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By Donald R. East

For many years there has been increasing pressure on mining companies, not without some justification, to design tailings storage facilities which guarantee some degree of protection against process solutions entering local groundwater and surface water flows. A fairly straightforward approach to achieving this has been to design conventional water-retaining type structures, and this has often been reinforced by the need to store large volumes of process water and seasonal run-off within the tailings facility. The conventional design uses naturally occurring geological barriers and a conventional embankment with possibly a grout curtain, or artificial barriers around the entire facility, using either imported low permeability soils, admixtures, or synthetic liners. While these approaches may achieve the primary objective of minimizing seepage, various other factors should be considered. Firstly, by depositing the tailings slurry in a sub-aqueous manner, the tailings, especially if relatively fine grained, will settle to very low densities and little or no on-going consolidation may occur over the operating life of the mine, There have been many cases where measured pore pressures within the tailings mass equal the total overburden stress, i.e., no consolidation and no strength, many years after decommissioning.

Jan 1, 1998

By Phillip Crowson

The annual surveys of corporate exploration expenditures carried out by Metals Economics Group (MEG) show that estimated total spending rose to a record $13.2 billion in 2008 ($14.4 billion including expenditure on uranium exploration). This compares with the previous peak of $5.2 billion in 1997. MEG does not adjust its figures for the effects of inflation but recognizes that the substantial increase in budgets of recent years has not been reflected in a proportionate rise in actual activity on the ground (MEG, 2009). In recent years, costs were boosted by rising prices of fuel and other inputs, and acute shortages of qualified staff and drilling rigs. The inflation rate for exploration substantially exceeded increases in general price levels as measured by national indices. Just how much exploration costs in-creased, however, is probably not widely appreciated. Figure 1 compares the MEG estimates of total annual dollar spending in Australia, Canada and the United States combined with their drilling activity, measured in meters drilled, as reported by national agencies. These three countries accounted for 39 percent of MEG?s estimate of total spending on minerals exploration in 2007 and 40 percent in 2008. Expenditure is shown both in money terms, as reported by MEG, and adjusted to real terms, with the U.S. implicit gross domestic product (GDP) price deflator. The comparison is of index numbers based on 2000 as 100. Comparisons for the individual countries are shown in Fig. 2 to 4. National figures for drilling activity and expenditures cover some minerals like iron ore, coal and industrial minerals (29 percent of Australian spending in 2007 to 2008 and 10 percent of Canadian in 2008) that are not included in the MEG estimates of spending. Such exploration has risen more than the total since 1997 so that the index of drilling activity in the figure may be overstated for recent years on a like for like basis with the MEG data. MEG only includes spending on uranium exploration in recent years whereas the national agencies include it throughout. Figures 5 and 6 com-pare national and MEG estimates for Australia and Canada. No comparable figures are available for the United States. The national estimates have been converted to U.S. dollars at the appropriate annual average exchange rates, and the Australian annual totals have been summed from the published quarterly data. Not only is spending cyclical, but so are unit costs. Australian and Canadian data show that total exploration spending per meter drilled rose by 2.5 to 3 times between 2001 and 2008. These rises may partly reflect increasingly complex drilling methods, and an increased share of assaying and testing in total exploration costs. In Canada, however, diamond drilling accounted for a lower footage of the total in 2005-2007 than in the preceding few years, and spending on exploration, as opposed to deposit appraisal, made up a higher proportion of the total. Figure 7 compares index numbers of exploration spending per meter drilled in U.S. dollars for Australia and Canada. These figures raise a number of questions. First, is the relationship between total expenditure and meters drilled in the three countries for which data are available typical of the global experience? Secondly, has there been a significant change in the make-up of exploration spending in recent years? The Canadian data do not suggest that there has been, but is Canada typical? Has drilling become less important relative to other exploration techniques, or is it just as important as ever in confirming the existence of viable ore deposits?

