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By Joel K. Witt, Jackson G. Jenkins, Wendy M. Gort, Brent A. Whitman

A major concern for operators of leaching/SX-EW plants is a phenomenon known as gunk or crud. The build up of gunk is typically the result of fine particulates which are transported via the leach operation to the solvent extraction plant. The particulates develop into a solid layer that concentrates at the aqueous/organic interface. This can cause phase separation problems which can result in entrainment and other deleterious operating conditions. Most operations physically remove the gunk from the SX settlers and process it via a centrifuge or coalescer to remove valuable organic. While both methods have served for many years, Magma is currently evaluating the use of a chemical surfactant which induces a definitive separation of particulates from organic.

Jan 1, 1994

By Thomas E. Shufflebarger

Glass sand is a virtually pure, fine- to medium-grained, quartz sand. It is produced in 25 states from a variety of chemically different source rocks. Deriving their raw material largely from either friable sandstones or loosely consolidated sands, 59 operations produced 14.9 million tons of melting sand in 1976. About one-third of this tonnage was obtained from the Oriskany and St. Peter Sandstones. Glass-sand prospects are discussed in this paper with regard to chemistry, environment, transportation, and costs. Demand for a quality product will range from 36 million to 46 million short tons in 1995.

Jan 1, 1978

By Ernest Baafi, David Drew

Microsoft Excel based ultimate pit limit (UPL) optimiser has been developed to determine the greatest economic value of an orebody. UPL manipulates a given geological block model to define an ultimate pit limit using either the moving cone, method or Korobov algorithm or Lerchs-Grossman technique. UPL is designed to be an adjunct to MS Excel; the cells in Excel are excellent analogue to the blocks in a geological model, providing X-Y dimensions in an easily viewable and editable state. The use of individual worksheets within an Excel workbook provides the necessary depth to a 3-D block model.

Jan 1, 2002

By Bruce S. Brown, Jeremy P. Haile

The construction of tailings facilities and heap leach pads for gold mining operations generally requires the installation of one or more impermeable liners beneath the structures to satisfy environmental regulations and prevent the leakage of gold bearing solutions. Liners are usually constructed using clay soils or synthetic membranes, such as high density polyethelene. Low permeability soil liners can be constructed from natural clays, well graded soils with only a small clay content, or sandy soils blended with a small quantity of imported clay such as bentonite or kaolinite. The permeability of such materials, prgperly pped with good quality control, generally lies in the range of 10- to 10- cm/s, and can be designed to meet minimum prescriptive standards set by regulatory agencies. Important benefits resulting from the use of clay liners include the slow rate of penetration of a saturated wetting front, the decrease in permeability of the liner as it consolidates under the applied load of tailings or ore, and attenuation of potential contaminants in the process solutions. Synthetic liners can be attributed a finite permeability as a result of minute perforations and seaming deficiencies, and possess none of the above properties. Synthetic liners are also, by their nature, relatively thin and even a small head build-up above the liner will result in a large hydraulic gradient across it.

