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By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Variations of cut-and-fill stoping are the most com¬monly used methods for mining narrow, nearly vertical veins in the United States. These methods offer the best results for maximum recovery with minimum dilution of ore from veins that are irregular along either the strike or dip. The vein is developed along the strike length in 61 to 122-m (200 to 400-ft) blocks or stopes and vertical lifts of 48 to 61 m (160 to 200 ft). The level development is either on the vein or parallel to the vein with crosscuts intersecting the vein at right angles. Chutes, manways, and haulageways are established in the center of the stopes at the lower level of the ore block. Mining then proceeds by blasting down horizon¬tal cuts of the vein in both directions from the chutes. As the blasting continues to the outer limits of the stope, the broken ore is slushed into chutes. When a complete horizontal cut has been made, the service raise, including chutes, is raised to a height of a few feet from the back and the stope is filled with either sand fill or waste rock. Then the mining cycle is re¬peated until the stope reaches the level above. LEVEL DEVELOPMENT PRACTICES FOROVERHAND CUT-AND-FILL STOPING The two basic methods used are: the development of the haulage drift within the vein, or development of the haulage drift paralleling the vein in the foot or hang¬ing wall with crosscuts intersecting the vein perpendicu¬larly at 61 to 122-m (200 to 400-ft) intervals. The il¬lustrated methods of development show some of the variations which can be used to develop a stope block. Development for cut-and-fill stoping does not require the more elaborate schemes outlined for shrinkage or sublevel long-hole stoping. Many factors must be considered before deciding on a particular development plan and usually the nature of the vein or host rock will dictate which specific method to use. If the vein is relatively straight, then develop¬ment on the vein is feasible. Excessive ground pressures will necessitate the leaving of a sill pillar (Fig. la) and highly fractured or "heavy" ground will require timber¬ing (Fig. 1b). Development on the vein has the advan¬tage of reducing capital outlay and waste handling prob¬lems since the mined rock has value. But removal of the sill pillar is costly and time-consuming. Also, where no sill pillar is left, excessive timbering and subsequent level repair is a drawback. Development in either the footwall or hanging wall is the more common procedure and offers greater flexi¬bility in overall development of consecutive stoping blocks by partially disassociating the level development from stoping (Fig. 2). Chutes, manways, and service compartments com¬prise the next stage of development and are established in a variety of ways. They provide access to the stope, ventilation ducts, and muck removal. Several different development systems are in common use and average costs for each have been developed. Raise borers are used to establish ventilation and ac¬cess from level to level. They are normally reamed with 1.5 or 1.8-m (5 or 6-ft) cutterhead and may either be left bald or lined with a 1.2-m (4-ft) diam corrugated liner plate, depending upon ground conditions. The use of an Alimak raise climber is another effec¬tive method of driving a raise from level to level. The size usually is 2 m (7 ft) square. This will accommodate square-set timber which is later installed. This method of raising is not common, although it is a valuable tech¬nique especially when level spacing is several hundred feet.

Jan 1, 1982

By K. Schonert

Jigging is a separtion process that has already been described by Agricola in 1556. Fine size material is normally treated in a hutch jigging process, in which the heavy particles are penetrating the jigging bed and are recovered in the hutch. In this process, a ragging consisting of the same or higher specific gravity particles is inserted upon the screen. The ragging is used mainly as a valve to maintain a constant upward flow of the water. Since the heavy particles should pass easily into the hutch, the ragging particles are up to ten times larger than the maximum feed size (1). In the fine size range - 1 mm the performance of jigs is mainly insufficient; some of the low density particles percolate the ragging while the hutch water flow prevents the fine heavy particles from entering the bed (2). By giving a thin layer of particles, which posesses a density sensitivity in case of certain adjustment of flow conditions, directly on the screen, the precision of separation can be increased. Stroke length and frequency have to be adjusted to give this filter layer a homogenous porosity between 60 and 70% during the pulsion phase, and the "en masse" lifting of the bed, as it is known in jigging of coarser material, has to be prevented (3). In this case, the jig cycle is only of minor influence.

