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By M. J. Zdunczyk

Sand is one of the most common materials on the earth’s surface. It is a naturally occurring material resulting from the physical and chemical breakdown of rocks. The definition of sand varies, as do the number of products made from the material. Sand is generally classified by grain size, however the measurement or scales differ depending on the context in which it is being used. Geologist define sand as particles of rock between .05 and 2 millimeters in diameter. The variability of the mineral composition generally dictates which sand can be used for different industrial purposes. For example, sand used for Portland cement concrete and asphaltic concrete mixtures can be very different in physical and chemical composition than sand used in the making of fiberglass. In this paper, construction uses of silica sand, which consists almost entirely of the mineral quartz (SiO2), will be discussed and analyzed.

Jan 1, 2001

By C. W. Schultz

The distribution of material passing a screen, or a set of screens, is described by the equation [ ] where p = the probability that a single particle will pass the screen on a single presentation k = number of layers of screen cloth n = distance from head of screen expressed as number of screen openings. This model can be used to predict the performance of conventional screening operations (i.e., k = 1) as well as the so-called "statistical screening" operations. Using this technique, in which multiple identical layers of screen cloth are employed, screening can be carried out at a finer size than is possible with conventional techniques with no loss of separation efficiency.

Jan 1, 1969

By D. Olsen, S. Cole, J. Strid, G. Colzani

This paper focuses on the underground tunnel work at the Hollywood Water Quality Improvement Project (HWQIP) in Los Angeles, CA. The HWQIP is being constructed to provide a water distribution system to bypass the existing Hollywood Reservoir System, which supplies drinking water for 500,000 people. The major project features include, a 2.9-meter diameter 1,829 meter long bypass tunnel, a 1.8-meterdiameter 366 meter long utility tunnel, 3 jack and bore tunnels, two 136 million liters underground circumferentially wrapped pre-stressed concrete water storage tanks, associated piping, and mechanical facilities. The physical project location coupled with strong community involvement and stringent environmental concerns was a considerable driving force for this project located in the Cahuenga Pass area of the Hollywood Hills north of Los Angeles, CA.

Jan 1, 2001

By Y. Yu

Florida phosphates ore typically processed by reverse flotation with an intermediate acid-scrubbing step. A single-reagent flotation process, using fatty acids or amines as the collector, would eliminate the intermediate scrubbing step plus an additional flotation step. This would not only simplify the process flowsheet but would also result in reduced reagent costs. The potential of single-reagent flotation of Florida phosphates using fatty acids as the collector is examined in this study. Introduction The beneficiation scheme now used for Florida phosphates involves the following three steps (Smith, 1976: Lawyer et al.. 1978, 1984: Moudgil et al, 1976, 1979, 1986, 1987): •rougher flotation of fluorapatitc particles using a fatty acid-fuel oil mixture as the collector. •scrubbing of fatty acid, and •cleaner flotation of silica particles using amine collector. Several attempts have been made to simplify the process flowsheet by floating either apatite or silica (not both, as is the current practice). However, single-reagent flotation schemes have had limitations in meting the concentrate-grade requirements of the phosphoric acid plants. In this study, an attempt was made to use a minimal amount of the fatty acid-fuel oil collector to obtain an acceptable grade (rougher) concentrate. Invariably, the recovery in such cases is expected to be lower than in the conventional rougher-flotation step. Therefore, scavenger flotation of tailings with additional fatty-acid collector was conducted to improve the phosphate recoveries.