Jan 1, 1988

By Robert Mermelstein, Robert W. Kilpper

In a long term toxicology study SPF, F-344 rats were exposed 6 hours/day, 5 days/week for up to 24-months to a special test toner at 0, 1 (low), 4 (medium), and 16 (high) mg/m3, or Ti02 at 5 mg/m3, or Si02 at 1 mg/m3 by the inhalation route. The test toner was composed of about 90% random, high molecular weight acrylic copolymer and 10% high purity, medium color, furnace type, carbon black. It was enriched in fine particles such that the respirable aerosol concentrations (ACGIH) were to 0, 0.35, 1.4 and 5.6 mg/m3. There were no effects of exposure on survival, causes of death, body weight, food consumption or clinical chemistry and hematology parameters. There was no evidence for systemic or upper respiratory system toxicity as all of the changes were restricted to the lungs and associated lymph nodes. The lung tumor frequencies in the control (air and Ti02) and toner exposed groups were comparable. There were no changes at the low exposure level, which is the most relevant in regard to potential human exposures. The lung changes at the high exposure level consisted of decreased alveolar clearance, increased toner retention, increased lung weight, chronic inflammation, decreased pulmonary function and a moderate degree of focal fibrosis in all animals. The changes in the middle exposure level were smaller in magnitude than at the high level, and a very slight degree of fibrosis was found in 20% of the animals. In a follow-up investigation, female F-344 rats were exposed to one of three toner concentrations; control, low (10 mg/m3) or high (40 mg/m3) for 13 weeks followed by five, 13 week observation intervals. At the low exposure level, slight clearance retardation and modest inflammatory changes were observed at the end of exposure. They returned to normal values within six months post-exposure. In contrast, at the high exposure level, increased lung weight, decreased alveolar clearance and a disproportionate increase in toner retention, as well as an inflammatory response, were noted, which persisted throughout the post-exposure interval. While no significant histopathological changes were noted at the end of exposure, a very slight degree of fibrosis was observed in all of the animals in the high exposure group, after 65 weeks of observation. All of the above alterations are ascribable to "lung overloading", a generic response to excessive amounts of any dust retained in the lungs for a prolonged time interval. Published accounts in the toxicology literature document that long-term inhalation of large quantities of even innocuous particles such as Ti02, carbon black, petroleum coke, toner, and other polymeric and insoluble materials has resulted in a spectrum of generically similar pulmonary changes. Scaling of the rat-data, using 1 mg dust/g of lung as a threshold for "lung overloading", normal ventilatory parameters and 300-500 day human retention halftimes project "lung overloading" upon long-term exposure to insoluble respirable dust above 0.2-0.33 mg/m3. Other types of information are consistent with an exposure limit in the range of 0.4-1.0 mg/m3 for respirable dust. The absence of definitive toxicology data in support of the current OSHA respirable nuisance dust limit, coupled with the lack of sound epidemiology data, indicate the need to reexamine the validity of the current nuisance dust health standard.

Jan 1, 1991

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

The term unsaturated flow refers to the conditions in which soil pores are filled partially with water and partially with air. This situation constitutes a porous material filled with immiscible fluids. Immiscible fluids by definition exhibit a distinct interface when in contact with each other resulting in a pressure discontinuity across the interface due to surface tension. The fluid that tends to wet the solid particles is called the wetting fluid and the other is the nonwetting fluid. In the air-water sys- tem, air is the nonwetting fluid and water is the wetting fluid. A somewhat analogous fluid system is that in a petroleum reservoir in which there may be up to three immiscible fluids-oil, salt water, and natural gas. The degree of wettability of the fluids is salt water + oil + gas. That is, salt water is the wetting fluid at any interface and oil is the wetting fluid at an oil-gas interface but is the nonwetting fluid at a water-oil interface. Since the oil-water and air- water systems are similar, much of the theory and methods of analyses for the air-water system or unsaturated flow have evolved through the petroleum literature. In the two-fluid system the pressure difference across an interface depends on the size of the pore space in which the interface occurs. The equation commonly used to represent this dependence is

Jan 1, 1986

By Jianliang Li, Xi Peng, Guorong Lu

This paper describes a wind tunnel modelling study on airflow and pollutant concentration in a large open pit mine located in southern China. A 1:1000 scale model of the surface mine was fabricated and placed in a wind tunnel. Airflow velocity in the test section of the wind tunnel can be continuously adjusted from 0.3 m/s to 6.0 m/s. In this study, both the present pit condition and the projected final pit condition were modelled. Airflow direction and velocity profiles were obtained. The impact of wind direction and velocity on airflow was also studied. The pollutant concentration distribution in the open pit was simulated using smoke, and a distribution profile was obtained. Respirable dust concentrations at the mine site were measured, then its concentrations at the projected final pit condition were estimated using the obtained pollutant concentration distribution profile. Under normal meteorological conditions, it is expected that the respirable dust concentration for the final pit should not exceed the government standard. Key Words: wind tunnel modelling, open pit, surface mining, scale model, mine ventilation, airflow profile, point source.