Jan 1, 2009

By H. E. Epstein, B. K. Parekh, W. M. Goldberger

Comminution is the single most expensive operation in mineral processing. It consumes about 50% of the energy required for mineral extraction (Agar, 1976). Current comminution technology is both energy-intensive and inefficient. A novel noncontact comminution process concept was developed in this study, whereby selective lib¬eration of minerals from an ore could be potentially achieved. The process involves application of electric and ultrasonic energy to liberate minerals from gangue particles. Introduction The mineral industry is a large user of energy in the US. Energy usage is increasing as the grade of ores processed decreases. It is es¬timated that comminution of ores uses about 32 000 kWh (115 trillion kJ), or 2% of the electric power produced in the US (NMAB-365, 1981). Among other minerals, ce¬ment, iron, and copper processing plants are the largest users of en¬ergy in comminution. Very little of this energy used in conventional grinding (about 1%) is used to gen¬erate new surface. The remainder is wasted (Table 1). Thus, there are substantial energy-saving and economic incentives to improve the efficiency of crushing and grinding techniques for mineral recovery. With an energy efficiency of only 1%, it would seem possible to devise methods to significantly improve comminution technology. This requires that breakage force be applied only where needed, not indiscriminately as in a conven¬tional ball mill (NMAB-365, 1981). Can apparatus and methods be de¬ veloped for large-scale commer¬cial use that allow energy to be fo¬cused at intergranular bounda¬ries? If this can be done, mono¬mineral grains would remain in¬tact and the grinding action would be selective and substantially more efficient. The novel comminution process concept described in this paper uses a combination of electric and ultrasonic energy. This energy breaks the ore and selectively lib¬erates minerals. The process has also been termed a two-stage or electroacoustical comminution process (Goldberger, Epstein, and Parekh, 1982). The mineral grain boundary is usually the weakest area in an ore. By applying electri¬cal energy to the ore, the rock frac¬tures mostly at grain boundaries. At the same time, the electric shock creates secondary hairline fractures in the ore. Additional application of ultrasonic energy to this ore provides further break¬age. The concept was proven on a molybdenum porphyry ore but needs additional study. Technical Background There have been substantial im¬provements in milling machinery and grinding operations, but rela¬tively few attempts have been made to develop alternatives to conventional impact milling. Sev¬eral methods, however, have been investigated that relate to this study. In the early 1970s there was considerable interest in a size re¬duction method known as the Snyder Process (Cavanaugh, 1972). It involved charging coarse ore into a pressure chamber, pressur¬izing with a gas, and activating a quick-opening (15 msec ) dis¬charge valve that connected the chamber to a discharge duct. This allowed solids to fluidize and ac¬celerate, subjecting the material to a variety of impulse phenomena that caused the desired size reduction. Other nonimpact means of achieving breakage and selective size reduction have been de¬scribed in technical and patent literature. Kanellopoulos and Ball (1975) considered using heat to induce thermal stress in quartz¬ite. This would cause cracking and, thus, increase grinding efficiency. The use of electrical energy to induce thermal stress was studied by the General Elec¬tric Co., in cooperation with the Montana School of Mines, under a research grant from the Anaconda

Jan 9, 1984

By Gregory P. White

A number of private companies and universities have developed computer programs which can be used to simulate surface coal mining by dragline. In this paper, we briefly discuss the concept of computer simulation, compare these computer programs with each other, describe the major changes required to convert these programs to an IBM computer system, and analyze their sensitivity to inaccurate input data. We also present a summary of the required data input for each program, core storage requirements, and output results. Sources for obtaining these programs and their documentation are listed at the end of the paper.

Jan 1, 1983

By S. S. Klima, D. E. Pollock, J. A. Organiscak

"Airborne respirable coal dust capture by water sprays or wet scrubbers has been studied and developed over many decades as an engineering control to reduce dust exposure in coal mines and combat coal worker pneumoconiosis. Empirical relationships and deterministic models for particular dust capture experiments have previously been devised to show the key parameters involved in airborne coal dust capture. Many of the results from these models show that the significant parameters related to airborne dust capture are water spray pressure, water quantity, water droplet size, relative water droplet-to-dust particle velocity, and total operating air pressure of the scrubber. However, many airborne dust capture efficiency relationships and models developed for particular experiments cannot be readily applied to forecast the dust collection efficiency of different spray and scrubber design configurations, which rely on several key dimensional engineering measures. This study examines engineering measures from previous water spray and wet scrubber experiments conducted by the U.S. National Institute for Occupational Safety and Health (NIOSH) and the U.S. Bureau of Mines (USBM) to develop empirical models for wet collection of airborne dusts. A dimensionless empirical model developed for predicting airborne dust capture efficiency of water sprays and wet scrubbers is presented. IntroductionWater sprays have been one of the earliest mainstay engineering controls used to suppress respirable dust in underground coal mines. Some of the earliest reports of water spray usage on continuous mining machines indicated respirable coal mine dust reductions of 20 to 60 percent (Kobrick, 1970). Not long after the enactment of the 1969 Coal Mine Health and Safety Act, the U.S. Bureau of Mines (USBM) and others conducted extensive research into water spray and wet scrubber systems to control respirable dust in coal mines. The primary aspects of water spray dust control included prewetting for dust prevention, localized ventilation redirection and airborne dust capture, which are summarized in handbooks or best-practices guides (Kost, Yingling and Mondics, 1981; Kissell, 2003; Colinet et al., 2010). Initially, underground studies were conducted to identify the primary sources of dust during the mining processes and effective dust controls (Courtney, Jayaraman and Behum, 1978; Jankowski and Organiscak, 1983). Laboratory research systematically studied ventilation and airborne dust capture effects of water sprays and scrubbers to improve their application in underground coal mines (Tomb, Emmerling and Kellner, 1972; Divers and Janosik, 1978, 1980; Grigal et al., 1982; Ruggieri et al., 1983; Jayaraman, Jankowski and Kissell, 1985; Volkwein, Ruggieri et al., 1985; Jayaraman, Schroeder and Kissell, 1986; Divers, Jankowski and Kelly, 1987; Jones and James, 1987)."

