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On my arrival in New York after the Russian examinations John D. Ryan, president of Anaconda, asked me what my plans were. I told him I wanted to be useful somehow in the World War that we had entered while I was away, but I did not know how. Ryan and Thayer told me that the new American Air Corps was gathering a group of executives, most of them young, to go to France as Army officers to help with the industrial side of our aviation activities. This would include construction of military airports, maintenance, and supply for our rapidly growing Air Corps. They thought I might be accepted in this group of about 30. It seemed to be just what I wanted, and I put in my application at once. My application was accepted about three weeks later and I was told to come to Washington to take physical and other tests and to meet others of the group. I liked them. They were a cross section of competent American manufacturers, bankers, automobile executives, and merchants. I was the only miner. I had many discussions in the days in Washington with members of our organization and others I knew who were going into war service. Attitudes were very different from those 55 years later. There was no discussion as to whether war was right or wrong. We were in it and must do all we could to help. At the same time there was a holdover from the days when, at least for the elite, war was high adventure. Even the most high-minded did not then realize that war is a horrible form of wholesale murder and, in 1917, it would have been treason to suggest this. I had two weeks before our industrial group was to start work which gave me time for a hurried trip to Berkeley where Dorothy and I announced our engagement.

Jan 1, 1976

By Osvaldo A. Bascur, J. P. Kennedy

Companies have installed powerful desktops for their personnel and instrumentation and control systems in their plants. These desktops are, in many cases, connected to a company intranet. Still, process engineers have to obtain their data manually. Engineers use process calculations from manually calculated averages from the plant; field operators collect equipment data on log sheets; and maintenance engineers collect critical information on log sheets. All of this information resides in individual machines. What are the strategic tools that will enable collaboration between them? How can operators, engineers, and managers access plant information to proactively increase their effectiveness in process troubleshooting, plant debottleneclung, aligning with business strategic and customer demand, satisfying environmental and safety regulations, developing capital projects, and implementing control strategies? A nonintrusive environment at the desktop is required to provide the process information to these workers. How can value be added to MS Windows technologies with the legacy systems at a minimum cost? There is no straight answer; however, there is a proven framework to leverage the large investments in new technologies. Objects and components are changing the way everyone relates to their computer and networks. Three examples will be described to demonstrate paradigm change. Today, process engineers can access data at the required level of detail for analysis of critical process variables at their desk or remotely. 1. Process Troubleshooting and Analysis: The engineer can develop an inferential calculation, trigger a process advisor to deter- mine the behavior of the units for the last day, and get the population of a set of batches (heats, blocks, rolls, sheets, etc). 2. Process Control and Optimization: The engineer can connect to a process simulator at steady state or a dynamic simulator of the process to evaluate new control strategies or process unit line-ups and use the process simulator as a process observer for fault diagnosis. 3. Collaboration with Accounting and Maintenance: The engineer can review reconciled data and the process mass balances, suggest maintenance cleaning or recalibration of flowmeters, and evaluate key control performance of all advanced controls.

Jan 1, 1999

By Eugene R. Palowitch, Albert W. Deurbrouck, Thomas H. Parrsons

INTRODUCTION Hydraulic concentrators utilize the physical characteristics-size, shape, and specific gravity-of particles suspended in a fluid medium to effect a concentration of desired quality. Fine-coal concentrators include generally those devices which are used to clean roughly 3/4, in. (9.5 mm) top-size coals. While this size designation is somewhat arbitrary, it is consistent with current practice although some fine- coal concentrators can treat coarser coals effectively. A variety of devices have been and are currently being used to wash fine coals and include: dense media cyclones, wet concentrating tables, hydrocyclones, jigs, spirals, launders, and upward-current classifiers. As launders and upward-current classifiers have been replaced by more efficient devices, they no longer warrant consideration as fine-coal washers. And as fine-coal jigs were addressed with the Baum jig under Part 1, Section 2, and heavy media cyclones and hydrocyclones were addressed under Part 2, Section 1, of this Chapter, only the wet concentrating table and the spiral concentrator will be discussed here. WET CONCENTRATING TABLES History The current wet concentrating table evolved from continuous belt concentrators which utilized a flowing film of water to recover heavy metals. In these devices the heavier concentrate moved up the inclined plane of the belt while the lighter impurities were washed away by water flowing down the belt. The natural outgrowth of the continuous belt concentrator was the vanner, a vibrating endless belt which was first developed about 1860 by Hartwig. Later models of the vanner, such as the Frue, incorporated oscillation in a plane transverse to the direction of the water current. Excellent descriptions of these washers are found in The Cleaning of Coal, by Chapman and Mott.80 Further development of tables included the circular stationary table (Linkenbach), the multiple round table (Anaconda-Evans), the bumping table (Campbell), and finally the differential-motion table which is now used extensively for cleaning fine coals.