Jan 1, 1997

Global mining company achieves significant productivity gains with three-dimensional (3D) mine models developed with Abaqus finite element analysis software. In a steep change beyond traditional processes, Abaqus finite element analysis (FEA) software from Simulia (the Dassault Systèmes? brand for realistic simulation), is being used to enhance mine design and engineering simulation at a number of major mines around the world. In North and South America, Africa and Australia some of the world?s biggest mining companies are applying FEA technology to evaluate safety and improve design planning, implementation and operations. Beck Arndt Engineering (BAE), a Sydney-based international consultancy, is a pioneer in the commercial development of engineering solutions for the mining industry. The consultancy has worked closely with engineers at Simulia Australia to expand the use of Abaqus FEA simulation software for mining applications. Among the early adoptors of mine-ready FEA technology is the world?s largest miner, BHP Billiton. With BAE?s help, BHP has already applied this technology to evaluate mines in Canada and Australia. At the BHP Billiton Nickel West Perserverance Deeps Project in Western Australia, Abaqus FEA software is now being used to help engineer the safety and productivity of planned deep-mining operations. To achieve this goal in the deep-mining environment significant technological innovation is required. Using measurements of site deformation and seismicity, Abaqus FEA models have been calibrated and, in a single day, used to simulate a full, 3D, inelastic analysis of a mine?s life cycle. In recent years, similar applications at Debswana?s Jwaneng Mine in Botswana, the Newcrest Mining Ridge-way Deeps Project in New South Wales, Australia and Rio Tinto?s Argyle Diamond mine in Western Australia have also established Abaqus FEA analysis as the lead-ing technology for multi scale, simulation-aided mining engineering. Joop Nagtegaal, a pioneer of FEA and a Dassault Systèmes Corporate Fellow, said that Abaqus FEA software is unique in its capabilities to enable mining engineers to investigate design innovations from the drawing board to full production. ?In the design stage, Abaqus models, which include rockmass volumes spanning several kilo-meters around the ore body and down to excavations just a few meters across, are used to compare and optimize engineering options,? Nagtegaal said. ?Then, as the mine goes into production, large volumes of data from the field are incorporated with the analysis models to allow them to be calibrated to a precision not previously available to the mining industry.?

Jan 1, 2009

By William Lyman

GENERAL DESCRIPTION Shrinkage or shrinkage stoping refers to any mining method in which broken ore is temporarily retained in the stope to provide a working platform and/or to offer temporary support to the stope walls during active mining. Since ore "swells" when broken, it is necessary to shrink the muck pile a corresponding amount by draw¬ing some of the broken ore out as the stope is advanced-hence the name. Broken ore retained during stoping is drawn out after the stope has reached its limits. The stope may be left empty or may be filled with waste contemporaneous with, or subsequent to, the final draw. Traditionally the method implies conventional overhand stoping methods with miners working between the muck pile and the stope back, in a space which advances updip with mining and is maintained by balanc¬ing "swell" with "shrink." The shrinkage classification is also applicable to so-called "semishrinkage" methods in open pillar-supported stopes where broken ore is temporarily retained as a working platform but offers no wall support; and to various blasthole shrinkage methods which utilize broken ore temporarily retained in the stope for wall support, but which do not require miners to work from muck pile in the stope. The method is generally applied to steeply dipping veins of strong ore between strong walls. APPLICATION Geometry The geometry of a shrinkable vein is described in terms of dip, width, and regularity along dip. Overall strike and dip dimensions and irregularities along the strike generally impose no restrictions on the method. Dip is ideally 1.2 to 1.5 rad (70 to 90°). As dip falls below 1.2 rad (70°), the shrinkage draw begins to strongly favor the hanging wall side, thus leaving a poor working platform for conventional overhand work. This is particularly true in relatively wide stopes. The sup¬port afforded to the hanging wall also diminishes with decreasing dip, reaching nil as the dip approaches the repose angle of broken ore. Dips below 0.78 to 0.87 rad (45 to 50°) are not generally shrinkable except by open stope "seinishrinkage" methods. Minimum mining width is fixed by working space requirements in the stope-generally about 1 m. Shrink¬age in narrower veins requires that waste rock from one or both walls be broken with the ore and the attendant dilution accepted to achieve the minimum width. Nar¬row stopes are less suitable, encouraging hang-ups and bridging of broken ore, with the attendant problems of erratic draw and incomplete recovery of broken ore. Maximum practical width may be 3 m or less to over 30 m, depending upon the competency of the ore and its ability to stand unsupported across the stope back. This is a vital safety consideration in conventional over¬hand stopes, but is much less of a factor in blasthole shrinkage methods. Very wide veins and massive ore bodies have been mined by transverse vertical shrinkage panels separated by transverse vertical pillars which are either abandoned or recovered later by other methods. Regularity along the dip is a prerequisite of shrink¬age as there must be no serious obstruction to the flow of broken ore downward through the stope to the sill level. Gentle rolls along the dip are acceptable if the local footwall dip everywhere exceeds 0.78 to 0.87 rad (45 to 50°). Off-dip hanging wall and/or footwall splits can generally be mined selectively from a conventional shrink stope as they are encountered without ad¬versely affecting subsequent continuation of shrinkage mining updip on the main vein. Vertical offsets or major rolls along the dip which cannot be "smoothed over" generally require that a sublevel be established with new draw control development. Blasthole shrinkage methods are much less flexible (and thus less selective) in their ability to accommodate any of these irregularities. Ground Conditions The wall rock must be strong enough to stand with the minimal support afforded by the dynamic mass of broken ore in the stope. During active mining, local sloughing from the walls is restrained, but the broken ore affords little, if any, useful resistance to closure of the stope walls. Such squeezing, if present, may bind up the stope and cause the loss of much ore. Pillars left between and/or within stopes are effective in preventing closure but reduce overall recovery. Walls may be re¬inforced by bolting after each stope cut in conventional shrinkage but not in blasthole shrinkage. Ore in place must be strong enough to stand with no natural support across the stope width, although tem¬porary artificial support or reinforcement may be used locally in conventional stopes. Some spalling or sloughing is permissible in blasthole shrinkage as men are never present in the stope. Physical and/or mineralogical characteristics of the broken ore may impose restrictions on stope design and/or operational plan¬ning, and may even preclude the use of shrinkage al¬together. Examples include: ores which, when broken, are cohesive or which tend to pack or cement together under the influence of ground water, wall pressure, and/ or chemical reaction. Such conditions precipitate er¬ratic draw during mining and often result in difficult and/or incomplete final draw; pyritic ores which oxidize very rapidly in the stopes and may generate heat, imposing a fire hazard by spontaneous combustion; sulfide ores which oxidize sufficiently in the stopes to adversely affect mill recovery by flotation; and ores (es¬pecially those containing uranium minerals) which ex¬ude radon gas and thereby impose ventilation constraints on stope design. In most cases these problems can be minimized by limiting the size of stopes, by minimizing the duration of mining activity in each stope, and by promptly drawing each stope empty following comple¬tion of mining.