Jan 1, 1987

By D. Hodouin

Abstract. Dynamic experiments were performed in a continuous open-circuit 40x40 cm ball mill using a pseudo-random binary sequence of the feed rate and measuring the variations of the discharge particle-size distribution. The impulse response is calculated by a cross-correlation technique. Then a model involving a discrete transfer function and a time-series stochastic equation is calibrated using maximum likelihood methods. Based on this empirical model. a minimum variance controller with constrained input is designed and its performances are compared. by simulation, to the performances of a conventional PID regulator.

Jan 1, 1987

By George H. Tousley

As a member of the vast fraternity of those that supply and service the mining industry, I can assure you that we are amused, amazed, and horrified at the poor drilling and blasting practices of a great number of operations. Yes, ?horrified?. After all, that is our product you are using and abusing. We all know that drilling and blasting is expensive. In some operations D & B costs may exceed excavating or hauling costs, or both. At any rate, it is the only portion of the whole operation that could be said to directly affect other costs. There is no argument that it affects excavating costs. If the overburden and/or product is drilled and hauled, it affects haulage costs. It can easily affect milling costs, if more or less fines is important. And poor fragmentation affects crushing costs. If scalping is done ahead of the primary crusher it can affect production. As has been pointed out, the D & B operation itself is costly, and therefore should lend itself to close scrutiny. Anyone that could immediately decrease haulage costs by 10% would be a hero with any mining company.

Jan 1, 1976

By J. C. Rhodes

The demand for good white minerals for many industries including paint and plastics have made it necessary to import high grade industrial minerals from various parts of the world. Some of the calcium carbonate, barite and talc minerals are from China, Mexico, Jamaica and Haiti. These materials can be processed by custom processors. The custom processor can provide a service to the Industrial Mineral owner. The imported materials can arrive in various feed sizes from lumps (3"-6") down to 325 mesh. The materials can also be in various size shipments or container loads. The economics for importing should be investigated thoroughly. The two factors just mentioned can influence the quantity of material to be processed. These two factors can also influence the total cost for each end product.

Jan 1, 1989

By William M. PhD. Schafer

The net acid generation (NAG) pH method is an important analytical tool that supplements static and kinetic tests for assessment of the acid generation risk of rock samples. This method is useful because it is simple, rapid and cost-effective; it combines features of the static and kinetic tests; and it can be conducted at mine assay labs. The NAG pH test is particularly effective for operational testing programs used to classify, selectively handle, and route potentially acid-generating waste rock. The NAG pH procedure is based on a 24 hour oxidation of a pulverized rock sample with hydrogen peroxide and subsequent measurement of the sample's pH. If the NAG pH is below a critical value, determined empirically, then the sample has the potential to generate acid in the field. The critical NAG pH value is typically within the range of 3 to 4.5, but the exact relationship between NAG pH and potential acid generation should be determined individually for each mine site, based on comparison with results of acid base accounting, humidity cell tests, and mineralogical analyses. Calibration and use of the NAG test for predicting ARD risk at several mines are described in this paper.

Jan 1, 2000

By W. E. Adkins

Ore and waste haulage usually constitute one of the major cost items in the operation of an open pit mine. This has always been true at the Chino mine, one of Kennecott Copper Corporation's western mines, located in southwestern New Mexico. The Chino Mines Division is an integrated operation. The open pit mine, dump leaching operation, and a precipitating plant are located at Santa Rita, New Mexico, and a power plant, mill, smelter, and fire refinery are located in Hurley, New Mexico, about nine miles south of the mine. Approximately 22,000 tons of ore are mined and processed each day; in addition, approximately 42,000 tons of waste are removed from the pit. Originally the pit was located along the side of Santa Rita Creek but it has deepened and expanded so that at the present time, it could better be described as a nearly circular pit about one to 1.5 miles in diameter. The bottom is at an elevation of 5750 feet, the west rim near the ore gathering yard is at an elevation of 6230 feet, and the highest bench on the east side of the pit is at an elevation of 6750 feet, giving a vertical mining depth of 1000 feet. The train waste dumps reach a maximum elevation of 6580 feet. At the present time, Chino uses electric locomotives, off-the- highway trucks, and a skip system for mine haulage.