Jan 1, 1991

By Robert F. Goss

OCAW appreciates the opportunity given to us by the sponsors of this Conference to present our position and policies on the issue of radiation hazards in mining. Our principal concern is the health impact that the mining of uranium has on our members. OCAW represents 1,500 underground uranium miners and more than 10,000 underground miners with 3,000 in the Rocky Mountain region. The U.S. Public Health Service has determined through mortality studies that the number one cause of death among uranium miners is lung cancer. It was also determined that exposure to radon daughters and mine dust correlates with the lung cancer experience of uranium miners. Data from the U.S. Mine Safety and Health Administration has also shown that not only uranium underground miners, but all underground miners, are exposed to radon daughters -- especially underground miners in the Rocky Mountain region. It is our position that any OCAW underground miner is at potential lung cancer risk. The dosages of radon daughters that our miners are exposed to are very many times the background levels of radon exposures in the communities where they live. We are also aware that cigarette smoking accelerates the onset of lung cancer; however, it has to be clear that the available scientific evidence shows that alpha radiation does initiate lung cancer and that cigarette smoke, as a recognized co-carcinogen, promotes cancer already initiated by radiation. It is true that cigarette smoke increases the risk of cancer significantly for miners exposed to radon, but nonsmoking miners have experienced lung cancer rates twice as high as the comparable members of the U.S. population. OCAW's position is that the occupational regulatory agencies should concentrate on the exposures that can be controlled; that is, occupational exposures rather than life-style exposures. Our Union has maintained a consistent posture in relation to carcinogens in the workplace -- that is, exposure to cancer-causing agents should be limited to the [lowest feasible level]. OCAW has interpreted lowest feasible level as the lower limit of detection of the collection and analytical method used to detect the carcinogen. Our posture is based on the available scientific information on carcinogenesis. We have asked the scientific community, many times, to provide us with safe levels of exposure to carcinogenic substances, including radon daughters. The answer has been: "We cannot determine levels of exposure low enough to assure that no cancer will occur." In short, there is not a "safe threshold" for any carcinogen. This statement does not come from one of the few so-called "pro-labor scientists," it comes from the National Cancer Institute and the National Institute for Occupational Safety and Health. I don't need to be a scientific sage, then, to conclude that the lowest level of exposure corresponds to the lowest risk of developing cancer. That is, then, our policy on exposure to carcinogens. It seems there has been an attempt to ignore the fact that lung cancer in uranium miners is the principal cause of death. Uranium miners are no exception from workers exposed to carcinogens. Our policy applies to them. Uranium miners should be exposed to the lowest feasible level of radon daughters and any decrease in the permissible exposure level is a decrease in their lung cancer risk. Accordingly, OCAW has petitioned the Department of Labor for a new permissible exposure limit to radon daughters in uranium mining, which lowers the current exposure standard from 4 Working Level Months (WLM) per year to 0.7 Working Level Months per year. We made our demand to the Department of Labor on April 20, 1980. We are still awaiting action from the Federal Government on our petition. OCAW is also very concerned with other important health impacts of uranium mining. We are concerned with a rate of disabling accidents and fatalities which is twice as high as the same rate in other underground mines, excluding coal. We are also concerned with the rate of respiratory disease fatalities among uranium miners which is almost four times the rate among a comparable U.S. population. We have expressed those concerns when the U.S. Senate proposed a Federal Compensation Act for uranium miners. That proposal, by Senator Dominici of New Mexico, found a quiet death in two Congressional sessions. In conclusion, our position on lung cancer induced by radon daughters is the same position we have taken with all other industrial carcinogens: The lower the exposure, the lower the risk. OCAW is demanding a drastic decrease of the permissible exposure limits. OCAW will never accept that a segment of our membership which mines uranium should take the lion's share of the risk while the uranium mining companies take all the benefits.

Jan 1, 1981

By Thomas O’Neil

Part 2 of this article continues with a discussion of features that distinguish contemporary investment analysis in mining from earlier methods. Part 1 traced the evolution of evaluation methodology from the traditional and discounted cash flow methods to the present. Quantitative investment models of the past 20 years played an important role in mine evaluation, but a new integrated approach to mine investment decision-making is now developing because of growing economic, technical, and political uncertainties. One of those uncertainties -inflation-was discussed in Part 1. Increased Risk At a time when mineral production technology and analytical evaluation tools have never been more sophisticated, it may seem anomalous to claim that mining ventures are riskier. Whereas technical risk has, indeed, been notably reduced, business risk the chance of a major loss due to market factors-has increased. This has been caused by two factors: the rapidly rising capital requirements of new projects and increasingly unpredictable future economic conditions. These difficulties are compounded by the long lead times of large, new mining projects. Rising Capital Requirements A $100 million investment is no longer considered to be a large mining project in most cases. For a large metal mining project, the total cost of the concentrator alone can exceed $100 million. Amoco Minerals' Thompson Creek molybdenum mine being developed in Idaho will cost over $350 million, for a relatively modest sized (23-kt/d or 25,000-stpd) ore operation with no extraordinary features. For a major grass roots copper operation producing blis¬ter or refined copper, the capital requirements are billions of dollars, not millions. Project costs are higher for two reasons. One, the projects themselves are larger; and two, the capital cost per unit of production is rising rapidly. Projects are larger because lower grade ores, yielding smaller margins, must be mined in greater volume-particularly in view of the high fixed costs of modern technology. When analysts contemplate any capital project, an important consideration is the minimum economic size of the facility. This is the smallest size that can be built and still be competitive on a total unit production cost basis. In the conventional copper smelting business, for example, a production rate of at least 91 kt/a (100,000 stpy) of blister copper is about the absolute minimum for reducing unit fixed costs to competitive levels. Clearly, many variables affect the minimum economic size of a mining facility, but with few exceptions that size continues to grow. Unit capital costs of new mining projects also are continuing to rise. The accompanying table shows that the capital costs in primary copper production have risen faster than the cost of living. Of particular note is the capital charge of over $1.70/kg (77411b) required for a new, fully integrated facility in 1981. Thus, the mid-1982 copper price would not even cover the cost of capital for a new copper producing complex without considering the direct operating costs for such a facility. The breath-taking costs of new mining projects are caused by factors such as increasingly complex technology, the rising social costs of creating attractive living and working conditions for employ¬ees, and the expenditures re¬quired for environmental protec¬tion. For whatever reasons, how¬ever, unit capital costs in mining continue to exhibit staggering growth rates. Future Economic Conditions Risk is not only a function of the size of the potential loss-an im¬portant measure of which is the capital intensity of mining dis¬cussed above-but also of the like¬lihood of incurring the loss. It is here that the ever increasing lead times of mining projects and the glaring limitations of economic forecasting are so important. In comparison to nearly all other in¬vestments, mining has more capi¬tal at stake for a longer period of time before commercial uncer¬tainty is resolved by distant eco¬nomic conditions that are subject to high forecasting errors. A sobering example of the pre¬ceding points occurred with the Lakeshore copper mine in south¬ern Arizona that Hecla Mining Co. undertook to finance and manage in 1969. The project was a joint venture of Hecla and the El Paso Natural Gas Co. and included two mines-one producing oxide ore and a second in the sulfide zone¬and a complex processing facility producing both electrowon cop¬per cathodes and copper precipi¬tates. The project originally was