Jan 1, 1982

By C. Newman

"Over the past decade, as conventional underground deposits have been depleted, mining operations have been forced to produce at greater depths and in more geologically and geometrically challenging conditions. As such, there has been a global increase in the application of cemented paste backfill (CPB) in tabular deposits utilizing open stope mining method with a delayed backfill placement. Despite the extensive use of CPB, many fundamental factors affecting the design of safe and economical fill structures are still not well understood. A critical issue in the design of backfilled stopes is the determination of stress states within the fill material itself as well as the surrounding rock mass. Analytical equations provide a means of quickly evaluating the effectiveness of a given design. However, in developing these equations large assumptions are implemented to simplify the design problem. It is important that one understands these simplifications as well as their effect on the overall design. This paper investigates common analytical equations utilized in the evaluation of single vertical backfilled stopes and their assumptions through comparisons to numerical modeling results. INTRODUCTION AND BACKGROUND As the mining industry continues to produce at greater depths and in more geometrically and geologically complex conditions, cemented paste backfill (CPB) has gained traction as a means of providing localized ground support in modern stope-and-fill mining operations. The application of CPB in stope-and-fill mining as a ground support material has led to the reduction of mine waste costs, while increasing both mine production and stability. In an attempt to achieve total extraction of the mining reserve, stope-and-fill mining operations often employ a primary-secondary excavation-support sequence (see Figure 1). In this practice, primary stopes are initially excavated and then backfilled in a two-stage pour; plug and final pours. During the first stage (plug), a CPB material with increased cementitious content is placed at the bottom of the stope providing a strong base from which to build a backfill structure as well as protect mine works from material breakthrough. Following placement, the backfill plug is allowed to cure to a given strength as designated by the mine design. Upon achieving the required cure strength, a less cementitious CPB material is used to backfill the remaining stope area. This is referred to as the final pour stage. With all primary stopes excavated and supported using backfill, the extraction of the secondary stopes commences as primary stopes are utilized as artificial pillars. While the application of CPB in a primary-secondary extraction-support mining sequence allows for a substantial increase in reserve recovery, it is imperative the mining personnel and planning engineers have a clear understanding of the material’s behavior and stress interaction with the surrounding rock mass to ensure the most efficient, cost-effective, and safe extraction of underground deposits."