Jan 10, 2018

By C. P. Blair

In discussing so-called "volume rates" for the movement of coal, I start from one basic premise: That it is the job of the railroads to devise rates which will move the traffic, while at the same time assuring a fair return on their investment. No matter how reasonable and nondiscriminatory freight rates may be, if they fail to attract and hold traffic to the rails, they are utterly useless - both to the carriers and the shipping public. And it will be my purpose here to show that under certain circumstances the use of volume rates by the railroads enables them. to hold traffic which would otherwise be lost, and to avoid needless dissipation of revenues, Properly employed, volume rates should benefit both the coal industry and the railroads, I do not, however, advocate the discarding of the carload as the basic rate-making unit. Far from it. Sound operating reasons dictate the making of rates based on car- loads, and the tradition of carload rates has a firm foundation in railroad operating costs, many of which vary directly with cars handled. In short, the carload was, Is and, in my opinion, will continue to be the basis for most rail rates on coal, as well as on other commodities. But change is an inevitable result of progress, and we should not be so bound by tradition that we are unwilling to try novel methods of ate making, should the circumstances require, The modern rate maker must, now and then, run the risk of being called an iconoclast if he is to keep up with the times. As T will point out, volume rates were for many years rejected by both the Interstate Commerce Commission and the railroads, But now both the Commission and the carriers have come to recognize that under proper circumstances and conditions volume rates are a powerful and effective tool.

Jan 1, 1959

By K. Sasaki, W. Astuti, N. Okibe, T. Hirajima

"The kinetics of leaching of a saprolitic ore from Indonesia by citric acid solution under atmospheric pressure was investigated. An examination of the effects of leaching temperature, citric acid concentration, pulp density and ore particle size on the dissolution rate of nickel found that they all had significant influence on the rate. The highest nickel recovery (95.6 percent) was achieved under the leaching conditions of ore particle size of 212-355 microns, citric acid concentration of 1 M, leaching time of 15 days, pulp density of 20 weight/volume percent, leaching temperature of 40°C and shaker speed of 200 rpm. The shrinking core model was found to be appropriate for describing the leaching kinetics of this ore in citric acid solutions at atmospheric pressure. The experimental data were well interpreted by this model with rate of reaction controlled by diffusion through the solid product layer. Using the Arrhenius expression, the apparent activation energy for nickel dissolution was evaluated as 12.38 kJ/mol. Finally, on the basis of the shrinking core model, a proposed empirical kinetic model for the leaching of nickel from this Indonesian saprolitic ore was expressed as a mathematical model, which was verified as consistent with the observed experimental results.IntroductionIndonesia has some of the world’s largest nickel laterite reserves and was the largest producer of nickel – in the form of nickel lateritic ores – in the world in 2013 (Merchant Research and Consulting Ltd, 2014). However, only high-grade saprolites are processed in Indonesia using pyrometallurgical methods to produce ferronickel and nickel matte. Lower-grade saprolites and limonites are removed, becoming waste or overburden in mining site areas. Hydrometallurgical methods, involving either chemical leaching or bioleaching, are now seen as the most important technologies for treating nickel laterites in the future because of their advantage of not only extracting nickel and iron but also other valuable elements. In addition, hydrometallurgical methods consume less energy and have lower capital and operational costs. Although they are recognized to be more suited to limonites, some researchers have studied their use in the extraction of nickel and other metals from saprolites (Guo et al., 2009; Lu et al., 2013). Hydrometallurgical processes for the extraction of metals, especially nickel, promise to be meaningful alternatives to pyrometallurgical processes to primarily treat low-grade saprolites for which pyrometallurgical processing is not favorable"

