hydrocarbons to penetrate the lower Castile gypsum. The largest faults are associated with biogenic calcite and sulphur for a distance of 250 meters up from the bottom of the Castile gypsum. It is apparent that two dynamic geological processes are generating permeability within the gypsum beds: (1) the Cretaceous sands decending with difficulty into gypsum and (2) the faults creating permeability in the lower Castile but losing energy through soft bed deformation in the overlying gypsum. Both of these sulphur ore controls are mappable, but another large sulphur deposit of the magnitude of the Culberson Sulphur Mine has yet to be discovered. It appears that massive sulphur deposits are associated with the largest scale examples of each system, that is (1) high densities of deep sinkholes and separately

Jan 1, 1992

By Howard M. Wells

Investment in minerals development usually entails big financial outlays. which will remain at risk for a long time before the venture can begin to generate a return cash flow. Before 'committing himself to such a venture the investor would be wise to carefully examine the possible consequences of his exposure to risk. In this context risk is defined in multiple terms, which relate the magnitude of possible loss with the probability that such loss may occur--normally the probability will decrease as the amount of possible loss increases. The analysis of risk is an essential component of the decision making process, especially so in the case of low grade high tonnage mining ventures where a very real possibility exists that future unfavorable circumstances may lead to losses that could cripple or ruin even the richest corporation. Wells (1979) postulates a rationale for making the investment decision under uncertainty based on the comparison of two independent risk criteria: 1) the risk inherent in the venture represented by the negative portion of the cumulative distribution of the net present value (NPV), as derived by a stochastic evaluation procedure, 2) the limits of risk the decision maker is willing to accept, taking account of the financial capacity-of the 'firm to absorb various levels of future cash shortfall should the venture prove disappointing. The chance the investor would be prepared to accept can be expected to decrease as the amount of the possible shortfall increases. The limits of acceptable risk are-defined by a curve called the "LL-curve" which can be directly compared with the cumu¬lative NPV curve. A positive investment decision is implied if the limits of acceptable risk are not exceeded.

Jan 1, 1982

By Chas. F. Jackson

INTRODUCTION This paper is one of a series dealing with mining methods and costs, prepared and published under the sponsorship of the US Bureau of Mines (USBM) and made possible through the cooperation of many mining companies in the US and foreign countries. The data presented have been obtained from papers previously published by the Bureau dealing with indi¬vidual mines, from the technical press, and from infor¬mation furnished by the mining companies. DEFINITIONS Top Slicing The term top slicing is commonly applied to that method of mining in which the ore is extracted by a series of timbered slices, beginning at the capping; the slice is caved, bringing down the capping to the floor of the slice; the succeeding lower slices are mined up to the overlying mat or gob, consisting of an accumulation of timber from the upper slices and broken capping, which follows the mining progressively downward (Fig. 1). The term has also come to be applied to the mining of ore bodies only one slice thick, by a series of tim¬bered slices one beside the other. The timbers are blasted out, bringing the capping down to the floor, suc¬ceeding slices being driven alongside those already caved (Fig. 2). In USBM classifications of mining methods, top slic¬ing has been placed under the heading caved stopes. The method differs fundamentally, however, from the cav¬ing methods of mining in which the breaking of the ore itself is accomplished by caving action induced by undercutting. In top slicing the ore is broken by the ordinary methods of drilling and blasting, and only the capping is broken up by caving. Main Levels or Haulage Levels Main or haulage levels are the main extraction levels on which the ore is hauled to the shaft. These levels and the shaft constitute the preliminary development of the mine.

Jan 1, 1982

By D. P. Tao, J. G. Groppo, R. Killmeyer, W. W. Wen, B. K. Parekh

Economical dewatering of ultra-fine (minus 74 m) clean coal slurry produced in column flotation to a 20 percent or lower moisture level will be an important step in successful implementation of the advanced fine coal cleaning processes. The main objective of the present study was to evaluate efficiency and effectiveness of three dewatering techniques, namely, hyperbaric, vacuum and screen bowl centrifuge, in dewatering of ultra-fine (D50 ~ 25 m) clean coal slurry produced by the `Ken-Flote' columns operating at the Powell Mountain Coal Company, St. Charles, Virginia. he hyperbaric filter using 3 bar pressure and 165o cake formation angle, provided a filter cake with 23.6percent moisture with a throughput of 800 Kg/m2-h and air consumption of 868 Nm3/t (460 scfm/t). The vacuum filter with 30 percent submergence provided a filter cake with 27.8 percent moisture with a throughput of 122 Kg/m2-h. The screen bowl centrifuge provided a dewatered product containing 32.9 percent moisture at a feed rate of 450 Kg/hr, however, the solids capture was only65 percent. Addition of 1 Kg/t of a cationic surfactant provided filter cakes with 19.1 percent, 23percent, and 34 percent moisture using hyperbaric, vacuum, and centrifuge, respectively. Similarly, addition of 0.5 Kg/ton of copper chloride lowered filter cake moisture to 20.7 percent and 24.8percent, using hyperbaric and vacuum. With the centrifuge no improvement in filter cake moisture was noted. However, use of 80 Kg/ton of `orimulsion' in the centrifuge provided dewatered product containing 26 percent moisture, and the solids capture increased from 65 to 95 percent. The dewatered product appeared dry and was easily handleable.

