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By Curt Morgan

This paper will discuss the value of cream a common information model for the mining industry. Information models are a proven technique for quantifying diverse information in a non-proprietary and easily understood manner. Combining information models with emergent data exchange technologies such as XML and SOAP, the mining industry could substantially reduce the costs of sharing information, reduce dependence on proprietary information systems, leverage inexpensive PC desktop technology, and provide a solid foundation for simplifying mine information systems. The paper will provide an overview of information modeling, discuss practical and proven applications of the modeling technique in the mining industry, propose how such a model could be developed in a collaborative manner and discuss the benefits of such an effort.

Jan 1, 2002

By Kerstin Malmqvist

Simulation technique has been used for a feasibility study of a deep exploration project for massive, sulphides in an old mining district, the Skellefte field. An interactive computerized mathematical model of an ore field and a drilling procedure has been constructed. It is possible to represent varying geological structures. The model also takes in consideration economical factors such as drilling costs and ore values. On this model Monte Carlo simulations and probability calculations are done in order to estimate the outcome of different drilling strategies under varying geological assumptions. Under the specific conditions of the well known Skellefte field, it is found to be possible to even calibrate the mathematical model.

Jan 1, 1979

By Julie Mitchell

The U.S. Bureau of Mines is conducting research on the application of expert sys tens technology to mining equipment diagnostics and predictive maintenance, to increase mining productivity and efficiency. One of the greatest hindrances to mining productivity is the breakdown of equipment, causing lengthy and costly delays. A diagnostic maintenance expert system (DIMES) is being developed for the hydraulic system of a continuous coal mining machine. Through the use of such a system, the frequency of equipment failures and repair time should be significantly reduced. This paper describes the application domain, the functional requirements, and the problem approach for DIMES, the design and implement at i on of the knowledge- based system, the sensor work involved with the project, and the future plans for the continued development of this and other expert systems.

Jan 1, 1989

By Stanislaw Wasilewski, Zdzislaw Krzystanek

Ventilation process in deep mines features relatively low efficiency. Commonly used passive control by means of regulators and flow throttling leads to remarkable energy losses that could be lower if effective monitoring and remote control of main fan operating point were implemented. In the paper, some results of investigation carried out in Polish coal mines to evaluate real energy losses in ventilation process are described. Different methods for main fans performance monitoring and control are considered and compared especially those using fan drive speed control to minimize power consumption. On the other hand some criteria for power evaluation are analysed and algorithms for optimum location of controllers in ventilation network are considered. The results of investigation are being implemented in the first system for monitoring and control of main fan operating point in Miechowice colliery.

Jan 1, 1995

By M. Kaya

The main purpose of this study was to design a modular/mobile coal preparation plant with a capacity of 300 t/h for Seyitömer B3 vein (almost 70% of the whole SLI deposit). Firstly, coal washing properties and washability curves were determined by sink and float tests. Then, equipment sizing, selection, flow sheet development, economic evaluation, investment and operating costs were determined for coarse (-150+18mm), fine (-18+0.5mm) and slime (-0.5+0.1mm) coals. Economic evaluation of the proposed flow sheet showed the feasibility and profitability of the investment.

Jan 1, 2006

By Sidney L. Hansen

This report was written because of concerns of the possible error involved when stopping leakage is incorrectly assumed to be turbulent. It is common practice to use turbulent flow to characterize each airway branch for computer models of mine ventilation circuits. This report addresses computer methods to accurately model air-leakage paths. A graphical method is described for deriving the equations of leakage through stoppings. The significance of modeling leakage according to the proper flow regime (turbulent or laminar) is demonstrated by the following method. A hypothetical mine was developed to represent present-day conditions. The stoppings of the mine were chosen to leak according to laminar flow properties. The mine was represented by constructing two base models. One base model used turbulent leakage equations for each stopping, while allowing the corresponding resistances to vary. The other base model used the correct laminar leakage equations and held corresponding resistances constant. The mine circuit was then expanded to demonstrate future development. The expanded portion was designed with stoppings that were identical to those in the base mine. The mine in the future state was represented by two projection models. Each projection model was an extended version of its associated base model; one used turbulent leakage equations and one used laminar leakage equations for stoppings. The laminar projection was theoretically correct and is used as a basis for comparison of differences in airflow quantities. Rather than modeling every leakage path for stoppings, a more practical approach is also discussed. A good level of accuracy is maintained by combining leakage paths if the proper flow equations are used. This method was used on the hypothetical mine projection and the flows were very similar to the theoretical values.