Jan 1, 1993

By M. J. McPherson, P. Mousset-Jones, G. Danko, I. M. Konduri, F. Calizaya

The flow of air through a bare rock, underground airway is affected by the size and frequency of the micro- and macro-asperities of the rough surface. The focus of this study is to characterize this roughness using the coefficient of friction, the friction factor and the fractal dimension. The rough surfaces were identified by two sets of 177 and 78 transverse profiles measured by a laser profiler along the length of an underground airway at the Sunshine Mine, Idaho. Based on the measurements, the friction factor was estimated from Atkinson's equation and the coefficient of friction from the Moody diagram, and power series and variography were used for computing a fractal dimension for the airway surface. Significant conclusions are reached relating the effect of error in the measurement of cross-sectional and rubbing surface areas of a drift, on the estimated friction factor for the airway. The potential role of the fractal dimension as an alternative to the friction factor for characterizing the roughness of a mine airway is discussed.

Jan 1, 1993

By John R. Alderman

INTRODUCTION Uranium production in Saskatchewan commenced in 1953 at the Eldorado Nuclear Mine in the Beaverlodge area of North Saskatchewan, followed over the next few years by nineteen other smaller mining operations. Production rose from 453 600 to 2 721 600 kg per annum by 1958. Over supply and the closure of the U.S. market to Canadian uranium led to an equally rapid decline to an annual production which varied from 453 600 to 907 200 kg during the years 1964 to 1976. As the market strengthened there was a resurgence of the Saskatchewan industry and production from one underground and two open pit mines reached 2 809 000 kg in 1980 for a return of 232 000 000 dollars (Table 1). [ ] Future output will be increased by a potential 4 536 000 kg from the Key Lake mine now under construction and scheduled to commence production in 1983. Three other projects are in an advanced stage of planning and are currently in the regulatory review process. Several other prospects are being evaluated. Figure 1 shows the location of uranium mines in Saskatchewan and table 2 illustrates the present status of uranium mines and deposits in Saskatchewan. At present, 1400 workers are employed directly by the uranium mining industry. Saskatchewan has diverse mining conditions ranging from the underground mine of Eldorado Nuclear Limited 1870 m deep with a low grade ore deposit averaging from 0.17-0.20% with extraction by carbonate leaching through an open pit operated by Gulf Minerals with a moderate depth of 110 m and 0.37-0.5% grade to the shallow open pit of Cluff Mining 30 m deep with a very high average ore grade of 7.0%. Extraction at both open pits utilizes acid leach extraction processes, the latter after gravimetric separation. Key Lake mining has commenced development of the next mine to come into production which will have two open pits to a depth of 80 and 175 m with an average 3% grade to be followed by the deep open pit of Canada Wide Mines Ltd. to a depth of 205 m with average grade of 1.5%. One existing mine will transfer its operation from the present open pit to another of slightly higher average grade. From the Saskatchewan point of view, these different operating conditions present varied problems associated with occupational health and safety and concomitant solutions to reducing worker exposure to radiation. REGULATION OF URANIUM MINES In 1946, the Government of Canada enacted the Atomic Energy Control Act which established the Atomic Energy Control Board (AECB) and vested in that board powers to regulate all aspects of uranium mining. In the early years, the AECB relied heavily on provincial occupational health and safety agencies for the surveillance of uranium mining and milling operations, particularly with regard to "conventional" occupational health and safety, the AECB having at the time no specific regulations covering the mining industry. There has been considerable dialogue between federal and provincial governments since, traditionally, occupational health and safety is a provincial jurisdiction. Saskatchewan is strongly opposed to the intrusion of the federal government into the provincial domain, particularly as there is little justification for a federal presence in occupational health and safety in uranium mining and milling. The province does, however, accept the need for federal involvement in certain aspects of the regulation of nuclear industry, for example, international sales, research and security of reserves for domestic use.