Jan 1, 1999

By P. Bissert, L. Yan, D. Yantek

"Federal regulations require refuge alternatives (RAs) in underground coal mines to sustain life for 96 hours while maintaining an apparent temperature (AT) below 35°C (95°F). NIOSH research has shown that heat and humidity buildup is a major concern with RAs because they have a limited ability to dissipate heat, and high internal air temperature and relative humidity (RH) may expose occupants to heat stress. The heat transfer process within and surrounding an RA is very complex and is not easily defined analytically nor experimentally. To investigate heat and humidity buildup in RAs, NIOSH conducted multiple in-mine 96-hour tests on a 6-person metal-type RA, a 23-person tent-type RA, and a 10-person tent-type RA. The test results show that when moisture was introduced into the tests to represent perspiration and respiration from miners (wet tests), the average measured internal air temperature at midheight increased by 7.7°C, 9.4°C, and 10.5°C, and the RH approached 95%, 94%, and 88%, respectively. For the dry tests when no moisture was introduced, the average internal temperature increased by 8.4°C, 10.3°C, and 12.6°C, respectively. INTRODUCTION If an accident occurs in an underground coal mine, miners who cannot escape from the mine can enter an RA for protection from adverse conditions, such as high carbon monoxide levels. One of the main concerns with the use of mobile RAs is the thermal environment inside the RA. The metabolic heat of the occupants and the heat released by the CO2 scrubbing system will cause the interior air temperature to increase. Moreover, the humidity within the RA will increase through occupants’ respiration and perspiration and from the chemical reaction within the CO2 scrubbing system. The internal thermal conditions can subject miners to heat stress, which can lead to heat exhaustion, heat stroke, or even death depending on duration and magnitude of exposure. MSHA regulations require that RAs must be designed to ensure that the internal apparent temperature does not exceed 35°C (95°F) when the RA is fully occupied [1]. Apparent temperature (AT) is a temperature-humidity metric for the perceived temperature caused by the combined effects of air temperature, relative humidity (RH), and wind speed. It is used to assess the perception of indoor temperatures when workplaces are not sufficiently heated, cooled, or insulated to provide comfortable or healthy conditions. Before a mobile RA can be deployed at an underground mine for emergency usage, it must be tested to show that its internal AT when fully occupied will not exceed the 35°C (95°F) limit. For practical reasons, RA manufacturers usually conduct their tests at above-ground test facilities. To investigate RA thermal response in an in-mine environment, NIOSH conducted multiple 96-hour tests on a 6-person metal-type RA, a 23-person tent-type RA, and a 10-person tent-type RA in its underground coal mine facility. The work described in this paper could be used by RA manufacturers to determine a final temperature rise based solely on dry and time-reduced testing."