Jan 12, 1982

By R. J. Pindred

Discussion by R.J. Pindred In his paper, Davis presents an overview of risk. He also introduces the Capital Asset Processing Model (CAPM) as a foundation for selecting the appropriate discount rate for a mining project. While applying portfolio theory is more defensible than the ad hoc adjustment of discount rates, the CAPM is not a panacea. CAPM shortcomings [The CAPM, as Davis stated, is expressed in the equation: ri=rf+pi4) where ri is the project discount rate rf is the risk free interest rate (3i is the project beta, and 0 is the market risk premium (rm - rf)] Application of the CAPM is more difficult than Davis indicates. Valuation is prospective, while the CAPM parameters are historical. Beta is determined from a regression analysis of historical data, while the beta needed for valuation is the expected beta. Betas are known to be unstable and the regressions that generate them often have low explanatory power. The difficulty of estimating a "project" beta must also be considered. Thus, the beta that is used in the CAPM will be based on the analyst's judgment. Like Cavender's discount rate, this judgment can lead to different project NPVs. Subjectivity in valuation cannot be avoided by a mechanical application of the CAPM. The risk-free rate, which Davis identifies as a short-term real rate of 4%, is also subject to scrutiny. A mining project is not a short-term investment and no single risk-free rate is appropriate for all of the cash flows. The hypothetical mine discussed in Cavender's paper is a six-year project. One might argue for the application of a risk-free rate from the Treasury yield curve at the duration of the project (in a bond-duration sense). This, too, is inappropriate. The risk-free rate should be matched to the timing of the cash flow. These rates can be determined by calculating the implied forward rates from the yield curve using a procedure known as "bootstrapping." It is likely that each of the project's cash flows would be discounted at a different rate. Commodity prices Davis criticizes the "ad hoc adjustment to the discount rate." Yet, in his discussion of the value of stochastic simulation, he suggests that the gold price be modeled as a "random walk, with or without a trend." This is essentially an arbitrary modeling of price risk. Consider that a liquid market in gold futures exists. The futures' price curve, which is closely related to the market's estimate of future spot gold prices, should be used to provide inputs to the model. This is especially true of a relatively short six-year project. Alternatively, as Davis correctly points out, a risk-averse investor can sell the commodity short to hedge price risk. Is it any more correct, in the portfolio sense, to account for price risk at all ?? References Cavender, B., 1992, "Determination of the optimum lifetime of a mining project using discounted cash flow and option pricing techniques," Mining Engineering, Vol. 44, No. 10, pp.1262-1268 Fabozzi, F.J., 1993, Bond Markets, Analysis and Strategies, Second Edition, Prentice Hall, Inc. Higgins, R.C., 1992, Analysis for Financial Management, Third Edition, Richard D. Irwin, Inc. Solnik, B., 1991, International Investments, Second Edition, Addison Wesley Reply by G.A. Davis Pindred discusses two issues related to my paper, the shortcomings of the Capital Asset Pricing Model (CAPM) and which commodity price values to use in the valuation exercise. Even though these topics are not directly related to the use or misuse of Monte Carlo simulation, they are important points to take into consideration in valuation exercises. Since I do not appear to have addressed these issues satisfactorily in my original paper, I will comment on each here. Pindred agrees with me that applying portfolio theory, and specifically the CAPM, to the selection of project discount rates is more defensible than ad hoc methods. But he then points out that the application of the CAPM to project valuation is more difficult that I indicate. It is true that the CAPM is a difficult tool for project valuation in general,. But the application of the CAPM to mining projects is one of the easiest I can think of. The biggest problem with using the CAPM for project valuation is coming up with an expected project beta. I suggest a project beta for gold projects of 0.45. The "true" value might be 0.35, 0.55 or whatever. Pindred correctly notes that the selection of the appropriate project beta is based

Jan 1, 1996

By Richard R. Klimpel

Part 2 of this article continues the discussion of a 10-year, multimillion dollar research and plant testing program on slurry rheology and grinding circuits. The first part of the article (ME, Dec. 1982) presented the basic concepts identified by the research and some laboratory test results. This section Illustrates typical Industrial scale test results and Identifies some industrial operating implications of controlling rheology by different methods. At least four controllable factors decide the rheological character of a slurry-slurry density or percent solids, particle size distribution, chemical environment, and slurry temperature. The second factor has two interrelated facets: the shape of the particle size distribution which controls packing behavior of the solids, and the fineness of the distribution. Finer particles increase interparticle forces and viscosity. As indicated in Part 1, during a given grinding test it is possible for all four factors to change. However, regardless of the particular settings of the four factors in a given test, if the resulting rheological character is either dilatant, pseudoplastic, or pseudoplastic with yield, the associated breakage rate is correlated with the current rheological character. It is obvious, for example, in batch grinding tests run at constant percent solids, that the second and third factors, where appropriate, are changing during grinding because size distributions are changing and the production of fresh surface area takes up unadsorbed chemical. Thus the corresponding rheological character change in batch tests with increasing grind time would be dilatant to pseudoplastic to pseudoplastic with yield. The degree to which this transformation occurs depends on the changing setting of the four factors over grinding time. In continuous grinding tests, one might logically expect that dramatic changes in any one of the four factors will be less likely to occur. It will be shown that continuous mill operations offer some unique opportunities to take advantage of possible rheological transformation by more direct operational control of the settings for the four factors. One extra observation noted in the rheological studies was the variability in the location and extent of region B (pseudoplastic behavior) for the various coals and ores tested. The location of region B was usually in the region of 45-55% solids by volume and was of the extent of 0-8%, or 2-11% with chemical addition. The corresponding increase in net production ranged from 0-10% in region B and 0-21% in region B&apos;. When region B is small or zero (no pseudoplastic character is exhibited) no increase in production will be observed and the use of chemicals is often marginal. There are several reasons for some materials exhibiting this quick transformation from dilatancy to high yield values often at surprisingly low percent solids by volume such as 30%. One condition identified was for materials containing high levels of viscosity-producing elements such as- carbonates or clays. A second condition documented was for materials that exhibited unusually fine primary fragment distributions (fine B1,; curves). This corresponds to materials having small y values of < 0.5. Such a slurry developed a yield value quickly during grinding because of the rapid buildup of fines. A related problem can be presented by materials exhibiting excessively coarse primary fragment distributions, y values > 0.9. Breakage of this type of material produces size distributions that give poor packing efficiencies even at long grind times, thus hindering the normal rheological transformation presented earlier. A final condition that can cause the occurrance of region B to be small or zero is when the media void volume filling of slurry is < 100%. Figures 5a and 5b use previously published data of this study to demonstrate the influence of solids loading and weight percent solids on the net production of taconite ore in a laboratory batch ball mill. In particular, Fig. 5a illustrates several trends not generally recognized until this study. They include these two: • Increased slurry density allows for increased solids loading before passing through the maximum in the net production curve where the fall-off is due to non first order breakage. • The use of rheology control chemicals such as GA-4272 allows this trend to be extended to higher slurry loadings with an increase in net production over any previous condition by keeping grinding first order. Figure 5b shows the same data as Fig. 5a plotted for constant weight loadings. These various figures of net production versus percent solids show the location and extent of the regions A, B, and C presented in Part 1 as a function of solids loading. It is obvious that the rheological transformation pattern described earlier does not hold true for slurry loadings corresponding to less than the void volume of the media, which is also a region of