Jan 1, 1981

Iron removal using biologically produced organic acids was investigated as a treatment option for the production of refractory grade bauxite. A leaching solution containing 98mM oxalic and 67mM citric acid was prepared from a sucrose medium using for fermentation the fungus Aspergillus niger. The leaching capacity of the metabolic product was tested on a bauxite sample containing 16-19% iron (as Fe2O3). Aerobic and anaerobic leaching tests were carried out at pH 3, 50°C and in the absence of light. Results indicated that more than 80% of total Fe can be removed from the ore, when the sample is finely ground and applying a two stage treatment. The presence of oxygen was seen to affect adversely the kinetics of the process, suggesting that leaching follows a reductive mechanism. Experiments that were carried out with different particle size fractions indicated that for obtaining efficient removal of iron particles should be finer than 0.63mm. Reduction of the operating temperature from 50 to 30°C caused a decrease in the rate of iron removal by more than 10 times.

Jan 1, 2006

By Jerry Motz, Gary Beerkircher

Through a joint development program between Bougainville Copper Ltd. and Nordberg, the MP1000 cone crusher was developed and tested during the period between 1982 and 1987. Following the successful prototype work, Bougainville Copper purchased an additional three units. A second prototype was installed at Codelco's El Teniente three-stage crushing installation in 1987. This unit has also completed its test period successfully. Three additional units have been sold, making a total of eight, to date. Two will be installed as secondary crushers in two separate large aggregate plants. The first unit installed in the United States was placed in operation at Vulcan Materials Company's plant in Manassas, Virginia in early May 1990. Another unit is being installed in Asarco's Ray Mines expansion. The Asarco installation will be a SAG MILL pebble crusher producing approximately 600 STPH of nominal 1/2" product. Application Objective The objective was to design a crusher that would fit on an existing 7' foundation with minimum modification and yet be capable of twice the production of a 7' cone crusher. In meeting this objective, the power rating was increased to 1000 HP, the maximum design crushing force was increased and the crusher was designed to operate at higher speeds than an equivalent size Symons cone crusher. The MP1000 was also designed to operate with a head throw that can be greater than the equivalent Symons design. Design The crusher design criteria involved the recognition of the crushing force increase required for the finer crushed products than