Jan 1, 2017

By Samuel A. Reipas

Introduction The ore mining industry has been successful because of technological progress and huge sums of capital investment. But new technology and plants have not come cheap. The problem now is companies will run out of money before they run out of ideas. Survival Strategies Industry survival strategies demand restructuring and new plans. Inland's prescription for the future includes: • Continued emphasis on technological improvements to upgrade pellet quality and lower costs; • Effective use of human resources resulting in employee satisfaction and improved planning and productivity; • Deliberate, well planned, and documented cost-control programs for the short- and long-term; and • Capital investments to support the first three initiatives. Technology Technological changes are not only required by the need for operational improvements, but by product quality improvement dictated by the user - usually the blast furnace operator. Established ambient measurements of pellet quality, such as tumble index, compression strength, porosity, size distribution, and chemical composition, are no longer considered meaningful in predicting pellet performance in a blast furnace. New specifications, called low temperature and intermediate temperature properties, have been recently introduced. Low temperature specifications are characterized by disintegration of pellets in the 500°C (930°F) range in the upper portion of the granular zone during the hematite-to-magnetite transformation in blast furnace stack. Intermediate temperature properties (measured by 1000°C or 1800°F) are characterized by swelling, reducibility, and compression strength after reduction, reflecting conditions in the lower portion of the granular zone of the furnace. These new parameters are being used for pellet quality evaluation at Inland, and will be introduced at its Minorca mine for more routine analysis and correlation with the pellet furnace operation. With increased emphasis on quality, effective communication between the blast furnace operator, research, and the pellet plant operator is critical. At Inland, this is accomplished through technical meetings where the opportunity is provided to solve problems, establish a common understanding on the value of quality, and develop future quality standards and programs. With the degree of sophistication in determining pellet quality as it affects blast furnace operation, the future pellet plant operator will actually be producing a custom pellet. Human Resources Technological change has been a strong point at Inland, and will continue to be. However, human resource management is equally important. Is management giving sufficient attention to employee motivation and involvement? It is a way to improve productivity, reduce costs, and enhance employee job satisfaction. The industry is beginning to understand the principles and the potential for tapping employee brainpower. The payoff can be handsome. Management has to be willing to share both information and power if participative management is to live up to its potential. It has to become more flexible and be willing to adapt to change on the human side. Management should embrace rather than fear employee involvement. Given time, joint problem solving will replace adversarial attitudes and pay dividends for both sides. Productivity and product quality will improve and new dimensions of employee job satisfaction will be achieved. But, it will not happen overnight. Attitudinal change must permeate the organization from top to bottom. This change is needed to support the working relationships and give everyone a sense of direct and meaningful involvement.

Jan 11, 1983

By Alfred Weiss

Selecting a minicomputer is not unlike selecting a costly piece of mining equipment such as a truck, shovel, crusher, or conveyor belt system. A marked difference might lie in the frequency at which such selection takes place; another difference might be mine management's familiarity with mining equipment vis a vis a minicomputer. These basic differences might prompt a mining company to seek the aid of a consultant when selecting a minicomputer system. Advantages and Disadvantages The fundamental advantage of employing an independent consultant is that it provides management with a source of knowledge and breadth of experience not readily available within the organization. If the consultant has a minerals background, such knowledge and experience are further enhanced by actual experience in exploration, mining, and mineral processing. If the consultant has worked within a mining company, such hands-on experience is supplemented with a meaningful understanding of the client's business; this facilitates communication between the consultant and members of the client organization. The consultant has been, so to speak, "in the client's shoes." There are several other advantages. Since computer systems cut across organizational boundaries, the independent consultant may be able to move within the organization more easily than an inside staff person to collect necessary information. Also, detachment from any manufacturer should provide the client with an objectivity and a set of recommendations as near optimal as feasible. The principal disadvantages lie in the area of continuity and accountability. Once the consultant's report has been submitted and his recommendations implemented, a discontinuity might occur. Also, if implementation does not provide the desired results, the consultant might not be available to correct the situation or account for actions taken. Proper selection of consultant services, relative to kind and duration, and the consultant's method of operation, will often prevent many of these aftereffects. Services Offered The professional consultant can help a client: •Define and summarize information systems technology* needs; • Evaluate alternatives such as in-house capability and capacity, time-sharing, or service bureaus; • Prepare a formal request for proposal (RFP) to be submitted to vendors; •Evaluate vendor proposals; •Evaluate the economics of lease vs. purchase; and •Assist in the conversion and installation. Thus, consulting services may range from evaluating an existing system and upgrading to a larger piece of equipment, to selecting a new first-for-the-organization computer. First, Define and Summarize Needs Regardless of circumstances, the most valuable document required before equipment selection is the information system plan (I/S plan). It helps insure that the company's information needs are met in a cost-effective manner. An I/S plan is usually a comporient of the mining company's overall business goals. Since an I/S plan is a formal vehicle for establishing and communicating the direction the organization intends to take to satisfy information requirements needed to achieve goals and objectives, most initial systems services are likely to center around creating such a document or updating an existing document. The trend toward company-wide I/S planning is increasing in the mining industry; however, smaller scale efforts focused at the information systems needs to support an individual