Jan 1, 1963

By S. Vivek

An atomic force microscope (AFM) was used to measure the surface forces between silica plate and glass sphere in sodium oleate solutions. Hydrophobic forces were observed when the two surfaces were activated by CaOH+ or MgOH+ ions. The force measurements conducted in the absence of the surfactant showed that the activated silica surfaces do not exhibit the primary hydration forces that have been observed between unactivated surfaces. The activation mechanism was studied by calculating the contributions from electrostatic and chemical interactions to the total free energy change. The results suggest that the adsorption of the hydroxo-complexes is controlled mainly by the lateral interactions between the adsorbed species on the silica surface.

Jan 1, 1999

By William F. Lawrence

Research performed by the West Virginia University Coal Research Bureau under the sponsorship of the U. S. Bureau of Mines has demonstrated that coal-ash can be converted into mineral wool insulation of a quality equal to commercial grade mineral wool. An unexpected benefit has been gained due to the discovery that the usual and costly fluxing agents employed when using conventional raw materials are not required when producing mineral wool from coal-ash. Because coal-ash sources are commonly located close to mineral wool marketing areas and because mineral wool is seldom shipped far as a result of its low weight and high volume, the use of coal-ash may offer the mineral wool producing industry profitable opportunities. Moreover, coal-ash produced from experimental limestone based air pollution control processes also yields high quality mineral wool which may aid in retarding corrosion of metals contacted by the wool. In addition, internal fluxing effects and possible sulfur oxide utilization could provide economic advantages to the use of these "modified" ashes.

Jan 1, 1969

By M. K. Boateng, K. Awuah-Offei

"Mining companies have started using quantitative tools, including computer models of community interaction, to gain intelligence on social risks surrounding their projects. Models of changes in community preferences regarding mining projects over time are useful for evaluating changes in project risk due to changes in the social license to operate. For example, agent-based models that use information diffusion models and social networks are useful for studying those changes due to information diffusion. However, such agent-based models are sensitive to many input parameters including the parameters of the diffusion model and the average degree of the social network. This work evaluates the sensitivity of such a model to diffusion and network model parameters (probability of imitation, probability of innovation, and average degree) using the first order and total sensitivity indices. The results show that the model is much more sensitive to the probability of imitation than the other two parameters. Thus, to reduce uncertainty surrounding the model’s predictions of community acceptance of mining, mines need to obtain accurate estimates of the probability of imitation. INTRODUCTION Mining companies have started using quantitative tools, including computer models of community interaction, to gain intelligence on social risks surrounding their projects. Models of changes in community preferences regarding mining projects over time are useful for evaluating changes in project risk due to changes in the social license to operate. For example, agent-based models that use information diffusion models and social networks are useful for studying those changes due to information diffusion. However, such agent-based models are sensitive to many input parameters including the parameters of the diffusion model and the average degree of the social network. This work evaluates the sensitivity of such a model to diffusion and network model parameters (probability of imitation, probability of innovation, and average degree) using the first order and total sensitivity indices. AGENT-BASED MODEL OF COMMUNITY ACCEPTANCE A few researchers have started using agent-based models to study mining systems (Fujiono 2011; Nakagawa et al. 2013; Bahr 2015; Boateng and Awuah-Offei 2017a, b). Nakagawa and co-workers used ABM to evaluate dynamics of a mining project’s stakeholder networks whereas Fujiono used ABM to study innovation diffusion in the mining industry. Our previous work has presented models and a framework for modeling information diffusion in a mining community and its impact on level of acceptance (a measure of SLO) of the mining project (Boateng and Awuah-Offei 2017a, b)."