Jan 1, 1992

By Andrew N. Peterson

US mine operators frequently fail to investigate more cost effective and productive bulk material handling systems because surface mines seem to lend themselves to truck ore haulage. In this country, as a result, use of conveyors to move heavy loads from mine to process facilities has been minimized, if not actually neglected. In contrast, there are more than 50 conveyorized surface mines in successful operation around the world. These mine operators have learned that properly applied conveyorized systems can offer major savings in capital and operating costs, which contribute to improved profits when combined with other proven mining technologies. Growing acceptance and application of conveyorized bulk material handling in surface mines also points up how unique each mine is and how careful planning contributes to maximum mine effectiveness. Because of these differences, mining executives and technical and operating staffs need to develop an understanding of three factors in applying conveyorized bulk material handling in surface mines: • Why each mine will benefit from the type of automation permitted by conveyorized operation, •What kind of equipment is available, and • What applications most effectively demonstrate the first two factors in action - hauling either ore or waste. The conveyorized systems considered in this presentation have production rates from 0.5-2.7 kt/h (500-3,000 stph). Worldwide, these systems have been operating since the early 1960s. Advantages of Conveyors Why do you want conveyorized bulk material handling? First, it almost always provides lower operating and maintenance costs. Second, it frequently requires lower initial capital costs and almost always requires lower capital costs over the life of the surface mine. Third, it provides comparable operating availability, and finally, it frequently gives comparable operating flexibility - depending on the mine plan. Cost avoidance can be accomplished with modern production methods. These, in turn, permit increased productivity and reduced operating costs such as those for energy, maintenance, and manpower. It has been demonstrated in European surface mines and elsewhere, that conveyor systems frequently require lower initial costs than does truck haulage. Almost always such operations require lower capital costs over the mine life. Those costs include the continual addition of haulage trucks to both accommodate the increasingly difficult haulage routes and fulfill replacement requirements when trucks wear out. Conveyor systems handling ore in numerous large crushing and port facilities, which have operated since the early 1950s, have clearly demonstrated a useful conveyor life of more than 25 years. In contrast, off-highway trucks have life spans of six to eight years. The following examples illustrate comparative capital costs to purchase conveyor systems and comparable truck haulage units. Example 1 The ore haulage route from point A to point B is level and 610m (2,000 ft) long. The material weighs 1.8 t/m3 (110 lbs per cu ft) and must be transported at a rate of 1.8 kt/h (2,000 stph). The installed capital costs to provide a properly designed conveyor that will transport the described material from point A to B is about $450,000. The capital cost to purchase three 77-t (85-st) off-highway trucks and one spare truck - which would provide equivalent capacity - would be about $1.2 million. The truck cost estimate is based on a 6 min. or 771 kt/h (850 stph) truck cycle time. Truck efficiency is estimated at 0.8. Each 77-t (85-st) truck would have an actual haulage rate of 617 kt/h (680 stph). Therefore, three trucks would be necessary to transport the designated tonnage of 1.8 kt/h (2,000 stph). A movable crushing plant would be located at point A for the conveyors and a permanent crushing plant at point B for the truck haulage system. Capital costs for these primary crushing plants were not included in the calculations for either system because the capital costs are frequently comparable. Example 2 The transport route from point A to point B is 610 m (2,000 ft) horizontally and 122 m (400 ft) vertically - on a 20% grade (Fig. 1). The material weighs 1.8 t/m3 (110 lbs per cu ft) and must be moved at a rate of 1.8 kt/h (2,000 stph).

Jan 6, 1983

By John A. Herbst, Per-Olaf Samskog, Jan Bjorkman, Yi Chang Lo

The study described here is concerned with the development of a model for the simulation and optimization of a rod milllpebble mill circuit. A series of laboratory tests was conducted to investigate the effect of critical operating variables. The pebble mill model structure used was a population balance type which accounta for impact breakage and fines production from pebble wear. Model parameterization and verification were performed using data obtained in batch tests and in a plant test campaign. The model developed was found to provide accurate predictions of plant-scale grinding. The prediction errors are less than 21 for the product size and 31 for the mill power required to achieve that product with a mill volume scale-up of 14,000 to 1. These results indicate that accurate extrapolations from laboratory batch data to plant scale performance are poaaible using the population balance model and scale-up prediction procedures for pebble mill circuits as well as ball mills. Preliminary indications are that the new model developed here will provide vital input for pebble mill optimization and process control as well as pebble mill design.

Jan 1, 1990

By W. J. Wuest, M. J. DeMarco, J. R. Koehler

Yield-pillar-based gate-road designs often fail to provide mine operators with adequate ground control. This is because of the unknowing m is employment of "critically" sized pillars. A "critical” pillar is one that is too large to either yield nonviolently or to yield before the roof and floor sustain permanent damage and is too small to support full abutment loads. To demonstrate the critical pillar concept, afield study of a tapering gate road was conducted by the US Bureau of Mines at the Sunnyside Coal Mines, Sunnyside, UT. Extreme pillar stresses and associated coal bumps characterize the response to first panel mining of a 16.8-m (55.1-ft) wide critical design. Significantly lower pillar stresses, early yielding of the pillar and adjacent panel rib, and an absence of coal bumps suggest that a narrower 12.2-m (40-ft) wide design more closely approaches proper yield pillar dimensions. A probe hole in several 10.6-m (34.8-ft) wide pillars revealed low stress levels and substantial pillar and panel rib yielding prior to abutment onset, indicating a properly functioning yield pillar design.