Jan 1, 1983

By Robert F. Kaufmann

INTRODUCTION In response to the Uranium Mill Tailings Radiation Control Act (Public Law 95-604, Section 114c), a report has been prepared by the U.S. Environmental Protection Agency (1980) which describes the potential health and environmental effects associated with uranium mine wastes. The author was responsible for the geology, hydrology and aqueous pathways analysis portions of the report. All potential sources of contaminants, solids, liquids, and airborne emissions, which might cause adverse impacts upon the environment or health of nearby populations were considered. Wastes at both active and inactive uranium mines (Hans and O&apos;Connell, 1981) are described and assessed with respect to dose and health effects (Blanchard et al. 1981). The emphasis of this paper is on the liquid emissions and the aqueous pathway. Model uranium mines were developed for surface and underground mines. The model underground mine was placed in New Mexico while the model surface mine was in Wyoming. Development of a generic assessment was complicated due to the inherent variability of uranium mines. Also, most information available relates to uranium mills rather than to mines, or reported measurements were often influenced (biased) by wastes from a nearby mill and did not specifically reflect mine discharges. Consequently, the assessment presented is often based on conservatve (maximizing) assumptions which are identified as such, and which were based on knowledge from field investigations, available literature, and consultations inside and outside the U.S. Environmental Protection Agency. Maximum use was made of available data, supplemented with limited field studies in Colorado, Texas, New Mexico, and Wyoming. PREVIOUS INVESTIGATIONS Stream survey work in Colorado by the Water Pollution Control Commission and the U.S. Geological Survey (Moran, 1974; Wentz, 1974) mentioned a portion of the Uravan Mineral Belt and uranium mines therein, but the work did not emphasize uranium mines. Significant amounts of acidity and total trace metal concentrations were found in streams from 18 different mining areas. Dilution and chemical precipitation below mine drainages decreased concentration and increased the pH. Given enough time and distance, the streams recover naturally, but the accumulations of trace metals in the sediments increase. Field observations in 1971-72 of streams in most of Colorado indicated that approximately 724 km of streams in 25 different mining areas were adversely affected by mine drainage (Wentz, 1974). Discussions of the impacts of uranium mining on water quality or quantity are incidental in numerous impact statements and environmental reports prepared by industry and (or) the U.S. Nuclear Regulatory Commission as an integral part of licensing or relicensing uranium mills. Coverage on mining is usually minor as the principal focus is on milling impacts. The same is true for the recently prepared generic EIS on regulation of uranium milling (U.S. Nuclear Regulatory Commission, 1979). Radiochemical assessment of surface and groundwater in uranium mining districts of New Mexico is done by self-monitoring programs associated with NPDES permits. Also, radiochemical assessment studies have been funded by the New Mexico Environmental Improvement Division and U.S. Environmental Protection Agency. Unfortunately, there has been no concerted effort to prepare broad assessments of the cumulative impacts of mining and milling in areas of interest in Texas, New Mexico, and Wyoming. Critical review and synthesis of these types of data can produce rather useful information. For example, the publication "Water Quality Impacts of Uranium Mining and Milling Activities in the Grants Mineral Belt, New Mexico" (U.S. Environmental Protection Agency, 1975) addresses the groundwater and surface water changes as the result of extensive uranium mining and milling production in a relatively confined area. A more recent report by the Wyoming Department of Environmental Quality summarizes 16 years of aqueous radium and uranium data. The study reports significant amounts of radium-226 and uranium were present in surface water in the Shirley Basin as a result of inadequate mine water treatment (Harp, 1978). In Texas, surface water and groundwater monitoring conducted by industry, as well as by State and Federal agencies, reveals little or no change of chemical quality attributable to uranium mining and milling (Gesell, 1977; Kallus, 1976). This conclusion is based on 586 samples collected from 198 stations over a period of 39 years, but primarily from 1961 to 1975. The State monitoring program by several agencies is continuing, but data reporting and interpretation is lagging. One survey (Itin, 1975) showed that none of the mine water from the 10 lakes that were sampled was suitable for human use. The lakes were also unsuitable for irrigation due to mineralization of the water by sulfate, chloride, and TDS. One of the 10 was suitable for stock watering.