Jan 1, 2018

By Daniel McNamee, Wern-ping N. Chen

The Amtrak East River Tunnels include four tunnels, Line 1 to Line 4, which extend from New York Penn Station in Manhattan, under the East River to Sunnyside Yard in Long Island City. Since fall 2015, Amtrak has initiated a program to repair and reconstruct the century-old tunnels, including structural repairs of all the tunnels and reconstruction of the Line 1 and Line 2 tunnels. The reconstruction includes bench wall demolition and replacing and installing the following elements in the tunnels: direct fixation track, traction power, signal, communication, and tunnel fire/life/safety components. This paper provides design and site investigation challenges and knowledge gained from this program.

Dec 1, 2022

By Kendall Y. Keuhey

The Mercur Mine, developed and operated by Getty Mining Company, is designed to process 3,000 tons/day of gold ore from the historical Mercur Hill mining district. The mill process involves crushing, grinding, and cyanide carbon in leach for recovery of gold ores ranging from 0.08 to 0.10 oz/ton. During the mining of normal grade ore, a substantial tonnage of material containing 0.02 to 0.05 oz/ton is mined and would be available for heap leaching. Also, Getty Mining is developing methods to test and prepare lower grade material for heap leaching by rejecting a course rock fraction from the grinding circuit SAG mill. This lower grade fraction is suitable for heap leaching, and its removal will upgrade the regular mill feed. Other test methods are directed toward leaching techniques on some 4.0 million tons of tailings that resulted from roasting and/or cyanide leaching Mercur ores in the early 1900s. The natural carbons associated with the Mercur deposits require special treatment to reduce their gold adsorption effects when leached with cyanide.

Jan 1, 1983

By James R. Boulton

Fire detection using Carbon Monoxide sensors is based on the assumption that a rise in the level of Carbon Monoxide in the atmosphere over a fixed, predetermined ambient level is indicative of a fire. The method is simple, reliable, and effective in those mines where the only source of CO is from undesired combustion. Ambience for this purpose is defined as that level of the gas that is present on a day to day basis whose source is not a fire. In a diesel powered mine the level of CO can and does vary so widely from hour to hour and minute to minute that the selection of a predetermined, fixed level represents a compromise between sensitivity and reliability. The compromise is a best of both and worst of both situations and the computer system often resembles the little boy who cried wolf instead of a state of the art early fire monitoring and detection system. Cyprus Coal personnel have developed method of determining the ambient level of CO in the atmosphere on a real time basis and using that dynamic ambient level to determine whether or not a fire exists in the mine. The method has been given the name of Wave Trend Crossing.

Jan 1, 1991

By Pramod C. Thakur

During the past ten years, the US coal industry has experimented with various ways to extract methane from coal seams. Drilling long horizontal boreholes in solid coal from mine workings and vertical gob degas boreholes are becoming standard techniques for methane recovery from coal seams. Prime motivations for such work have always been safety and productivity in mines, but the commercial value of gas so produced is encouraging utilization. The paper reviews various ways to produce methane from coal seams and discusses some utilization projects extant in the industry.