Jan 1, 1997

By William Mayrsohn

INTRODUCTION The net smelter return (NSR) calculations made on various smelter-refiner alternatives, is usually done by taking the gross metal value less metal deductions, smelter-refiner charges and transportation-insurance costs. An important further step in the calculations is the financial charges imputed by refining time, transportation delays, to the time gold is returned (if tolled) or paid for if bought by the refiner. Metal that is priced in advance of payment also requires a forward premium that the mine may or may not be getting. This is a hidden cost that is often omitted from the calculations. Segmentation of each of the above costs are vital to see where improvements can be made through negotiations, better choice of metal pricing or transportation alternatives. BASIC NSR CALCULATIONS SMELTER-REFINER CHARGES Payable Metals - Treatment Refining Charges: Since these can vary from one smelter to another, the effect of a $25 per ounce price increase/decrease should be calculated and kept in mind. This effect is compared to differences in refining/treatment charge and a total smelter/refiner cost on a cost per ounce of dore, gold, or a dry ton of concentrate. Transportation - Insurance: Research has to be done on transportation alternatives perhaps with some assistance from the smelters, railroads or trucking companies. For dore, the U.S. postal service is viable and inexpensive. For railroads, deal with the regional office ie., B&N El Paso, if production warrants regular bulk shipments. Spot rates tend to be higher. A small mine producing concentrates usually goes to the nearest smelter due to extremely high transportation costs. Here is where a refiner can help. Insurance-Quotes for rates on dore/ concentrates are easily obtainable, but the mine has to shop for the best rates. Your insurance broker is the best source of this information. Differences between Mine and Smelter/Refiners Metal losses can be caused by the following: - Concentrate dry weight differences caused by moisture determination or faulty scales (often at the mine). - Sample bias by poor preparation practices on either side ie., grinding of the sample removing gold; or a poorly mixed concentrate. Dore bullion has no such difficulties.

Jan 1, 1987

By Jan K. Wolski

This paper presents a simple method for determining air temperature gradients along a tunnel with time-dependent intake air properties. The variable intake air properties may be the flow direction and rate, air temperature and humidity, and the carbon dioxide content. The transient intake air properties are simulated by a series of steady intake air conditions, using a superposition method. The mathematical model was verified by reduced-scale experiments in the Waldo mine. The method may be usefull in conditions where the computation speed is more important than the precision of the results.

Jan 1, 1995

By G. Gregory Bryan

G. GREGORY BRYAN 33.1-DEFINITIONS 1. Engineering-Engineering may be defined simply as the science and art of making materials and energy serve the purposes of mankind. There are many fields of engineering, each with specialized branches. The basic fields are: civil, mechanical, electrical and chemical. Others may be specialized phases of one of these or a combination of two or more of them. Mining engineering, for example, requires a working knowledge of construction, mechanics, electricity, metallurgy, geology, accounting and economics, in addition to mining techniques or methods. Mining geology and geological engineering involve rocks and their structure and the forces that act upon them, and chemistry, physics and mechanics also are important to practitioners in these fields. Metallurgical engineering requires chemistry and physics but also a working knowledge of mechanics and electricity. The fields of engineering with which this section are concerned are those primarily involved in the finding, mining, processing and utilization of the natural mineral resources of the earth. They include geology, mining engineering and metallurgy, each with several specialized subdivisions. 2. The Engineer-The engineer is an individual who has had specialized training to qualify him to practice the profession of engineering. The term engineer is used here in its broadest sense to include such professions as certain applied branches of geology, physics and chemistry, whose practitioners may not have acquired engineering degrees but who, nevertheless, by training and experience in some area of the minerals industry, qualify under the above definition. To become a successful engineer, a person should have certain pronounced aptitudes and preferences, even before embarking upon the requisite course of training. Among the chief qualifications that the engineer-to-he must possess are: (a) a natural inclination toward mathematics and the physical sciences, and (b) intellectual and moral honesty. Among the qualities the student engineer must develop are: (a) independence of thought, open-mindedness, sound judgment and the ability to observe, deduce, plan and apply solutions, (b) the ability to articulate and convince, and (c) the ability to organize and direct the activity of others. These qualities are normally developed through sound education, followed by pro¬gressive experience. 3. The Consulting Engineer-The consulting engineer is an engineer primarily in the business of furnishing specialized advice and engineering services on an intermit¬tent, part-time or occasional basis to clients who do not have continuous or full-time need for those services. He must be able to make and take full responsibility for decisions and recom¬mendations where original thought is required. He should not be confused with the engineer, surveyor or technician who also may furnish services on an intermittent basis, such as regular periodic surveys, the overseeing of drilling programs and other such routine work as may require little original thought and planning. 33.2-THE PROFESSION OF CONSULTING ENGINEERING 33.2.1- QUALITIES REQUIRED A successful consulting engineer must first have, as a minimum, the qualities and preparation of an engineer. In addition, he must have sufficient knowledge,