Jan 1, 2018

By Daniel de la Rosa, Juan Manuel Mayoral, Damián Vital

"This paper presents a numerical study regarding the loads, and the corresponding damage, taken by the primary and secondary lining of a tunnel built in cemented fine-grained stiff soils. This study aims at establishing more rational design criteria that take into account explicity the lining performance. A 3.5 km long tunnel to be built in a densely populated northwest area of Mexico City, was considered as case study. Series of three-dimensional finite difference models, developed with FLAC3D, were used to simulate the tunnel construction, and to compute the ground settlements, internal forces, bending moments, and displacements acting both on the primary and secondary tunnel linings. Interaction diagrams were used to study the lining performance. Several advancement lengths, and tunnel covers were considered to establish their impact on the loads transferred to the tunnel lining. A practice-oriented damage criterion for tunnels was proposed based on the interaction diagram analysis. INTRODUCTION Lining performance during conventional tunnelling, CT, as defined by the International Tunnel Association, ITA (ITA 2009) is strongly affected by the construction procedure. Special attention must be taken in cemented fine-grained materials, such as those found in the northeast portion of Mexico City, due to the potential of fragile failure (i.e. the soil fails suddenly without important deformation) that these soils may exhibit. During CT, it is assumed that the ground strength contributes to the stability of the excavation, as long as the lining is built correctly. It is common practice to consider two structures, which together form the tunnel support: 1) A primary lining, which is a flexible support, which allows excavation walls to deform to mobilize part of its shear strength, and thus contributes to the tunnel self-stability, and 2) A secondary lining. Commonly, the primary lining consists of several layers of shotcrete. In this case the installation method guarantees perfect contact between the lining and the excavation perimeter. The secondary lining, if required, is used to withstand the load conditions in the long term and also has an esthetic function. This paper presents a numerical study aimed at establishing a performance damage criterion for the primary and secondary lining design in cemented fine-grained stiff soils. Due to the lack of data, the strength and deformation parameters were established through empirically-derived relationships. Three-dimensional finite difference models were developed to simulate the tunnel construction, and compute ground settlements, internal forces, bending moments and displacements acting both on the primary and secondary tunnel linings. Interaction diagrams, along with a damage criterion were used to study the lining performance. From the results gathered insight was gained regarding tunnel lining performance during CT."

Jan 1, 2016

By Wen-Long Cai, A. Frederick Banfield

This paper investigates, through case studies on a porphyry copper and a coal deposit, the differences and relevances between an approximate production schedule approach, i.e., the best case and the worst case schedules and a practical production schedule approach. The comparisons were with regards to: 1) choosing an optimum ultimate pit size, 2) determining processing capacities, 3) selecting 'optimum' cutoff grades, and 4) blending. The comparison criteria was to maximize the project net present value. Conclusions were drawn based on the study results.

Jan 1, 1996

By William R. Yernberg

The Mineral Processing Plant Design, Practice and Control Conference was held from Oct. 20through 24, 2002, at the Fairmont Hotel, Vancouver, British Columbia, Canada. The conference was attended by nearly 400 of the world’s top mineral-processing specialists, including plant-design engineers, mineral-processing consultants, academics, suppliers and equipment manufacturers. The meeting was sponsored by SME and cosponsored by the Canadian Mineral Processing Division of the Canadian Institute of Mining, Metallurgy and Petroleum (CIM). The conference covered all aspects of mineral-processing design, construction and operation. The meeting included more than 130 technical presentations. Topics included sampling programs, feasibility studies, design criteria, process technologies, construction, start-up operations and shutdowns. To stimulate interaction and discussion throughout the five-day conference, the sessions were presented consecutively. Attendees had the opportunity to attend every session.