Jan 1, 1984

By Louis A. Panek

INTRODUCTION Block caving is distinguished from other types of mining in its requirements that the mineral body to be extracted must cave after a relatively large tabular open¬ing is created underneath it, and that the caving must be predictable and controllable for both operating and safety reasons. A mineral body that caves readily may present special problems with respect to ground support in the mining access openings, where stability, rather than failure, is the desired objective. The complete ex¬traction of a caving block, which may measure 50 to 100 m on a side and 100 to 300 m in height, tends to create a large void, resulting in significant changes in the ground stress system and a migration of rock from all directions toward the void; these events begin very early in the life of the mine and lead to additional spe¬cial problems with respect to control of ground move¬ments in the mine working areas, in the rock adjacent to the caving stope, and on the subsiding surface above the mine. Thus, the mechanics of an undercut-cave operation involve a broad range of geotechnical subject matter, including the determination of the lithostatic earth stresses, the structural properties of a large jointed rock mass, the flow properties of the fragments after caving, the stability of tunnel-like openings, and the stability and displacements of the rock structure (1 to 2 km in size) that surrounds the ore body. The objectives are achieved by in-place measurement of rock properties, laboratory tests, theoretical structural analyses of the related prob¬lems, measurements of structural behavior in and above the mine, and comparison of theoretical predictions with actual performance in order to interpret and refine the predictive capability. Undercut-cave mining (Julin, et al., 1973) differs from most other methods in that the caving is an essen¬tial and inherent part of the excavation process and in¬volves the movements of millions of cubic meters of rock over distances of hundreds of meters, as compared to the gradual 2-m settling of the overburden typical of full¬caving longwall coal mining. Undercut-cave mining may involve a rock mass on the scale of a 1-km diam open¬pit mine, but mining is several times deeper than that in the open-pit mine; furthermore, a large portion of the rock mass is in a failed, rather than a stable, state. Ground-control problems are inevitable in a situa¬tion where ground movements are occurring on such a large scale, and where the rock mass is expected con¬veniently to cave at one location but to be stable at an¬other. To devise effective responses to these ground¬control problems, one needs to know first of all what is happening, not merely at the point where the symptoms are observed, but more broadly with respect to the be¬havior of the entire mine rock-mass structure over the total area of extraction, including the effects on the sur¬face as well as those underground. Monitoring structural behavior on this scale would require a sizable group of people, if they attempted to measure all the significant mining-induced changes of ground displacement and pressure, even on a 100-m grid of points. The practical approach is to conduct the mea¬surements activity by a judicious sampling, making in¬termittent, concentrated efforts in selected areas of lim¬ited extent to study problems of immediate importance; simultaneously a minimum level of surveillance is main¬tained over the total structure to delineate long-term be¬havior trends in the major structure, which provide the baseline reference data that are so essential for interpre¬tive purposes. Mine-structure behavior depends not only on the mine extraction configuration and sequence, but on the rock structural properties and how they change from one location to another in the mine. Increasing attention is being given to determinations of rock strength and de¬formation by performing mechanical properties tests in place and in the laboratory, and by drawing inferences from the mapping of geologic structure in mine openings as they are driven and from the logging of cores ob¬tained from exploration drilling that may be performed primarily for delineating ore reserves. Monitoring of mine-structure behavior and inter¬preting it in terms of the growing excavations and the structural properties of the rocks have significant im¬pacts on designing the major undercut-cave operations of undercutting and drawing the ore, supporting the ac¬cess openings, and controlling surface subsidence. The purpose of this chapter is to call attention to the com¬mon ground-control problems encountered in undercut¬cave mining, to relate them to the mining process, and to explain the geomechanical approaches to analyzing these problems with emphasis on methods that are most readily available to the mine staff. UNDERCUTTING, CAVING, AND DRAWING A distinguishing characteristic of an undercut-caving method of extraction is the objective that the rock mass fails of its own weight in a predictable and controllable manner, after it is undercut by excavating a tabular¬shaped void, horizontal or inclined. Induced or forced caving refers to procedures that may be resorted to when the ore will not self-cave over the desired span. Block caving, panel caving, and mass caving are terms used to denote the particular sequence of undercutting and ex¬tracting the total area of the ore body on a single min¬ing level. The discussion herein is not specialized to any of these variations; the word block as used does not imply anything as to sequence. CREATING THE UNDERCUT A common undercutting method is to drive a series of drifts and blast the intervening pillars. Theoretical structural analysis and practical experience show that ground-support problems in the undercutting drifts and in underlying access openings as well (the same princi¬ple applies even if undercutting is done by belling out drawpoints from the slusher drifts) are minimized by retreat-mining across the undercut slot according to a sequence such that the growing undercut area retains a