Jan 11, 1981

By H. Nematollahi

The comminution laws The grindability of the ores or minerals is an important factor that could not be deduced from their mineralogical characteristics (Testut, 1958). Many theories have been proposed to determine the amount of work required in a comminution operation as a function of the nature of material, size and shape of the minerals before and after of comminution. The most important laws are proposed by Rittinger, Kick and Bond. According to Rittinger (1867), the work is directly proportional to the new surfaces produced during the comminution. Kick (1885) believes that the work needed to produce similar modification is proportional to the volume or mass of the material. Bond, during his 25 years of experience and research, reached to the conclusion that on the average the energy required for a known comminution is between the Rittinger's and Kick's laws (Bond, 1952). According to Bond's Third Theory, the work input is proportional to the new crack tip length produced during the particle breakage, and equals the work represented by the product minus that represented by the feed. The work input is given by the following equation: [111W = 1 Owi( FP FF(1)] where W is the The energy input to the mill, (kWh/st); F is the Feed size based on 80% passed, (µm); P is the Product size based on 80% passed, (µm); and W; is the Work Index, (kWh/st) Methods to determine the Bond Work Index Various methods have been proposed by Bond to determine the Work Index. There are some problems with the Standard Bond Grindability Test: •Requiring a special mill. •Requiring a relatively large amount of sample. •Being relatively time consuming and tedious. [ ] During the recent years, numerous attempts have been made to install faster and easier methods to determine the Work Index. A complete description of these methods is presented by Yap et al. (1982). Determining the Bond Work Index by a new size ball mill During the recent years, researchers at the University of Tehran determined the Bond Work Index, using the Anaconda simplified methods by its Denver 305- x 127-mm (12 x 5-in.) laboratory wet ball mill. The equation used to calculate the Work Index for the above mill containing 16.12 kg (35.5 lbs) of steel balls is as follow: [1 _ 1-1W;=0.54x [P [)(2)] In addition, an attempt has been made to determine the Bond Work Index similar to standard Bond ball mill procedure but by a different ball mill, smaller and in consequent portable and less laborious than the Bond ball mill. Method description Just as the standard Bond grindability test, this method is a locked-cycle dry grinding and screening process that is carried out until steady state condition is reached. Test equipment The test is conducted in a special ball mill, called the New Size Ball Mill (NSBM). The mill is scaled down with a coefficient of two-thirds from the standard Bond ball mill. It measures 200 x 200 mm (8- x 8-in.) inside (instead of 300 x 300 mm or 12 x 12 in. in Bond ball mill). Just like the Bond

Jan 1, 1995

By Ramon S. Pizarro, W. Joseph Schlitt

Introduction Trying to synthesize innovations occurring in gold processing is not simple. At present, this technology is exploding, brought about by the high price of gold. This price has compelled the mining industry, as a matter of survival, to produce gold. The frantic search for gold deposits continues, from glacial areas to tropical jungles. The search has been fruitful. New exploration techniques have helped to locate deposits. Some are amenable to conventional processing techniques. Others are in unusual settings that challenge the ingenuity of extractive metallurgists. Ore deposits containing traces of invisible gold, refractory mineralization, and impurities and deleterious materials that previously defied treatment are now being profitably processed. Metallurgists have developed various processing methods and equipment for gold ore treatment that reflect differences in ore characteristics, grade, mineralogy, reserves, topography, climate, and location. These processes can be categorized as: • For nonrefractory or readily leachable gold ores - vat leaching, heap leaching, and agitation leaching. • For ores containing free gold - gravity concentration methods, followed by mercury amalgamation, and • For refractory gold ores - flotation, roasting, pressure oxidation, followed by cyanidation leaching. In actual practice, a combination of these processes may be required. This is because gold mineralization, although usually simple, may be associated with other mineral species or with clays that make the ore difficult to treat. Furthermore, after the gold has been solubilized to form a cyanide complex, one of two important processes must be applied to recover the gold. These ore zinc cementation (Merrill-Crow process) and Carbon adsorption, desorption, and electrowinning. New techniques for leaching and recovery are also advancing. These will become the bases for the next generation of gold processing plants. Nonrefractory Gold Ores Nonrefractory ores with readily leachable gold are among the .easiest to treat. Even so, non-refractory ores from epithermal deposits are likely to contain too much clay. When these ores are damp, they may clog the screens and cone crushers. Depending on the type of process, scrubbing the crushed ore is very effective in overcoming clay problems. After dewatering, the slimes enter the cyanidation process and the thickener overflow is stripped of gold values and recycled. Semiautogenous (SAG) mills are also gaining popularity due to lower capital costs and ability to treat wet, sticky ores. If properly sized, the secondary crushing and screening plants, together with a secondary ball mill, can be eliminated. When the ore is of such a low grade that it can not support fine grinding, vat or heap 'teaching may be applicable. These systems are being actively applied in the US and elsewhere. Rapid construction and low capital costs are important features of these plants. Vat leaching is ideally suited for fast leaching ores and small tonnage deposits located in areas with harsh winters or wet tropical climates. For effective vat leaching, the ore is crushed to less than 25 mm (1 in.), agglomerated with lime or cement to control fines, and leached with sodium cyanide solution (upflow or downflow). The degree of size reduction is a tradeoff between increased costs of comminution and improved gold recovery from smaller particles. However, the size of material