Jan 1, 1989

By Henry T. Miller

INTRODUCTION The basic purpose of this paper is to suggest procedures and practices to insure that no employee working in a uranium milling operation receives exposure to radioactive, toxic, or other materials or agents that might produce a permanent, deleterious effect upon his physical health and well-being. This program is also designed to insure that each employee can carry out his assigned duties without risk to his health or to that of his fellow workers. The assumptions underlying this program are by necessity, conservative, with emphasis being placed upon prevention rather than correction or reaction to an undesirable situation. The limits specified-either as Threshold Limit Values (TLVs), Maximum Permissible Concentrations (MPCs), or Radiation Protection Guides (RPGs)--represent both standards of best practice and those of law and regulation. The total program is envisioned as a balanced combination of occupational hygiene and radiation monitoring. This includes surveying, air sampling, personnel dosimetry, bioassay, medical surveillance, epidemiology, and training--all backed by a thoroughly tested and evaluated set of emergency procedures. The program, as presented, is keyed to the results of monitoring, surveying, air sampling, medical surveillance, and epidemiology--it being obvious that no problem can result when no hazard can be identified. The lack of an identified hazard, however, does not negate the need for such a program and prudence suggests the need for routine surveillance programs in the best-designed facility. [The Model Mill] The model mill used in this paper is that presented by the U. S. Environmental Protection Agency in its study of the uranium fuel cycle1. The model mill processes 600,000 metric tons of ore per year into 1,140 metric tons of U308 as yellowcake per year using an acid leach process. Other characteristics of the mill such as location, tailings dam design, number of light water reactors supported, and geographical location are not pertinent to this paper. The usual remote western location of these facilities produce some difficulties in delivering the required occupational medical services and is not within the scope of this paper. The operating lifetime of most mills is commonly estimated as from 12 to 15 years in actual practice; however, the U. S. Environmental Protection Agency estimated a 20-year lifetime for its model. The principal steps in the acid leach process include :2 (a) Ore is blended and crushed to pass through a 2.5 cm (one inch) screen. The crushed ore is then wet ground in a rod or ball mill and is transferred as a slurry to leaching tanks. (b) The ore is contacted with sulfuric acid solution and an oxidizing reagent to leach uranium from the ore. The product liquor is pumped to the solvent-extraction circuit while the washed residues (tailings) are sent to the tailings pond or pile. (c) Solvent extraction is used to purify and concentrate the uranium. (d) The uranium is precipitated with ammonia and transferred as a slurry. (e) Thickening and centrifuging are used to separate the uranium concentrate from residual liquids. (f) The concentrate is calcined and pulverized. (g) The concentrate or yellowcake is packaged in 208 liter drums for shipment. This process is illustrated in Figure 1.3

Jan 1, 1981

The last twenty years of development in Process Mineralogy are reviewed. Since the milestone review by Henley in 1983, which projected the concept and suggested closer working relations between mineral processing and mineral science, the second generation (1983-2003) has made significant progress. Examples are to be found in improved sampling and high-confidence flotation testing, and in automated quantitative mineral measurement. The recent addition of end-member studies has provided a clearer view of the overall ore mixture milled from its constituent geological components. In all, better predictive abilities have resulted from these improvements to Process Mineralogy. These improvements have led to more significant business impacts in mineral resource characterization, and in optimization of existing concentrator operations.