Jan 1, 1982

By Rafael Moure

On April 21, 1980 the Oil, Chemical and Atomic Workers Union (OCAW) submitted a petition urging the revision of the Mine Safety and Health Administration (MSHA) permissible exposure limit to radon daughters in underground mines. This petition asks for: 1. Annual worker exposure to be lowered from the current 4 WLM/year to 0.7 WLM/year; 2. Full economic protection for workers whose exposure exceeds the proposed 0.7 WLM/ year; and 3. The primary means of lowering workers' radiation exposures shall be environmental controls. I will present the scientific rationale for the lower standard requested in our petition to the Department of Labor. [1. Radiation Exposures to Uranium Miners] Two types of radiation exposures are present in the environment of uranium mines: A. [External Exposure] to gamma sources which until 1978 was considered as negligible compared with internal alpha radiation. R. E. Utting (1), in an investigation of three Canadian uranium mines, found levels of exposures adding to an annual dose equivalent of 3 rems (30 mSv)/year. He concluded that gamma dosages may be a major fraction of the maximum permissible dose for many individual miners. The U.S. experience has shown that there is significant gamma exposure in U.S. uranium mines and, therefore, it could contribute to the total radiation exposure to the lungs in combination with internal alpha radiation. Before ventilation improvements in uranium mines, the contribution of external radiation to total lung dosage was relatively small. After improvements in ventilation, the reduction in internal radiation exposure appears to be in 1980 as low as 1/10 of the 1950 dosages. However, ventilation improvements do not affect external gamma radiation. Gamma radiation remains at the same 1950's level, and it represents today a higher proportion of the total radiation burden to the miner's lungs. MSHA regulations (57.5-47) state that if gamma exposure is below 2.0 milliroentgens per hour, then workers are not required to wear gamma dosimeters and cummulative gamma radiation records need not be kept. Assuming an exposure of 1.9 milliroentgens per hour per 173 hours/month for 12 months,a miner would accumulate a dose from external radiation of 3.9 rems/ year. This dosage would, of course, go unrecorded. MSHA, therefore, should consider a downward revision of the external radiation permissible exposure limit of 2.0 milliroentgens/hour. MSHA should make clear that this exposure level, when added to the internal exposure, does not make for sound health physics -even though the measured current levels do not approach 2.0 milliroentgens/hour. B. [Internal Exposure] to alpha radiation is still the larger contributor to lung irradiation in uranium miners. The target organ is still the lung. The parameters to be considered in determining the biological radiation load appear in Table I for radon daughters present in the upper respiratory tract (URT). TABLE I RADIATION DOSE PARAMETERS IN THE URT OF URANIUM MINERS Parameter Uranium Underground Mining [1. Quantity of the Absorbed Radon daughters (Po-218, Nuclide Pb-214, Bi-214) deposit in bronchial and lung tissue. (Measured by air concentration, Working Levels) 2. Type of Radiation Emitted Alpha Particles 3. Energy of Radiation High-Linear Energy Transfer 4. Half-life of the Nuclide Ra 222, 3.8 days in the Organ Po-218 (RaA), T1/2 = 3.1 minutes Pb-214 (RaB), T1/2 = 26.8 minutes Bi-264 (RaC), Vi = 19.7 minutes 5. Selection Factor of the Direct inhalation with mine Organ dust to be deposited in the tipper Respiratory Tract 6. Mass and Form of the Organ Especially cancer sensitive, tracheobronchial region and bronchi* 7. Distribution of the Activity Dependent on deposited dust in the Organ distribution in the Upper Respiratory Tract 8. Chemical Nature of All compounds (RaA, RaB, RaC) Incorporated Substance biologically insoluble. (Only localized, irradiation) *See Reference (2)] The parameters of Table I depend, in turn, on the demography of the exposed population, especially: 1. Age 2. Sex 3. Health Status 4. Genetic Constitution 5. Lifestyle (Smoking, Diet) Some of these demographic characteristics of uran-