Jan 11, 1984

By Y. Luo, S. S. Peng, C. T. Holland

Introduction A well-designed mining plan is one that minimizes cost and mining induced problems and maximizes recovery of the coal reserve. To minimize the negative effects on surface structures caused by ground subsidence due to underground mining activities, two kinds of practices are commonly employed by mine operators. One is the support technique, which leaves an area of coal unmined or partially mined beneath the structure to be supported. The other is the protection technique, which protects the structure when it is subsiding with the ground by taking advantage of the ground subsidence process (Luo et al., 1992). In designing the supporting technique, the only currently available method in the US is the one stated in the Pennsylvania Underground Coal Mining Rules (1985). However, the size of the support area designed by this method will be excessively large and result in a wasteful use of coal reserves when the overburden depth is more than 350 ft. A more scientific method is proposed in this paper for designing the support area. This method is based on a large amount of subsidence data collected throughout a number of major US coal fields and the authors' experience with monitoring and protecting surface structures that have been undermined by longwall panels. It considers the critical deformation indices that a structure can tolerate, and depth and height of underground extraction. A support area designed by this new method will not only assure the safety of the structure to be supported but will also maximize the extraction of the coal reserve. Design of support area using the Pennsylvania rules The Pennsylvania Underground Coal Mining Rules (1985) state that no underground mining activity should be permitted beneath structures where the overburden depth is less than 100 ft. An exception to these rules is provided for mine related openings that provide access to the surface for such as entries, shafts, boreholes and site-specific variances for entry development. If the overburden depth is greater than 100 ft, the mine operators should utilize either of the following: •Extraction techniques that provide for planned subsidence in a predictable and controlled manner, or •Support techniques designed to prevent subsidence damages to features identified as: •Public buildings and non-commercial structures [ ] customarily used by the public, including churches, schools and hospitals • Dwellings, cemeteries, municipal public service operations and municipal utilities in place on April 27,1966 •Impoundments and other water bodies with a storage capacity of 20 acre-ft or more • Aquifers, perennial streams and water bodies that serve as a significant source for a public water supply system, and •Coal refuse disposal areas. The support techniques require that a support area be left beneath the surface structure or feature to be protected (Fig.1). The support area should consist of coal pillars, with their size and pattern being such that their bearing strength is maximized. The total coal extraction within the support area should be limited to a maximum of 50%. The support area should be rectangular in shape and determined by projecting a line at a 15° angle from the surface to the coal seam with an offset distance of 15 ft from all sides of the structure. For a

Jan 1, 1994

By Patrick M. James

Many in mining believe that the industry is slipping behind in public image, in public mineral education and, particularly, in mining's basic right to do business. It is critical to turn things around - to change the rules of the game soon - before mining loses any chance of recovery. In the industry's experience, there is a technique of project development and operation that can help gain ground on public opinion. Only recently has this technique had a name, and that is sustainable development. What is sustainable development? Sustainable development is broadly accepted and supported by the international business community. And significant thought and work has been devoted to the subject. Research into the topic was initially confusing because the definitions of sustainable development are written in terms the writer, and his or her constituency understands. To some, it refers to environmental protection. To others, it is about social responsibility. And to the developer, it is about economic benefit. Slowly a theme emerged from the research that seemed to be part of all the definitions. That theme has three parts. First, the development, whatever or wherever it is, must contribute economic value to the developer and the stakeholders of the development. Stakeholders in this context include employees, neighbors, residents of the surrounding area, citizens of the country, suppliers and all others affected by the development. Next, sustainable development must contribute environmental value. This means that at a very minimum it must be neutral and not cause irreparable damage. And last, and what is new for miners and most other businessmen, it must contribute real social value to the stakeholders. How to use the concept? With that three-part theme as a working definition, the mining industry must find ways to use that “the challenge is for an individual company to translate what sustainable development means for them. There is no one model..." Donald W. Gentry is president and chief executive officer of Polymet Mining and former professor emeritus of mining engineering at Colorado School of Mines. He added: "While simple to state, the concept is anything but simple to implement. Present trends indicate ... one wins the sustainable development argument by staking out the high ground early..." In other words, start using sustainable development principles at the beginning of the project. New ways of doing business The mining industry's challenge is to implement sus¬tainable development universally around the world. Paternalism as a way of doing business is dead and a new model of doing business is emerging. All types of businesses are learning to deal with the stakeholders of new developments. Mining cannot be an exception to these trends. To understand how the international mining business came to sustainable development, it is useful to look back a few years. For many North American mining engineers, there were few opportunities for foreign assignments from the 1960s to the 1980s. American companies had been asked to leave many of the Latin American countries. Africa was going through a continent-wide upheaval. The Soviet Union and China had Eastern Eu-