Jan 1, 2003

By Jan Wolski, Jack Barry

Ventilation ductlines with a single fan or with a series of fans are analyzed. The duct friction loss factors are discussed. Numerical solutions, presented in the paper, applies to any fan number and locations along a duct line. The findings are compared with formulas available in the technical literature. Values of the air leakage factors are presented in the SI, technical metric, and US unit systems. A method is shown of how to determine the leak- age factor for a given ductline. Optimum solutions for the sample duct lines with two and five fans in series are discussed and presented in a graphical form. Performance of fans in series is discussed in conjunction with both friction pressure loss and leak- age.

Jan 1, 1997

By R. B. Coleman, T. P. McNulty

HISTORICAL PERSPECTIVE Environmental protection is not a newcomer to the minerals industry, there having been numerous examples dating from decades ago. Sometimes the awareness of a need for preservation or restoration derived from the ethics and principles of the engineers and management responsible for development and operation of a property. On the other hand, pressures to clean up industrial messes frequently were applied by special interests such as farming and ranching, recreational and conservational groups, and local governing bodies. Mineral industry environmental controls in this century have therefore been a patchwork of good intentions and bad, clean operations and dirty ones. The Hoboken siphon converter for converting copper matte while effectively capturing sulfur dioxide at a relatively high concentration was developed in Olen, Belgium in 1928, a full sixty years ago. The alternative was closure of the Hoboken smelter by Belgian government health officials who had become intolerant of sulfur dioxide emissions which were causing serious respiratory ailments in the river valley in which Olen is situated. As early as 1958, the Anaconda Company had a full-time botanist on its metallurgical research staff with the responsibility of learning how to vegetate the extensive tailings deposits at Anaconda, Montana. By that time, considerable progress had already been made in restoration of Silver Bow Creek, which at one time was “dead” for as far as 40 miles down- stream from the point of entry of acid mine drainage in Butte.

Jan 1, 1989

By D. Kaja

The Symons Cone Crusher was invented by Edgar B. Symons in 1923. It was quickly accepted as a superior secondary and tertiary crusher. Nordberg manufactured crushers for the Symons Brothers Co. and in 1928 acquired the Symons Cone business. Since then, the Symons cone has become the standard of the Mining Industry. This is due to the basic rugged design and Nordberg's continued efforts to modernize and upgrade the crusher design over the years. In 1930, the Heavy Duty model was introduced with a 225 kW (300 HP) rating. Then in 1931 the Short Head version was developed to handle finer crushing applications. In 1950, the 7 Ft. XHD design was produced which provided higher crushing force capabilities for harder ores and for smaller product size requirements. Finally, in 1978, the 7 Ft Super Heavy Duty design was created to expand the power rating to 500 HP and to further increase the crushing force limits. The first 7 Ft. Super Heavy Duty crushers were delivered to Bougainville Copper Ltd. in 1978 for the production of primary ball mill feed in a three stage crushing plant with the third stage in closed circuit. The product size was 80% minus 6.35 mm (1/4 in).

Jan 1, 1998

By Tedd B. Bodimer

Why complicate an already complex machine such as a continuous miner by adding roof drills? Immediate permanent roof support is an immediate thought of the mining engineer just as elimination of frequent time-consuming places changes is a thought of the methods engineer. The realization of those two objectives has the potential for improved safety with increased production is the thought of the general manager. Miner-Bolters Used a Generation Ago These are not new thoughts as the 1CM continous miner equipped with two integral roof drills was introduced twenty-five years ago. Approximately two hundred of these machines were sold in the next five years for use in medium and high seam operations. The 1CM miner chassis that remained stationary in the center of an entry or room while the ripper bar was swung and sumped hydraulically provided an ideal platform for mounting two roof drills that could be operated during the continuous miner's production cycle. Two roof bolts could normally be installed in less time than required for the mining cycle. A typical roof drill unit is shown in Figure I.