Jan 1, 1973

By J. D. Miller, C. L. Lin

As the techniques and resolution for three-dimensional spatial analysis have advanced in the last decade, it is now possible to map in detail the microstructure of multiphase materials in three-dimensional digital space. In this regard, three-dimensional X-ray microtomography offers a unique imaging capability. Spatial resolution on the order of 10 µm can be achieved with the use of microfocus X-ray generators. Recently, a state-of-the-art cone beam X-ray microtomography system was installed at the University of Utah for the quantitative analysis of multiphase materials in three dimensions. This facility was designed to obtain 2,048 by 2,048-pixel reconstruction over a 10-mm diameter, while also allowing for the imaging of somewhat larger (40-mm) objects. The system is capable of handling high-density materials, even materials having a density as high as 8.0 g/cm3. This unique, one-of-a-kind instrument can be used to obtain three-dimensional spatial reconstruction for many applications, such as three-dimensional liberation analysis, pore structure analysis of particle beds during filtration and the air-void system of concrete structures. The utilization of X-ray microtomography not only allows for the quantitative analysis of multiphase systems but also allows, for the textural characterization and determination of phase continuity. This paper presents information regarding the current use of this new facility and a review of potential applications for the advanced analytical system. Key words: Cone beam X-ray microtomography, Three-dimensional analysis, Material characterization

Jan 1, 2003

By Tracy E. Barnes

The process of computerized reserve estimation consists of assuming a model of the spatial continuity of a deposit and then, working within the assumed model, estimating the grade and tonnage of the deposit. The basis for geostatistical analysis is a statistical definition of this continuity or the variogram. The true variogram is essentially impossible to define. So an experimental variogram is defined from the sample data and kriging proceeds assuming that this experimental variogram is a good approximation of the true variogram. Therefore the closer the experimental variogram is to the true variogram the better the estimate will be. Realizing that the variogram is nothing more than the definition of the spatial continuity of the deposit, taking the known geologic continuity of the deposit into account will give a superior variogram. The use of known geology is a widely recognized necessity for good structural analysis of a deposit, but is rarely fully incorporated into the structural analysis.

Jan 1, 1982

By Scott Jones, Tony R. A. Durkac, David W. Stephens

Williams Mine was considering spending a large amount of capital to install a pumping system underground to rid the mine of slimes; with the introduction of Delvo Stabilizer this challenge would be met. This paper will address the process to introducing the Delvo Technology into the Cemented Roddill Program. Some areas that were worked on included: testing of the mix design, bet dosage rate for an economical solution, temperatures at different levels of the mine, the amount of time needed once the grout was put to sleep, preventing plugs in the delivery system A controlled dosage that gave the Operator a 12 hour window of flowable material was developed and an added feature to the introduction of Delvo was also noted; an increase in MPa strength over the design. Collectively with the help of operators at the batch plant and the underground miners involved with the program all concerns where addressed and all goals met. A cost effective program was implemented thus increasing productivity, preventing costly plugs, and improved the actual maintenance on the entire delivery system, which included the batch tanks, aggitator tanks, all pumps required in the system and of course reducing buildup of slimes at the cones.

Jan 1, 2001