Jan 1, 2015

By J. R. Walls

The United States potash industry had its beginning in the early Colonial days. Potash was initially manufactured by leaching of hardwood ashes in large iron pots. Production of potash was recorded as early as 1631 in New Hampshire, and in Maine a few years later. According to the census of 1850, the production of potash was America's most thriving chemical industry and, at that time, 569 plants were reported producing potash from wood ashes. Interest in potash was so great prior to the 1800's that the very first United States patent, signed by George Washing- ton, was granted for the production of potash on July 31, 1790. A new era for the production of potash began when potash was discovered by the analysis from a brine well in 1839 near Stassfurt, Germany while exploring for salt. The predominant use of potash at that time was in the chemical industry rather than agriculture. The discovery, by Justice von Liebig in about 1858, that potash salt contributed greatly to the essential plant food in soil opened up an enormous potential market for a suitable potash material.

Jan 1, 1985

By M. A. Berdine, C. N. Thompson, D. Brock

The raise bore method of shaft construction has been success-fully used in the mining and construction industries for over thirty years. In the mid 1990’s, Twentymile Coal Company (“TCC”), now affiliated with RAG American Coal Company, used this method for a moderately deep small diameter shaft and the raise bore and slash method for a shallow larger diameter shaft. In late 2001, TCC contracted to raise bore a 431-meter deep, 6.2-meter diameter shaft. During the raise bore and shaft lining processes, regional horizontal stress fields resulted in extensive elliptical caving beyond the original design diameter. This paper discusses the engineering approaches and procedures used to safely complete construction of the raise bore shaft.

Jan 1, 2004

By Ted H. Eyde

Improving the mineral industry's tarnished image describes in a nut- shell the reason why the Speakers Bureau of the Tucson Subsection of the AIME was formed. Let's go back in time about 3 years to the fall of 1971. Literally thousands of bills detrimental to the mineral industry were being introduced into Congress and State Legislatures. Unfortunately, many were being passed and signed into law. Clean air and water, zero pollution, preserve the environment, ecology, biodegradeable, wilderness were the words that the self- styled ecologists and environmentalists were preaching. The mineral industry either sat mute or in place of acting, reacted in a way certain to attract a maximum amount of bad publicity with a minimum amount of rhetoric. A mining company president, for example, commenting on the primary air standards in Arizona was reported by the press as having said, "we'll shut all the smelters down and put thousands of Arizonans out of work if we have to meet the Arizona standards." Or another defending a poorly planned highly visible open pit nonmetallics operation, "we'll take the whole mountain down if we want to, and nobody can stop us" and still another manager of Operations, "the bottom line is all that counts, to hell with all of this environmental nonsense."

Jan 1, 1975

By Paul B. Valenti

The Nevada Moly Operation of Anaconda Minerals Co., (a molybdenum/copper ore body) was initiated in 1978, with a contract to Bechtel Civil and Minerals Inc. for construction of a mining and milling facility near Tonopah, NV. Ground was broken on the concentrator portion of the project in August 1979$ with a multi-phased startup of the mill beginning in June 1981. Processing design tonnage of 20 000 t/d (22,000 STPD) was reached in early 1982. Major pitfalls of rapid design and implementation of engineering, construction, and startup are considered. The operation of the concentrator including SAG milling, ball milling, bulk rougher flotation, copper/moly separation, moly regrinding and cleaning, thickening, and product loadout are discussed.

Jan 1, 1983

By Güldane Atesoglu, İsmail Atilgan

The changes in the elemental contents and the structural properties of chalcopyrite concentrate when roasted were studied, and the main reactions contributing to the leaching process revisited. Firstly, TG/DTA measurements were performed to identify the roasting temperatures that influenced the type of species reacting with sulphuric acid leaching liquor. The temperatures of main reactions, whose output may affect the subsequent leaching process, were analysed for the concentrate at hand, and concluded to be generally lower than those reported in literature. Secondly, the obtained results were confirmed by images and XRD measurements on the roasted concentrate at temperatures in between 350 and 600 °C with increments of 50 °C. Finally, the leaching behaviour of the roasted specimens in a stirred sulphuric acid solution examined and the Cu/ Fe selectivity investigated that the maximum Cu/Fe ratio in the pregnant solution would nearly be achieved for the roasting temperatures lower than 500 °C, since the oxidation temperature of iron sulphate is found lower than that of copper sulphate.