Jan 1, 2004

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

1.1 GENERALIZED METHODS OF SUPPRESSING THE INFLOW OF GROUND WATER DURING THE CONSTRUCTION OF SHAFTS, DRIFTS AND TUNNELS Various methods are used to prevent or minimize the inflow of ground water into underground workings during their excavation. The two most common methods include freezing the saturated rock and grouting using cement, sodium silicate, polyurethane and/or other chemicals. Each of these technologies for combatting the inflow of ground water is effective only under specific hydrogeological conditions. For example, although the freezing of saturated ground is among the more universally adopted methods, it is designed to provide only temporary protection during construction. Before the saturated rock thaws it is necessary to emplace a waterproof liner which is labor intensive, time consuming, and expensive. Consequently, freezing is used only in exceptionally complex hydrogeological conditions, namely in those cases where the water-bearing strata consist of unstable rock or the ground water has an anomalous hydrochemistry. The classical grouting of saturated rock, carried out from the surface or from the face of the underground workings for the purpose of limiting the inflow of ground water during excavation, utilizes both cement and a variety of chemical grouts. Cement grouting has been regarded as the main method of combatting the inflow of ground water in fractured rock throughout the world. In such countries as the Federal Republic of Germany (FRG), Canada, the Republic of South Africa and Great Britain, cement grouting is the main method of reducing the inflow of ground water during the excavation of mine openings. In Great Britain and the FRG, cement grouting has been used in 80% of all shaft excavations. In the Republic of South Africa cement grouting has been used in almost 100% of all the shafts that have been constructed. The installation of grout curtains into permeable water- bearing strata significantly reduces their permeability and increases the rock strength. Grouting has the greatest effect in fractured sandstones, certain well indurated shales, fractured granites, fractured quartzites, and karstic limestones or dolomites. The following principal factors must be considered when assessing the expediency of grouting rock with cement: the geometry of the network of fracture openings, the saturated hydraulic conductivity of the fractured rock, the hydraulic head acting on the water in the rock, and the chemical composition of the ground water. On the basis of the geometry of the fractures and the thickness of each hydrostratigraphic unit, the characteristics of the cement grout are selected. As a rule, the cementing of large fractured zones with high ground water velocities is carried out using inert fillers (sand, mill slag, loam, loess, crushed limestone), special types of cement, setting accelerators, and high cement concentrations in water. Calcium chloride, soda ash, sodium silicate, sodium nitrate, amino alcohols, tin bichloride, trisulphate nitrate, and lumnite are used extensively as the setting accelerators for cement grouts. In the FRG and Poland, special cement injection compounds that embody a mixture of cement and active cement metal salts, water additives and binding substances are being used for the cementing of saturated zones with rapid ground water velocities. These reagents accelerate the rate of structure-forming reactions. In spite of the wide variety of additives used for cement grouts, the effectiveness of the method in large fractures below the water table is either poor or it leads to a very large consumption of cement due to the erosion of the grout through the cracks before it hardens. At the present time, a large number of cement types and brands are being produced by various countries. These variations permit cementing to be employed in a variety of geological conditions. However, both pure cement grout and grouts with fillers constitute unstable systems with a high water loss rate. Therefore during the grouting of finely fractured rock the cement grout's premature loss of a large amount of free water causes its consistency to increase, whereupon the grout solidifies. Consequently, it inadequately penetrates into the fine fractures of the stratum. In addition the high water loss rate in the cement grout causes unreacted cement particles to remain, which greatly reduces the grout hardness and the resulting rock strength characteristics are weakened. As a result, the binding properties of the cement are utilized perhaps up to 60%. The remaining portion is left in the grout as a filler. All these conditions significantly decrease the efficiency of the isolation effort both with respect to strength and cost. In order to expand the cement grouting method, re- search is being conducted to improve the quality of cement grout, to increase its range of application with respect to permeability control, to reduce its water loss rate and to increase its capacity to withstand the erosive and corrosive effects of poor quality ground water. For example, mixtures of bentonitic clay, carboxymethyl cellulose, aerated sodium sulfide, sulphated alcohol distillery waste, nitrolignin, gip-

Jan 1, 1993

By D. C. Mahon, P. G. Sagert, M. R. Legeyt

INTRODUCTION Public opinion, over the effect of industrial development on the environment, has required government to institute guidelines for the protection of the environment both during the development of and after the decommissioning of mines. In no other area of mining has this requirement been more consistently applied in recent times than in the development of uranium mines. Uranium mines have been a focal point because of the very real uncertainties associated with the detrimental effects of radionuclides and the possible hazards of radiation. In order to determine if there is an environmental effect caused by a development or the storage of the mine waste as tailings, it is necessary to know what the quality of the environment was prior to the production of the tailings. This is the reason for baseline data programs. It is neither technically nor economically feasible to completely define the environment in any given region. Therefore, judgement decisions must be made prior to the development as to what aspects of the environment are more likely to be affected by the presence of the tailings, and to what degree. These decisions have evolved over a period of time based on information generated from existing facilities, scientific research, and the development of strong public opinions. This process is not static, but evolves with the growth of information and the technological developments which permit greater control of industrial pollution and waste management. A well-designed baseline program will reflect changes in science, technology and public attitude, even in advance of government regulations to enforce policy changes. RATIONALE FOR PREOPERATIONAL BASELINE STUDIES There are three distinct phases to be considered in the construction of an environmental assessment for a uranium mine tailings containment facility. I. Pre-operational 2. Operational 3. Post-operational (decommissioning and long term).

Jan 1, 1980

By G. Gaidajis

For the environmentally safe management of the sulphidic waste rock produced from the Stratoni mining facilities, TVX Hellas has examined, within the frame of the funded European research program ROLCOSMOS, the effectiveness of an alternative rehabilitation scheme involving the segregation of the sulphide rich components of waste rock and the engineered application of alkaline additives. The examined waste rock management scheme was evaluated with laboratory test work followed by pilot-scale field applications. In this paper, the results from the two-year monitoring period of the pilot scale application are presented.

Jan 1, 2002

By P. Otto, G. Borm, Th. Dickmann, F. Amberg, R. Giese

Based on many years experience, the well approved Tunnel Seismic Prediction System (TSP) has been substantially enhanced for online geological prognosis in underground construction. The method provides high resolution seismic images of the rock mass ahead of a tunnel face via an array of standard anchor rods. These contain3D-geophones and can routinely be installed during the excavation process. A further outstanding feature of the system is the newly developed high frequency repetitive seismic source which can easily be fixed at a tunnel boring machine. A user-friendly software package enables online data acquisition, processing, 3D-visualization, and interpretation.