Jan 1, 1981

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

10.1 I~RODUC1&apos;ION In order to understand the effects of surface subsidence due to underground mining on the ground water hydrology, a brief introduction to the rate, direction, and general pattern of ground water movement in aquifers is necessary. Ground water movement in an aquifer is controlled by aquifer lithology, structure, permeability, water levels, topographic set¬ting, and pumping patterns (Sgambat et al., 1980). Fig. 10.1 shows an idealized flow pattern of ground water in a hilly terrain (Coe and Stowe, 1984). The hill is made up of a series of interbedded permeable aquifers (mostly sandstone) and confining units (mostly shale). Since sandstone is porous and permeable, ground water moves through it downward and laterally until it reaches the interbeddings of the confining unit immediately below. The ground water then moves laterally along the interbeddings, be¬cause the confining unit is less permeable, and comes out as springs on the surface at the intersection of the aquifer and the confining unit. If there is a ground water table in each aquifer, it is a perched water table (e.g., right side of Fig 10.1). On the other hand, if the ground water recharge is larger than can be drained out, the interbedded permeable aquifers and confining units will be filled with water and a ground water table will exist only at the uppermost permeable aquifer. This is the semi-perched water table (e.g., left side of Fig. 10.1). There are three ground water flow systems: shallow, interme¬diate, and deep. Shallow flow systems occur near the land surface and the surface drainage system. The flow depth ranges from a few to several hundreds of feet with a resident time of a few weeks to a few years after entry, depending on the permeability of the rock strata. In a shallow flow system, the drainage coincides with sur¬face run-off water divides. Intermediate flow systems occur in minor hills some distance away from surface drainage. The travel distance ranges from several hundreds to several thousands of feet, the travel time from several months to tens of years. These systems are less dependent on seasonal recharge or discharge. In deep flow systems, the ground water travels from thousands of feet to miles, and transfers between surface basins. The travel time ranges from several tens to several hundreds of years. The systems generally receive recharge over a large area on major mountains and dis¬charge into major stream systems. 10.2 CASE STUDIES There are two ways of monitoring water wells for studying the effects of dewatering due to underground mining, i.e., water level and piezometric level. The water level is the water elevation in a water well that has penetrated two or more aquifers. Thus it indi¬cates the combined flow of all the aquifers that the water well intersects during the time of measurement. The piezometric level is the level to which water rises in a piezometer cased to a specific aquifer, and coincides with the pressure level of water in that aquifer. Therefore piezometric level reflects the water condition in that particular aquifer to which it is confined, while the water level represents the total yield of a water well without regard to the source of water. For instance, if a water well intersects two aqui¬fers, and if the lower aquifer is affected by undermining and goes dry while the upper aquifer does not, the water level in the well will drop, but may never go dry. Under this condition, the data obtained by monitoring the water level do not indicate the true behavior of the individual aquifers, especially if they are located at widely separated heights. 10.2.1 WATER LEVEL EFFECTS The effects of underground mining on ground water flow have been the subject of many investigations (Booth, 1984; Coe and Stowe, 1984; Owili-Eger, 1983; Rauch et al., 1984). However, the problems are very complicated, as have been exemplified by the conflicting conclusions reached by various investigators in var¬ious case studies. The most dramatic examples were described by Coe and Stowe (1984). In the first case (Fig. 10.2), the coal seam was 5 ft thick with cover from 400 to 800 ft deep. The majority of water supplies came from a sandstone aquifer in most ridge tops. When the face of the Panel No. 3 passed under it, dug well WW-1, which was approximately 670 ft above the coal, went dry, while the water level in Borehole B- 1, which was located at the center of Panel No. 4 with the bottom reaching the coal seam, also dropped about 160 ft. In Panel No. 5, the developed spring DS-2, which originated from the surface of the sandstone/shale contact and was approximately 750 ft above the coal seam, stopped when the face passed under it. The water level in the dug well W-2, which was dug into the shale about 700 ft above the coal seam dropped about 22 ft, but the well did not go dry. Therefore, nearly every spring or well located above or near the active panels was affected by longwall mining. In the second mine (Fig. 10.3), the seam was 5 ft thick with cover from 200 to 400 ft thick. Dug well DW-2, the bottom of which was 35 ft deep and approximately 330 ft above the coal seam in a perched sandstone aquifer, went dry right after the longwall passed under it. Drilled well W-3, in the sandstone aquifer, about 220 ft above the coal seam, experienced a 75-ft drop in water level when the face was under it, but showed some re¬covery after the face had passed under it. Ponds P- I and P-2 were developed in shale bedrock and were spring fed. They were not affected by longwall mining. Pond P-3 was developed in the sand¬stone aquifer, and its water table showed some decline when the face passed under it. Spring ST-1, which originated from Pond P-1 and was fed from several tributary valleys, was affected by longwall mining, probably because it passed across the panel edges where tension cracks drained off the water. Although the overbur¬den in case 2 was about half of that in case 1, the majority of water resources were not affected by longwall mining. Coe and Stowe (1984) attributed this discrepancy to: (1) a higher percentage of clays, shales, and claystones, resulting in a lesser degree of inter¬connected fracturing, and (2) a less rugged topography resulting in a less localized ground water flow system. Therefore, in order to assess the effects of underground mining, the premining ground water flow patterns, topographical, lithological, and meterological conditions must be known. Booth (1984) also monitored the ground water effects of a longwall mine (Fig. 10.4) in a seam 5 ft thick and 300 to 800 ft deep. There were three major aquifers: Morgantown sandstone approximately 30 ft thick and 630 ft above the coal; Saltsburg sandstone of up to 50 ft in thickness and 520 ft above the coal; and

Jan 1, 1992

By A. Simon, J. Weyer, J. Brune, V. Zhiteneva

"Reiche Zeche is a lead-zinc-silver deposit in Freiberg, Germany, that was mined for over 800 years. Since 1969, Reiche Zeche has been a teaching and research mine with numerous adits for dewatering large amounts of inflowing water. Infrequent metal water quality monitoring in the 1980-1990s and flooding of the mine in 2002 led to a lack of relevant data. New measurements were taken at different level adits and ore veins for comparison to historical values, revealing pH levels as low as 1.95, conductivity as high as 22000 µS/cm, and an overall trend towards less acidic waters. INTRODUCTION Reiche Zeche is part of the Himmelfahrt Fundgrube, a former iron ore mine in Freiberg, Germany, in the Erzgebirge mining region of Saxony. The polymetallic sulfide ore found in Freiberg was mined from 1168 to 1969, and though it is no longer producing ore, Reiche Zeche remains as a teaching and research mine for the Technische Universität Bergakademie Freiberg. The most important ore minerals mined were sphalerite, galena, pyrite, chalcopyrite, and silver, and important gangue minerals include calcite, fluorite, barite, siderite, and quartz (Zänker et al 2002). Over one thousand ore loads were mined from this region during operation, which was ceased in 1969 after the lower part of the 700 meter depth was flooded, leaving only the top 200 meters under the surface open for research and teaching. Mined materials from this region can be separated into high grade ore and waste rock. High grade ore, composed of pyrite, sphalerite, galena, chalcopyrite, and arsenopyrite, was brought to surface for smelting during mine operation, and spaces were backfilled with waste rock, which is the primary site of sulfide oxidation and acid mine drainage (Haubrich and Tichomirowa 2002). The waste rock includes remnants of ore from ore loss, veins which were too small to mine, or ore minerals within host rock. Waste rock is primarily composed of biotite gneiss, with sulfides including jarosite, gypsum, minor anglesite, scorodite, and sulfur (Haubrich and Tichomirowa 2002). When weathered, the gneiss forms clay minerals such as illite and kaolinite (Haubrich and Tichomirowa 2002)."