Aug 11, 2022

By Lynn Huntsman

INTRODUCTION The Sunnyside mines of Kaiser Steel Corp. are lo¬cated on the west flank of the Book Cliff Range, a tongue of the Wasatch mountain range in Utah. The mines are 193.12 air km (120 air miles) southeast of Salt Lake City and 40.23 km (25 miles) east of the town of Price. The area has a semiarid climate with an an¬nual rainfall of 304.8 to 381 mm (12 to 15 in.) and a temperature range of -23.3 to 32.2°C (-10 to 90°F). The base of the Book Cliff Range is approximately 1981 m (6500 ft) in elevation with some mountain peaks in excess of 3048 m (10,000 ft) (Fig. 1). The Sunnyside mines were opened in the late 1890s and have been operated continuously since that time. Kaiser Steel leased the No. 2 mine in 1942 and pur¬chased the entire property in 1950. There are three mines adjacent to one another, each with its separate access, ventilation, and haulage. The three mines en¬compass an area extending 12.87 km (8 miles) along the outcrop and dipping northeasterly 2.41 to 3.22 km (11/z to 2 miles) into the Book Cliffs. During 1975, only the No. I and No. 3 mines were operating to sup¬ply 908 184.7 t/a (1,000,000+ stpy) of clean coal to the Kaiser Steel mill at Fontana, CA. Sunnyside coal is blended with coal from the Kaiser Steel mine near Raton, NM, and with purchased coal from noncompany sources to form the mix for the coke ovens. At that time, Sunnyside operated three longwall units and five to seven continuous miner units. Typical production for that time came from one miner unit pillaring an area not conducive to longwall mining; one or two miner units on construction or modernization projects, and the remaining units used to develop future longwall panels. Longwall mining has been a part of the Sunnyside scene since 1961 when the first panel was set up in the lower Sunnyside seam of the No. 3 mine. Through 1975, 27 longwall panels were mined (Fig. 2). Of the more than 47 173 604 t (52,000,000 st) of clean coal that were taken from the property through 1975, 6 350 292.9 t (7,000,000 st) were mined by longwall methods. Coal was brought to the surface in 4.5-t (5-st) mine cars and crushed to -152.4 mm (-6 in.) before entering the preparation plant. Coarse coal is cleaned by Baum jigs and fine coal (-28 mesh) is cleaned in flotation cells. The coal is reduced to 50.8 x 0 mm (2 x 0 in.) and belted to an aboveground storage pile with a ca¬pacity of approximately 113 398.1 t (125,000 st). The coal is gravity fed through two 1.52-m (5-ft) diam culverts directly into railroad cars, which pass beneath the storage pile through an arched excavation. Eighty-four 90.7-t (100-st) cars make up a unit train which makes the round trip to the steel mill in four days. Additional trains are scheduled in as demands dictate. Loading rates in excess of 13 607.8 t/h (15,000 stph) have been attained. Geology The coal seams of this area are of Cretaceous age and are classed as a medium volatile bituminous coal. coal, with a sulfur content less than 1 % and 6 to (20-mile) zone of the Book Cliffs, and this area is owned or under lease by either Kaiser Steel or US Steel. Coal in adjacent areas is noncoking but an excellent steam coal, with a sulfur content of less than 1 % and 6 to 7% ash. Sunnyside coal contains 40% volatile, 53% carbon, less than 1 % sulfur, and 51/2 to 61/2 % ash after washing. There are two minable seams throughout most of the Sunnyside property known locally as the upper and lower Sunnyside. Separation between the seams varies from less than 300 mm (1 ft) near the southern boundary adjacent to the US Steel Columbia mine, to over 12.19 m (40 ft) in areas of the No. 3 mine in the center of the Sunnyside property. The upper seam varies in height from 1.2 to 1.98 m (4 to 61/2 ft) while the lower seam varies in height from 1.52 to 4.28 m (5 to 14 ft) in areas of No. I mine where the two seams apparently come together. The majority of early work was con¬centrated in the upper seam which has a stronger roof and a lower sulfur content. When Kaiser Steel began mining the No. 3 mine in the early 1950s, it mined in the lower seam beneath the old workings with conven¬tional and continuous mining equipment. Numerous bounces and pressure points created by unextracted pillars and barriers in the upper seam made this a low¬production, high-cost mine and brought about the ad¬vent of the first longwall system. The first unit, while small by today's standards, successfully mined panels that had been all but given up and saved the mine from closure. There were areas that could not even be mined by longwall because of the erratic extraction in the upper seam. The system was successful enough to allow operations to gain access to and develop virgin areas of the upper seam. Development of longwall mining in the No. 3 mine has been confined to the upper seam since late 1964. With the almost complete extraction afforded by the longwall system, it is felt the lower seam can be successfully worked in future years (Fig. 3). In most of No. 1 mine, only the lower seam (or lower seam and upper seam combined) is minable, except in the northwest portion where a rock split has again separated the seam into two 1.83 to 2.13-m (6 to 7-ft) thick beds. In this area, the upper seam or upper split is being mined. The separation between seams at the present workings in this area varies from 0.609 to 6.1 m (2 to 20 ft). It is hoped that most of the lower split can be mined using shield supports and/or what¬ever new technology is developed in the coming years. The coal seams in this area dip to the northeast at an average grade of 10%. The dip near the outcrop reaches 22% and gradually decreases downdip. Over¬burden ranges from a few meters at the outcrop to 762 m (2500 ft) over some of the deeper workings. In the next several years, some overburden will be in excess of 914 m (3000 ft) as the lower mine workings extend ever deeper under the Book Cliffs. To cope with the