Jan 1, 1981

By M. Valdebenito, J. Rajala

An OCS Expert control system was implemented by Svedala Cisa at the Minera Maricunga Refugio Crush-Leach Operation near Copiapo, Chile. The purpose of the Expert system was to maximize production through the primary and secondary crushing circuits and help balance crusher usage between each stage to maximize total plant production. Interesting aspects of the control strategies are discussed. Benefits of the control were verified through a randomized On/Off test program. Tonnage increases ranged from 5% to approximately 20% at good confidence intervals. Benefits analysis results are discussed. The crushing plant at Refugio is very remote and is located at approximately 15,500 ft. This poses many challenges. The effect of altitude and remote location on plant operation are discussed.

Jan 1, 2002

By R. C. Gilmore, S. A. Lolon, G. E. Bogin, S. A. Saki, J. F. Brune, J. W. Grubb

"Gob ventilation boreholes (GVBs) are widely used in underground coal mines for longwall gob degasification purposes. GVBs are often drilled within 10 to 30 m above the top of the coal bed into the fractured zone, completed with 20 cm (8 in.) diameter casing and 60 m of slotted pipe at the bottom. The purpose of this completion strategy is to create a pressure sink to capture the emissions before they can enter into the underground workplace. For GVBs to work effectively, they must be drilled close enough to the working areas to capture methane from the fractured zone while the setting depth must be kept above the caved zone to minimize the amount of ventilation air that is drawn into the GVBs. Well completion parameters are important for creating safe working conditions. In this paper, computational fluid dynamics (CFD) studies will be presented to analyze the effect of different practices for GVBs completion parameters for methane extraction, formation of explosive gas zones in the gob and methane concentrations at the longwall face and tailgate. Authors have identified the optimum completion parameters for GVBs, which can maximize the benefit and minimize fire and explosion risk. INTRODUCTION Coal mines may be considered gassy based on the presence of CH4. Gas from the mined coalbed or from rider seams above or below the mined coalbed can migrate into the gob and then potentially move towards the working areas. In longwall coal mines with high methane emissions, sufficient ventilation and additional methane control measures are usually required to reduce the methane concentration in working areas to below statutory levels. Several techniques or combinations of techniques are available to control the methane hazard. These may include mine face ventilation to dilute methane, coal bed degasification, nitrogen injection into the gob to render a methane atmosphere inert and the use of gob ventilation boreholes (GVBs). Longwall face ventilation and dilution alone can only handle methane emissions up to a certain amount. When methane emissions exceed the face ventilation capacity, additional control measures are required. Drilling horizontal or vertical methane drainage holes into the coal seams as well as gob ventilation boreholes from the surface into the gob area to extract methane are common practices. GVBs are usually drilled from the surface above the longwall panel ahead of mining. When the face moves outby the GVBs and the gob begins to form, methane production starts from the boreholes. It is important to control the performance of GVBs to maintain a safe working environment in the mine. The performance of GVBs depends on multiple factors, including borehole location, length and diameter of slotted casing, setting depth of casing above the coal seam being mined, wellhead vacuum pressure, and permeability of caved-fractured zones."