Jan 1, 2000

By G. D. Naugle, L. C. Atkinson

Introduction Deep open-pit mines invariably affect the local and regional hydrologic systems. Pit dewatering, occurring during mining operations, puts an obvious hydrologic stress on these hydrologic systems. However, post-mining hydrologic impacts resulting from the pit refilling with groundwater following the cessation of mining activity can also be significant. The prediction of the rate at which the post-mining pit will fill with groundwater is a critical aspect of assessing the long-term hydrologic impacts. Numerical groundwater flow modeling provides a method for predicting the groundwater refilling rate of the pit. The rate at which pit "lakes" fill depends on several factors: •the rate and duration of pit dewatering; •the depth and size of the ultimate pit and •the pre-mining hydrologic regime. These factors can be incorporated into a detailed numerical groundwater flow model that can then be used to assess the effects of dewatering and post-mining recovery on the local and regional hydrologic systems. A sufficiently detailed, numerical groundwater model provides the oportunity to: •account for complex geology near the pit; •assess the impact of active pit dewatering and •predict the long-term impacts of post-mining groundwater flow into the pit. A detailed groundwater model incorporating these items has been developed and applied at an operating open-pit mine. Developed by Durbin and O'Brien (1987), the three-dimensional, finite-element, groundwater flow model was used to represent the hydrologic system of an approximately 253-km2 (98-sq mile) area surrounding the pit. Historical groundwater elevation data, stream flows and meteorologic, geologic and geophysical data were used to establish the dimensions and initial conditions for the model. Steady-state conditions, representing the pre-mining local and regional hydrologic systems, were simulated using the initial conditions incorporated into the groundwater model. The groundwater model was then utilized to simulate various dewatering programs, to predict the filling rate and the groundwater depth in the ultimate pit once mining activities are complete and to assess the long-term impacts on the regional groundwater flow system. Development of pit lake model Groundwater modeling efforts were completed in two phases. The first focused on pit dewatering activities, while the second phase concentrated on the post-mining effects on the hydrologic system. The final estimates of groundwater elevations calculated during the pit dewatering simulations were used in predicting the post-mining recovery of the hydrologic system. The groundwater model was also modified prior to the second phase to account for the volume of rock removed during mining activities. To account for the actual volume of rock mined, the geometry of the post-dewatering model grid was modified to approximate the final pit geometry. The depth and width of the ultimate pit were divided into eight idealized stages that represented significant changes in the bench geometries. These eight stages were then introduced sequentially into the model according to the predicted water elevations within the pit. In this way, changes in the volume and depth of water within the pit were accounted for through time. Once the ultimate pit geometry was accounted for in the model, it was necessary to assign new hydraulic characteristics to those parts of the model grid (elements) that represented excavated rock. The solution of the numerical model requires that finite hydraulic conductivity values be assigned to the portion of the groundwater model that represents excavated rock. Therefore, the calculated groundwater elevations differ, somewhat, between the edges and the center of the open pit. These model-calculated water elevations at the edge and in the middle of the open pit represent the elevation of water that would occur in the pit lake. To minimize the error in the estimated level of water within the pit lake, the hydraulic conductivity was increased to a value that would: •minimize the predicted difference between the groundwater elevations across the open pit and •produce a numerically stable solution. Specific storage is the hydrologic parameter that accounts for the water produced by compaction of the aquifer matrix. To predict the groundwater volume that would flow into the ultimate pit, this parameter was assigned a value equivalent to the compressibility of water. This value of specific storage reflects the post-mining groundwater storage occurring as an open body of water. Additionally, a specific yield of 1.0 was assigned to the pit elements to represent the 100% porosity of the open pit. In

Jan 1, 1994