Jan 1, 1982

By A. E. Sears

The intent of geomorphic landform design is to mimic the function of the natural landscape. In a surface mine reclamation application, geomorphic landform design has the potential to mitigate stream loss. This research quantified features of reference landforms and performed computer modeling of a conceptual valley-fill design using geomorphic landform principles. A design was created using published design parameters with a software design tool (Carlson Natural Regrade with GeoFluv). Three additional designs, incorporating field-obtained regional geomorphic landform parameters, were completed and resulted in a geomorphic landform design specific to southern West Virginia. This work illustrates the importance of field determination of the ridge to head-of-channel input and drainage density parameters. Values of ridge to head-of-channel distance measured in this study were consistently an order of magnitude greater than software recommended values (SRV); SRV = 24 m (80 ft) versus field= 220 m (720 ft). Similarly, for the drainage density parameter, the software recommended range is 6 to 9 km-1 (80-120 ft/ac). The field measurements for central Appalachia quantified the drainage density to range from 5 to 5.3 km-1 (67-70 ft/ac). This affects Appalachian valley-fill stream construction by making the stream lengths shorter and the land slopes steeper with straighter headwater channels compared with areas in the southwestern United States. While the application of geomorphic landform design to surface mine sites presents challenges, this work provides support for future applications.

By James J. Scott

I am indeed honored to be the 1991 recipient of the Daniel C. Jackling Award. To be voted this honor by my peers in the Society for Mining, Metallurgy, and Exploration and AIME means more to me than I can truly express. Daniel Jackling's position and prominence in the mining industry first came to my attention in the fall of 1949 when I transferred from Wisconsin Institute of Technology, then a three-year school, to the Missouri School of Mines to complete my last year for a bachelor's degree in mining engineering. I played basketball in the Jackling gym and ran track on the Jackling field. I didn't know a lot about Jackling then, except he was an honored graduate of the school and, obviously, supported the school financially. Over the years, Jackling's name kept coming up. He was a man with a reputation for being a doer, a mining leader who was always a step ahead, a mining engineer who knew how to motivate men in the working place, to organize mining systems, to finance and organize corporations. to open up new, low-grade ore bodies for exploitation. If Jackling were alive today, I would expect he would bean excellent candidate to apply practical rock mechanics to his projects. Looking over the list of Jackling Award recipients, I have known many, have met many personally and had a few who were dear friends that affected my own professional life. Ab Shoemaker, '58, and I consulted at the White Pine Copper Co. on mining problems in the 1960s. We arranged for our visits to the mine to coincide so that we could collaborate on our ideas. I would rent an auto and drive for Ab. He said he was the only mining engineer in the country that had a PhD for a chauffeur. James Boyd, '67, was my mentor as regards the White Pine Rock Mechanics Program that we organized in 1963. Only through Boyd's forsight. leadership, dedication and willingness to spend money on research and development did the Practical Rock Mechnics Program at White Pine succeed. He never wavered in his support to that program or to me personally. Tom Howard, '75, was my predecessor in 1970 as Director of Mining Research for the US Bureau of Mines. When I joined the Bureau in August 1970, he had the research centers up and running. Any successes I may have had with the Bureau of Mines were due in great part to Tom's leadership and organization throughout the 1960s. Gene Cameron, '68, was my teacher at the University of Wisconsin and I remember him fondly. He was the one professor who lectured in a manner that made me welcome going to his classes each day, a truly remarkable educator. If you have come today to listen to me talk about instrumentation, rock properties, stress-strain relationships, rock mechanics theory, etc., you will be disappointed. Large volumes of data have been collected by well qualified theoreticians to define stress-strain conditions in mines. Unfortunately, very little of this data has been used by industry due to the lack of initiative and understanding by the researchers to the true problems of implementation in the industry. I hope my remarks will assist future investigators to work through the maze of operational problems so they will experience the satisfaction of seeing their research used. I fully realize that the Jackling Award is given by the Society for Mining, Metallurgy, and Exploration and is primarily offered for accomplishments in hardrock mining. I have found it impossible to separate hardrock and coal mining when I talk about "Applied Rock Mechanics." There is too much overlap and the same mechanics principles are applied to both. For example, coal mine designs have been applied to salt, trona and potash. Also, the original layout and design of White Pine Copper Mine was based on coal mine designs. Any developments to apply rock mechanics ideas must