Jan 1, 2016

By William R. Bond, Jim V. Rouse

Introduction The pressure/vacuum (p/v) lysimeter or soil water sampler has become useful for monitoring vadose zone or unsaturated zone water conditions. The soil water sampler was first introduced around 1961. It was used as a monitoring and research device for studies in agriculture and soil contamination around landfills. Since then, it has been used for many purposes. The use of p/v lysimeters at mining facilities to monitor for solution loss is a relatively new idea. The primary purpose of the lysimeter here is to function as an early warning system. It detects leakage through liners of tailings ponds, heap leach pads, and leachate holding ponds. Their use has proven valuable for collecting baseline data and detecting loss of precious leachate fluid. Different lysimeter designs are capable of vadose moisture sampling from near the surface and down to 90 m (300 ft). They are also capable of sampling soil moisture or saturation plume contaminants from areas not accessible by conventional water well installations. The most common installations are usually simple and cost-effective. In many cases, they can be installed without the use of heavy drilling equipment. A properly placed and installed lysimeter can detect solution leaks in several days or weeks. A conventional perimeter monitor well may take months or years to detect a leak of the same magnitude. A discussion on the general design and installation of p/v lysimeters will give an understanding of its versatility. Three case studies of lysimeter installations at mining sites will be examined in detail. Design, Installation, and Sampling Design The basic design of a p/v lysimeter chamber consists of a porous material through which the soil is drawn into a collection chamber. Access tubes run from the unit to the sample station at the ground surface. Most units currently available range from 25 to 76 mm-diam (1 to 3 in.-diam). The sample collection chamber, or body, can vary from 0.6 to 1.8 m (2 to 6 ft) in length. The body can be constructed of PVC casing or Teflon, depending on the installation desired. The collection chamber functions as a holding reservoir for the collected soil moisture. The intake, or porous material, on the lysimeter is the most complex and vulnerable portion of the unit. The pore size selection varies based on the needs of the installation, type of fluid to be sampled, and type of soil material surrounding the unit. Teflon (PTFE) and porous ceramic are used in fabrication of the porous inlet portion. When using a Teflon material, which is hydrophobic, a larger pore size of 70 µm is needed to overcome surface tension and draw water into the collection chamber. It is typically used at sites involving organic contamination, because of its inert properties. Ceramic material is generally hydrophilic, having less surface tension. Therefore, it has a smaller porosity range to conduct the moisture through the porous medium while under vacuum. Pore sizes may be selected from 0.16 to 6 µm, and can be manufactured for the specific needs of the installation. The final major design component of the p/v lysimeter is the access or sampling tubes. The tubes are usually of 6.3-mm-outside-diam. (0.25-in.-outside-diam). They are made of rigid polyethylene or flexible Tygon with a wall thickness of 1.5 mm (0.06 in.). The con-

Jan 4, 1985

By Louis W. Cope

Mongolia is known in western countries as the location of rich dinosaur fossil fields. However, the country is even richer in mineral resources. The Ministry of Industry and Trade reports 6,000 mineral occurrences of 80 metals and minerals. Mongolia has already attracted much foreign investment in the mining sector. Mining is the country’s largest industry. It is responsible for 17 percent of the gross domestic product, 65 percent of industrial output and 57 percent of export earnings.

Jan 1, 2006

Gary R. Steele is vice president, corporate marketing with Energy Fuels Resources. As such, he knows the global uranium market. The following are some of his thoughts on uranium?s place in the global energy picture. ME: What are Energy Fuels Resources? thoughts on the global nuclear power industry? Steele: The global nuclear power industry is a major contributor to the electrical needs of the world. Though generation by nuclear reactors has primarily attracted attention through the reactor accidents at Three Mile Island and Chernobyl, nuclear power continues to quietly and safely provide power to a growing and power-hungry world population. As of March 2008, 30 countries were operating 439 nuclear power reactors. Nuclear power plants provided about 15 percent of the world?s electricity production in 2007. France leads the world in nuclear utilization, with 77 percent of its power coming from nuclear plants. And Lithuania, Slovakia and Belgium each derive more than 50 percent of their power needs from nuclear plants. For much of the developed world, nuclear generation has already been contributing significantly to the supply of electricity. Additionally, 34 new nuclear plants are under construction in 14 countries. Leading the rush to build new nuclear generating plants are China, Russia and India. Each of these countries is building six new nuclear power plants. These 18 plants, totaling almost 12,000 MW of generation capacity, will require about 7.25 kt (16 million lbs) of U3O8 for their initial fuel load and will consume about 2.45 kt/a (5.4 million lbs/year) for the rest of their operating lives. The rest of the world, outside of the United States, has used nuclear power as a viable and economic generation alternative since the commercialization of utility nuclear reactors. And current activity supports the strong future of the nuclear alternative internationally.

Jan 1, 2008

By J. J. Gusek, T. C. Richmond, R. W. Reisinger

A high-altitude (9,500-ft [2,900-m]) historic underground copper mine in Wyoming discharges neutral but copper-laden water into a pristine creek. Under a Comprehensive Environmental Response, Compensation, and Liability Act (CERCLA) deferral agreement, the Wyoming Abandoned Mine Land Program performed site characterization and treatment studies in preparation for reclamation of the site. An integral part of the project included addressing the needs of a variety of interested and affected parties, including property owners, local populations, government agencies, and the media. This paper discusses an overall approach and specific methods employed to enlist public support for the project as well as some of the successes and failures experienced.

Jan 1, 2004