Jan 1, 2016

By T. R. Ley

"INTRODUCTION Brutally honest mining executives will tell you that “safety is job #1 but.. dollars are also extremely important”. When economic times are tough, as they are now in mining, it can feel like you need to choose between safety and cost control. Leading edge companies know that it is excellence at both that provides the best performance. FLUID POWER SYSTEMS Fluid power systems are ubiquitous in mining – longwall systems, continuous miners, drills, haul trucks and excavators, etc. And fluid power systems are getting more powerful and using higher pressures to move greater volumes and achieve higher productivity. Average operating pressures in mobile equipment has been rising to 450 bar (6,500 psi) over the last decade. With increased system pressure comes ever increasing levels of stored energy. Stored energy that, if released in an uncontrolled fashion, can do tremendous harm to people and things.Fluid power systems are composed of pipe, flexible hoses, pumps, valves and actuators. They are filled with fluid (mineral oil or fire-resistant) to transmit power to where it’s needed for performing work. With time and use fittings can loosen, component performance diminishes and hoses wear and tear. In hostile environments, like coal mining, hydraulic hoses can fail in as little as 2 years according to DGUV (umbrella association of German Accident Insurers), maybe sooner in some particularly challenging environments. FLUID POWER SYSTEM SAFETY While appearing simple and benign on the surface, fluid power systems, particularly such systems on mobile equipment, have many potential hazards. Hazards that most workers aren’t aware of – particularly young, new hires. And those hazards that can quickly turn deadly. During the past couple of years, including at last year’s SME Conference, I talked to a wide variety of mining people about high pressure fluid injection injuries. Some people, often with little or no field experience, never heard of it. But, people with mine operations / maintenance experience were almost always aware of such injuries.Some had even seen it happen right in front of them. But most claimed it didn’t happen all that often. It just wasn’t seen as a problem!"

Jan 1, 2016

By Klaus Rieker

"Since the mid-19th century the River Emscher in the German Ruhr District has been used for disposing of wastewater. In the early 1990s, it was decided to replace the existing open wastewater system with a sewer system and to restore the River Emscher to its natural state. The major Emscher conversion project is divided into a large number of individual schemes, with construction phase 30 being the largest. By applying the pipe jacking method, interlinking tunnels with section lengths in excess of 1,100 m are constructed. USAGE OF THE EMSCHER OVER TIME The 85-km long river Emscher flows right through the Ruhr District. It has its source to the east of Dortmund and flows into the river Rhine near Dinslaken (Figure 1).The Emscher has been used to dispose of wastewater since industrialization began about 200 years ago. Intensive coal mining over the years led to considerable subsidence causing stagnant sewer areas which in turn led to typhus and other deceases. Thus in 1899, the Emschergenossenschaft was founded to create an efficient interceptor system for a catchment area of 865 km² through lining, raising and Emschergenossenschaft decided to free the waterways of the Emscher system from sewage and to improve them in ecological terms by 2020. A budget of 4.5 billion euros was earmarked for this purpose.The construction of sewer conduits began in order to no longer transfer the wastewater into the Emscher. The currently ongoing construction of the “Emscher Interceptor” - the main sewer - represents the heart of the Emscher conversion project. The interceptor will run along the river over a length of 51 km. Three pumping stations are required to confine the sewer’s depth to elevations between 15 and a maximum of 40 m below the surface. Since early 2012, the Emscher has been free from wastewater east of the Dortmund-Deusen clarification plant. There are two further clarification plants situated along the Emscher, which treat the still contaminated river water down-stream. Once the Emscher Interceptor is in operation, they will receive wastewater solely from the tunnel system."

Jan 1, 2016

By Z. Sample, S. Kaushik

"INTRODUCTION Industries around the globe have been making significant advancements in recent years due to innovations in automation technology. With the implementation of modern automation technology, professionals in process automation or plant operations can expect significant and exciting change, and the mining and mineral processing industries are no exception. From 2000 to 2010, the International Council on Mining & Metals (ICMM) reported that an increased output and increased value of most metals has resulted in a rise in global metal and mineral process industries from $214 billion to $644 billion. [1] Since most productivity increases in mining and mineral process industries have been achieved through the ability to process lower grade ores with improved efficiency and larger equipment, these industries have begun to adopt advanced automation and control technologies. Despite a growth slow-down, the mining and mineral processing industries are poised for continued and sustainable growth and development due to an increasing demand for energy minerals and emerging technologies requiring minerals and metals previously not demand. However, as these industries becomes more automated, new challenges are arising on the horizon. DEMOGRAPHIC TIME BOMB John Payne from BP is credited with first using the term Demographic Time Bomb to describe the upcoming crisis in the process industry in finding experienced operations personnel. From a survey of 912 mining professionals, Stratum International found that a mere 0.4% of senior level candidates were between the ages of 25 and 34 and a similar proportion aged over 75 years. This indicates that there is both a deficiency of young skilled workers and a skilled workforce that is retiring earlier than the previous generation. Combined with data published by the US Department of Labor stating that the median age of the mining workforce is approaching 50 years old, this implies that these skilled mining and mineral process professionals are an aging population with irreplaceable knowledge and experience that are expected to retire in the next 5-10 years. While it is projected that the labor pool will be sufficient to fill open positions, the real question is if this young workforce will have necessary skill sets and experiences."