Jan 1, 1992

By M. Lotfi

The criteria for evaluating equipment operator performance were identified as the changes in reaction time, the changes in physiological responses, and the changes in subjective reporting of stress. Data were collected while the Subjects operated a front-end loader simulator. This report out-lines the components and the tasks generated by the simulator. It explains the simple and basic assumptions underlying the design of the simulator and the methods of data collection. Improvements that can facilitate the collection of data, including the connection of the simulator to a portable personal computer for on site use are discussed. Statistical analysis of data showed significant deterioration in performance. The report concludes that operator performance can be measured and improved. This report shows that the operator tasks can be simulated by means of simple simulators, the complexity of which is determined by the availability of funds and the area of interest of the investigator.

Jan 1, 1989

By P. K. Hopke, A. Hubbard, K. H. Leong, J. J. Stukel, K. Nourmohammadi

INTRODUCTION In order to understand the transport and deposition of radon daughters in mine atmospheres, it is necessary to know the variation in the attachment of the daughter atoms to particles as a function of particle size, composition, number density, relative humidity, temperature, and radon concentration, the free gaseous diffusion coefficients of the daughters, and the variation in the mass transfer of the activity, both free and attached to particles, to mine surfaces as a function of particle size distribution, surface roughness of the mine walls, and the flow conditions. If all of these parameters are known in a model system, it should be possible to understand the transport and fate of the airborne radioactivity in real mines under certain well-defined flow conditions. There have been a number of recent investigations of the attachment of radon decay products to particles 1-4, but there are still a number of unanswered questions regarding the process. However, it is clear that for most real mine atmospheres, the vast majority of the activity is attached to particles. The size distributions for the activity-bearing airborne particles have been studied 5,6, and it has been found that most of the activity resides on particles with diameters in the range of 0.05 µm to 0.3 µm with an average mass median diameter between 0.1 and 0.2 µm. The behavior of the unattached radon daughter species has also been recently studied[ 7], and many of the previous problems regarding the value of the diffusion coefficient for Po-218 have been resolved. A major problem in the understanding of the airborne transport of radioactivity in mines is the lack of detailed knowledge of mass tranfer to and fluid flow over rough walls under fully developed turbulent flow conditions. This paper will report the progress on a project that is designed to obtained that information. MATHEMATICAL MODEL DEVELOPMENT Deposition of particles on smooth surfaces in turbulent flow has been extensively studied. A comprehensive review of these results has been prepared by Sehmel 8. There has not been such a comprehensive study of particle deposition on rough walls under such flow conditions. In recent years, only a single model has been proposed to explain such deposition 9,10 and in both of these papers the flow structure in the rough walled pipe was not taken fully into account. As part of the work being conducted on this project, a more complete model was outlined in a previous report [11]. The basic theory will be reviewed to provide a context for the flow measurements to be reported. The flux of particle deposited on the walls of a pipe in a turbulent flow is derived from the one dimensional form of Fick's law as given by [N = Dpdpp/dr (1) where N is the flux of particles deposited per unit area per unit time, D is the total eddy diffusivity of the particles, p is the airborne concentration of particles, and pr is the distance measured from the center of the pipe. The rate of deposition is best expressed by a deposition velocity Vp = NIP pb (2) where P b is the mean particle concentration in the sulk flow. The shear radius, V/ut and the shear velocity, u , are used to calculate a nondimensional distance, and velocity, respectively, where v is the kinematic viscosity of the fluid. The nondimensional form of equation 1 is given by Vd = DP dpp(3) V dr+ where Pp = Pp/ Ppb(4) By integrating equation 3 from the rough wall stopping distance, S , to the center of the pipe, the deposition velocity can be obtained. In order to make this calculation, it is necessary to have accurate descriptions for the particle eddy diffusivity, stopping distance, and shear velocity in order to insure that the influence of the flow structure has been properly accounted for. The shear velocity can be determined experimentally from the shear stress evaluated at the wall, Tw, and the fluid density, ut =VT w/p = ub V f/2 (5) where ub is the mean bulk axial velocity. The wall shear stress for a given pressure drop, dP/dL, and hydraulic diameter, Dh, is]