Jan 1, 1981

By Hsu-Hui Hsing

In several previous studies from this laboratory (1-4) the autophobicity concept of Zisman and co-workers (5, 6) has been extended to gas phase adsorption processes. Low molecular weight alcohols adsorb on silica or alumina to yield novel isotherms in that monolayer adsorption occurs at low relative pressures indicative of strong bonding, but subsequently multilayer adsorption is quite sparse and diminishes with molecular weight of the alcohol. Such isotherms fit no standard classification scheme, however, in the vacinity of monolayer coverage alcohol isotherms give an excellent fit to the Langmuir or B. E. T. , Type I model (7). Monolayer adsorption is relatively irreversible and it is thus possible to prepare these chemisorbed layers and perform subsequent adsorption studies thereon. All indications are that the adsorptive properties on such films is radically different from that on the underlying alumina (4, 8) and in one study finite contact angles for hydrocarbon fluids were observed. One perennial conceptual difficulty is that there is no observed upper bound to the adsorption isotherm which would be necessary for autophobic adsorption to occur. The cause for this has been studied, and it is currently believed that adsorption is complicated by reversible capillary condensation at contact points between powder particles. The evidence for this is that large particle powders give less "apparent" adsorption at high relative pressures and in one very tedious study adsorption on planar sheets, oxidized aluminum displayed similar reduced uptake at high relative pressures when compared with high specific surface area powders (3).

Jan 1, 1974

By John Newby, Ulf Gwildis, Fred Marquis

From the Brightwater Conveyance System in Seattle to the Second Narrows Water Supply Tunnel in Vancouver, BC, the design and construction of deep-lying tunnels in the Pacific Northwest often require the incorporation of deep portal shafts. Alluvium overlying marine and glacial deposits combined with the high regional seismicity is the common geologic setting. The latest example illustrating the development of shaft design for this scenario is the 40-m-deep shafts of the Annacis Island Wastewater Treatment Plant New Outfall System. The main components of the shafts include the initial lining, providing the excavation support system to be used for launching and receiving a 5-m-diameter TBM; and the final lining, designed to withstand all static and seismic loads and to remain viable for the intended design life. This paper presents design and construction considerations for these shafts against the background of and lessons learned from past project experiences.

Jun 13, 2021

By Y. C. Kim

This paper reviews the past 18 months of geostatistical work in a coal deposit performed jointly by the University of Arizona and the Homer City Owners. The objective of the study was to better predict the in-seam sulfur values of the deposit for the purpose of future emission control. Based on initially available 112 DDH information, variogram models were developed for Sulfur, Ash and Thickness for the entire deposit. The deposit was then kriged using 333 m x 333 m (1000 ft x 1000 ft) panel size. The estimation variance information obtainable from geostatistics was next utilized to improve the development drilling activities that were underway. The results of this effort are presented in this paper. In addition, this paper describes (1) the reasons for preferring kriging over the conventional isopachs, (2) the results of empirical validation of the kriged estimates using 19 additional DDH information, and (3) up-to-date experience on the use of conditional simulation of the same deposit for the future emission control strategy development.

Jan 1, 1980

By William F. Riggs

Inspiration Consolidated Copper Company, located in a semi-arid climate in Arizona, continually struggles with a water problem. This applies to quality as well as quantity. When the EPA considered installing a new centralized sewage plant to handle the combined waste from the municipalities of Globe and Miami, Arizona, they automatically assumed that He could use the sewage effluent for process water in flotation. The treatment plant was to be a tertiary system producing 1600 gpm of sewage effluent. Inspiration decided that this could possibly be a source of additional process water. It was decided to test sewage effluent from Miami Treatment Plant at the rate of 400 gpm since it was only a primary treatment plant. The 400 gpm was combined with the 5000 gpm of process water going to the Concentrator. After one week the results were disastrous. The recovery and final grade had each dropped approximately 5%. With the results from this brief run, it was concluded that sewage effluent from a primary treatment plant could not be used. Research was then brought into the problem to begin an evaluation of sewage effluent in flotation.

Jan 1, 1977