Jan 1, 2015

By R. L. Taylor, S. D. Abbott

"INTRODUCTION Mines are filled with a mixture of gases, some of which are toxic, flammable, explosive. Others can be early reporters of potential problems. It has previously been shown that analysis by gas chromatograph can the most accurate conventional analytical tool for common mine gas mixtures when used correctly. There are some other detection methods (e.g hand held sensors), but they do not detect many situations which are immediately dangerous to life or health (IDLH). Examples include explosive hydrogen mixtures, toxic carbon dioxide levels, and some explosive mixtures of methane with oxygen. (Reference our previous paper). There are several reasons for mines to consider onsite GC analysis. (1) Timeliness of results. Samples can be taken and analyzed immediately. Samples do not have to be shipped to a lab and wait for results. (2) Ability to run greater number of samples a lower for the same cost. This allows mine to increase the number of samples taken to be more informed about the gases/conditions of the mine without the worry of financial constraints. (3) Fewer sample media needed, and there is no wear and tear or cost from shipping them offsite. When samples are shipped offsite, those sample media cannot be used to take until the lab returns them when the analysis is done onsite, the sample media can be used again almost immediately. However there are several challenges associated with using GC to analyze mine gas samples onsite. To date there has been a clash of the requirements for a successful analysis, the capabilities of the technology, and the practicalities of implementation. The purpose of this paper is to discuss some of the difficulties in getting accurate reliable data and doing so in an economic way. This whole measurement activity is a cost center for a mine, so there are economic concerns for the mine to address. These concerns are mainly staffing costs, capital costs, maintenance, and consumables costs. A properly staffed and equipped analysis lab, including burdened overhead, costs about $500K per year. At a sampling rate of 10 per day, this is about $140 per measurement for a proper onsite lab."

Jan 1, 2016

By J. B. Fletcher

Miami Copper Company has used conveyor belts underground since 1947, At that time, the draw level for a portion of the orebody was only 25 feet above the haulage level To get the main haulage drifts as far away as practical from under the stoping area, 301, belt conveyors were used for lateral transfer, If the haulage drift was kept in good ground., a satisfactory draw of these stopes could be effected in that repairs to the conveyor drifts could be accomplished while drawing ore from the stopes. Some 7,400,000 tons of ore were mined with the 301d belt conveyor at a belt cost of approximately $0,004 per ton of ore. When the mining plan for a new section of the mine was laid. out in 1953, it was decided to use belts to collect the ore from the higher draw levels, and. move the ore laterally to centrally located concrete ore passes for transfer to the main haulage level (See Cut #1), The final plans used belt conveyors too collect from 3 stopes: on the 570 level, 2 stopes on the 720 level, and one stope on the 900 level. The six stopes to be mined in this manner would produce about 8 million tons of ore, and the sequence of mining was such that not more than four 400 feet conveyors would be in use at one time, Four conveyors using 36 inch, 4 ply belting were designed by the mine engineering office and fabricated at the mine shop, so that no piece was longer than 10 feet, and. the heaviest piece was the 25 horsepower motor, The head drive section, and the tail-pulley-takeup section were of bolted construction so they could be broken down for easing moving, Due to the size of the mine openings, the belt had to be transported into the mine in 400 feet rolls, requiring two splices for an installed belt. A torque arm speed reducer was used with a V -belt drive which made it possible to change the speed of the conveyor by changing the diameter of the V-belt pulley (See Cut #2).

Jan 1, 1958

By Donald C. Wotring, Douglas A. Gabriel, Michael G. Vitale

"The Northeast Ohio Regional Sewer District (NEORSD) recently completed the 24-foot diameter Euclid Creek Tunnel using a steel-fiber reinforced concrete segmental lining. During construction HMM, the McNally/Kiewit JV, CSI/Hanson JV, and the NEORSD performed a test program to evaluate replacement of steel fiber with synthetic fiber. A section of synthetic-fiber-reinforced cast-in-situ lining was also placed as part of the permanent works for the starter and tail tunnels. Laboratory and field testing were performed on two complete segmental rings reinforced solely with synthetic fibers. This paper summarizes the testing program, construction, and results of the synthetic-fiber-reinforced CIP lining and concrete segments. INTRODUCTION The principal feature of the Euclid Creek Tunnel (ECT) project is a CSO storage tunnel that is one component of the Northeast Ohio Regional Sewer District’s (NEORSD) Easterly CSO storage and conveyance system. The tunnel, shafts, and appurtenant facilities will be used for the collection and storage of CSOs from the northeastern part of Cleveland, Ohio and subsequent conveyance to the NEORSD’s Easterly Wastewater Treatment Plant for treatment. The ECT project includes five drop shafts that extend as deep as 220 feet below the ground surface. The ECT alignment and shaft locations are shown in Figure 1. Traditionally in northeast Ohio, tunnel excavations were supported with a two-pass system using steel ribs and timber lagging, followed by cast-in-place concrete. Control of rock overbreak proved difficult with large-diameter two-pass operation, and invert degradation led to train derailments. To mitigate these risks, the ECT was supported using one-pass, precast, bolted, gasketed concrete segments reinforced with steel fiber alone. Tunneling operations were very successful and experienced no significant delays or claims. As a result of the positive ECT experience with steel fiber-reinforced segments, Hatch Mott MacDonald, LLC (HMM), McNally/Kiewit (MK) joint venture, CSI/Hanson (CSI/H) joint venture, and the NEORSD performed a test program to evaluate replacement of steel fibers with synthetic fibers in segmental rings. This paper presents the laboratory and field testing program, construction, and results of the synthetic-fiber-reinforced cast-in-place (CIP) lining and concrete segments."

Jan 1, 2016

By K. Lagnese, C. Robinson, S. Muddasani

"Mine water generated from underground coal mining operations contains both dissolved and particulate solids. Dissolved solids primarily consist of sodium, calcium, magnesium, potassium, chlorides and sulfates. When discharged to a receiving stream without treatment, these constituents create a potentially toxic environment for aquatic life. A case study is presented to discuss how a centralized treatment plant treats mine water from six locations to meet discharge limitations for chlorides. Located in West Virginia, this facility consistently achieves less than the NPDES permit limitation of 218 mg/L chlorides in the discharge while generating almost zero liquid waste. The dissolved solids concentration in the influent ranges between 5,000 and 10,000 mg/L, with chloride concentrations of 1,000 to 2,000 mg/L and sulfate concentrations of 2,000 to 6,000 mg/L. The mine water is treated using advanced treatment technology to produce clean water for discharge or reuse. The treatment process comprises aeration, softening, filtration, reverse osmosis, RO reject softening, evaporation, crystallization, final effluent remineralization and sludge dewatering. The solid waste generated in the treatment process is landfilled on site. The leachate is sent back to the facility’s thermal treatment process. Because no liquid waste leaves the property, this plant is termed a “Zero Liquid Waste” (ZLW) facility. INTRODUCTION The advanced water treatment system was built to meet a new 218 mg/l regulatory limitation imposed by the West Virginia Department of Environmental Protection for chlorides discharged to surface waters. The system is designed to treat a maximum flow of approximately 5 million gallons per day (795 m3/h) of mine water. The treatment process utilizes reverse osmosis (RO) membrane treatment to achieve the effluent criteria, and evaporation and crystallization technology to manage the RO concentrate from the water treatment process. As a result, the system produces clean water for discharge while generating zero liquid waste. The effluent water can be used for industrial purposes or discharged to a receiving stream. A Water Impact Index (WIIx) study indicated that discharge of this high-quality water into the Monongahela River watershed may improve the water quality for downstream water users. The waste produced in the treatment process, including softening sludge and salt, is disposed of in an on-site landfill."