Jan 1, 1981

By James F. Fraser

Construction of a major linear project in any environment is difficult, at best. Construction of a major linear project which includes major subway elements, both cut and cover and tunneling, in the center of an urban environment, approaches the impossible. I believe such a project would be impossible for a private company or a privately owned utility for a variety of reasons. The major two reasons are the funding necessary for such massive projects and the myriad of individual interests associated with private ownership of land. Such projects are possible in the public sector only because the need for the project... as is the case with mass transportation in Atlanta. . .overrides the profit motive. Without the profit motive, private companies, whether utilities or not, are precluded from practical interest in the development of any project. While it is safe to conclude that a profit-oriented business would have no practical interest in developing such massive projects as may include major tunnels, it would be useful, I think, to compare some of the significant differences between private profit-oriented businesses, and an agency, or Authority, such as MARTA, in building a major linear project. SLIDE 1. Comparison of a Private Company and a Public Agency [Item Private Company Public Agency Simplified To make a profit To provide a pre Purpose scribed service]

Jan 1, 1979

By T. Lee

Ongoing economic success for mining companies is dramatically enhanced when they really know what they are harvesting, such that they can accurately forecast their head grades for any time period in excess of a month. This is an uncommon experience. It is a very common scenario for resource/reserve estimates to overstate the concentration (grade). Grade is a critical input of revenue, and hence, it is to cashflow projections. With the current direction being taken by the legal climate, it is no longer adequate for directors to accept the resource/reserve estimates of experts at face value. This is because resource/reserve estimates are the foundation of primary mineral economic decisions. Directors of mining assets need to undertake their own due diligence analysis of these estimates, in the best interest of themselves, and the company's shareholders. In the light of the industry's past achievements, this is quite a challenge. This paper suggests a set of new objective tools for directors to undertake an overview of the unique commodity concentration pattern and assess associated economic dynamics of each mineralization. These tools allow the examination of global graphical trends that can be helpful to directors and others in undertaking due diligence of resource/reserve statements. Ensure that it smells right.

Jan 1, 1995

By Arnold E. Lamm

The degree of economic, progress of a nation is measured by its possession and use of energy, sources. The future of coal in the West hinges on its ability to participate and grow in the energy market. A source of energy is, in the long run, only as good as the ability of the industry to develop its reserves economically. The economic development of the tremendous coal reserves, literally hundreds of billions of tons of coal in the States of Colorado, Utah, Wyoming, Montana, New Mexico, and Arizona, poses a very real question in the light of the known mining technology of our day. Basically in America today, coal is mined from the earth either by the strip mine or underground mine method; in the West, the large long-term, contracts now in force with utilities are mostly from coal mines utilizing the strip mine method. However, the preponderance of the known coal reserves of the West are located deep underground and will have to be mined by the various; underground mining processes. The problems encountered in the West with regard to underground mining vary from state to state but, on the whole, do not present what might be considered a miner's bonanza. I do not mean to' infer that the underground coal of the West is unmineable, but I do want to emphasize that the tremendous reserve figures do not present a true picture in the light of present day technology.

Jan 1, 1965