Jan 1, 2015

By Takayoshi Otsuka, Taher Elsamni, Ashraf Bou-Krisha, Takaya Takimoto

"INTRODUCTION Tunnel is a closed space structure and it is necessary to ensure the sufficient time and route for safe evacuation for the tunnel users in case of a tunnel fire. In order to do so, damages and collapse of the tunnel linings and the secondary disasters can be avoided. Furthermore, the road tunnel is one of important infrastructural facilities and stopping its function for a long time for a tunnel fire could cause a great economic loss. Therefore, the tunnel structure shall be the one which can be repaired as soon as possible, and the fire protection methods need to be selected among those which enable the service to start as early as possible. The necessity of the fire protection works is evaluated for the tunnel structure with the consideration of the expensive tunnel construction cost and the economic effects of the tunnel service. Naturally, the fire protection works, which can protect the tunnel linings from the tunnel fire with small repairs, are inevitable. Design engineers of the fire protection works estimate the scale of fire from the degree of damages on the linings and its repair cost, the effect of fire extinguishers placed inside of the tunnel, the secondary disasters in accordance with the tunnel lining damages if any, and the ratio of fire protection cost calculated from tunnel operation cost. Based on this information, the engineers design the fire protection works according to the permissible limits Among the design cases of the tunnel fire protection in Japan, as shown in Table 1, there is no cases that used the fire temperature-time curve to design the fire protection works for large fire caused by vehicles catching fire. However, the most of cases used 30-60 minutes for fire duration based on the entry prohibition of dangerous goods carrying vehicles or the effect of early fire fighting. This paper will study the variation between different methods used for fire protection in road tunnel in Japan to evaluate them and try to found some recommendations. Especial l care will be attention to Polypropylenes (PP) which is starting to be applied lately."

Jan 1, 2016

By P. Knudsen, S. Rosenthal

INTRODUCTION In 2010, Montana Tech was gifted 65 acres of land immediately west of the campus. The parcel of land included an old silver mine called the Orphan Boy Mine. The Orphan Boy and its sister, Orphan Girl, were developed about1895 and operated until 1956. In addition to the gift of land, a significant industry gift was obtained to develop a first class training facility, which we call the Underground Mine Education Center (UMEC). The UMEC has been funded primarily by industry through donations of dollars, equipment, and supplies. The UMEC is a unique hands-on, educational environment for today’s students who are being trained to find, develop and process the worlds' natural resources. The center complements courses in mining engineering, geological engineering, environmental engineering, metallurgical engineering and occupational safety and health. It also serves as a as research facility utilized by students and faculty, and other organizations interested in doing research a hard rock underground mine. LOCATION The Orphan Boy and Orphan Girl Mine are located approximately 700 feet (ft.) apart and the Orphan Girl Mine is part of the World Museum of Mining. Underground they are connected on various levels and all levels below the 100 ft. level are currently flooded. The Orphan Boy mine was 804 ft. deep and Orphan Girl was 3200 ft. deep. The surface locations of the two mines and some of the underground workings on the 100 ft. level are shown on Figure 1 (2015). CONSTRUCTION Land ownership was gifted from BP/ARCO in 2010 and planning started immediately. In June of 2011, site development for the decline to the 100 ft. Level of the Orphan Boy began with a satisfactory portal site developed (Figure 2). We hoped to be able to start driving the decline in July. However, the portal was sited approximately 125 ft. away from a natural gas line that bisected the Montana Tech campus. While outside the right of way for the line, Northwestern Energy requested that no blasting be performed near the gas line until such time as the gas line was abandoned in the fall so decline development was put on hold. The gas line was officially abandoned in January of 2012.

Jan 1, 2019

By J. Castaneda, J. Schmidt

"INTRODUCTION Chino is an open-pit copper mining complex faced with intricate material routing requirements. Historically, this was challenging at the mine, in part because the shovel dig plan, haul plan, and dump plan were prepared completely separately. The mine now uses an integrated workflow for faster evaluation of schedules, which facilitates preparing balanced cutoff grade-to-equipment requirements, supports better mineral production, and improves schedule viability, while considering operational implications. Located in the state of New Mexico, Chino is one of nine Freeport-McMoRan mining operations in North America. This presentation discusses the benefits of implementing an integrated approach to tactical mine planning, and shares practical examples that were learned as Chino implemented an integrated tactical planning method. This includes improvements in completion time for preparing a schedule, minimal risk from manipulating Excel macros, data integration required to make short-term planning decisions, and better understanding of the mine plan’s effect in relation to the business’s key performance indicators. Ultimately, the integrated approach to tactical mine planning at Chino has led to production forecasts which are more achievable and reduced planning and execution variance. PROJECT OVERVIEW Chino mine is a porphyry copper deposit wholly owned by Freeport-McMoRan, the largest publicly traded copper producer in the world. Chino mine is located in Grant County, New Mexico, a short distance east of Silver City. Ore material mined from the pit today includes oxides, primary sulfides, and secondary sulfides. The mineralized material is sent to a mill, with a rated throughput of 42,000 tons per day, or to ROM leach pads with variable dumping capacities. At Chino, the short-range objective is to maximize total copper pounds sent to the mill, but no fixed cutoff grade accomplishes this for all periods. Instead, 12 different material classifications, including four sulfide classes that can feed either the mill or other destinations, and 38 cutoff bins, are aligned with the needs from operations, leaching, and the concentrator to create a schedule. Cutoffs vary from one week to another to cope with a highly intermittent ore body."

Jan 1, 2018