Jan 1, 2001

By Jim M. Johnson

INTRODUCTION The facility water balance and maintenance of sufficient excess storage volume in the solution control system are often amongst the most critical facets of design and operation of heap leach facilities in cold, snowy climates. The purpose of this paper is to summarize and discuss climatic and operational factors which have a significant impact on heap leach water balance evaluations in high snowfall climates. Example water balance simulations are presented for proposed heap leach projects located in areas subject to high snowfall in the northwestern United States. The climatic data input to these simulations represent both net evaporative and net precipitation climates. FACTORS AFFECTING WATER BALANCE EVALUATIONS The combined effects of seasonal operations, re1atively low evaporative losses, and potentially large inflows to the solution control system from snowpack and spring storms require an enhanced awareness of the factors which impact facility water balance. Unlike operations at mines in more arid climates, seasonal excess water rather than a shortage of make-up water is the rule. A thorough understanding of a number of factors, including regulatory constraints, climatic influences, and impact of operating philosophy and techniques, is critical to successful development of an incident-free heap leach operation under high snowfall climatic conditions. A process water circuit schematic for a typical heap leach operation is shown on Figure 1.

Jan 1, 1988

By Martha Caudill

When planning for the required amount of storage capacity at a proposed open-pit coal mine or determining when additional storage capacity should be added to an existing operation, the decision must be based on a breakeven- analysis relating the discounted cost of capital for storage capacity to the marginal costs incurred for additional crew call-outs on weekends and demurrage charges. The model developed in this paper uses simulation techniques applied to two coal mines in the Powder-River Basin to determine how much storage capacity is needed and when it is needed. The paper also generalizes the simulation model so that it can be applied to any type of open-pit mining operation. Two important results of the model are (1) all of the mines in the Powder River Basin have excessive amounts of storage capacity, and (2) management is provided with estimates of the total number of hours that crews will have to work on the weekends as annual tonnage increases, and the levels of demurrage costs incurred.

Jan 1, 1986

By Roy E. Williams, Muriel S. Robinette

Ground-water contamination from uranium operations could be avoided with proper site analysis techniques. Contamination from tailings disposal sites is carried downgradient within permeable hydrostratigraphic units which can be located and delineated with a combination of drill holes and borehole geophysical logs. Engineering safeguards can be designed into a disposal site by determining if permeable hydrostratigraphic units exist which will dominate the flow system and provide seepage paths for contaminates. It is more cost-effective to accurately assess seepage potential with geophysical logs before contamination becomes a problem than to be faced with a major clean-up effort after contamination has occurred. If a uranium waste impoundment already has a ground-water contamination problem, then borehole geophysical logs may be used to delineate the contaminated zones which need to be cleaned up. It is less expensive to withdraw the contaminated plume from only within the affected zones rather than attempt to dewater the entire area. Consequently, the geophysical logs are invaluable in assessing which zones should be pumped and where packers should be set for construction of piezometers to allow monitoring of water-quality changes. A surface geophysical technique is being tested and developed for monitoring of seepage frum uranium tailings impoundments. Geoelectric soundings and profiles from direct current resistivity surveys can locate anomalous areas for the placement of drill holes to better assess seepage. This reconnaissance technique should eliminate random hit-or-miss drilling at sites where monitoring is required. Tests and experiments are being done on developing the capabilities of this technique to incorporate complex resistivity data to assess cultural impedence noise which complicates simple resistivity data.

Jan 1, 1980

By G. L. Schroeder, C. H. Summers, D. B. Lindsay

INTRODUCTION The requirement that is placed on owners and operators of underground mines to protect workers against the health hazard of inhaling radioactive materials which are short-lived decay products of 222Rn can be satisfied by applying a considerable variety of what we may call "engineering" solutions as well as a number of "administrative" remedies to the problem. The most obvious of the "engineering" approaches has always been that of forced ventilation, in which relatively clean (i.e., radon-free) air from aboveground is drawn or pushed through the mine workings by a system of fans and ducts. This relatively clean air, in sweeping through the drifts, stopes and haulageways, dilutes the radon and radon-daughter concentrations in the air of the mine, and performs the added beneficial function of removing the daughter-mixture quickly enought to limit grow-in of the longer-lived nuclides in the group that make up the "toxic trio" on which the Working Level (IM) unit is based. Effective as the dilution-ventilation method is for control of WL in most underground mining situations, however, the increasing strictness of control measures that have been imposed on the mining industry over the last two decades have demanded measures of even greater effectiveness. In times of poor markets for yellow-cake and other products of the mines, mine operators are pressed to reduce operating costs, and the installation of additional primary ventilation capacity can be a severe burden on a mine that is already laboring under an unfavorable earning power. When traditional dilution-ventilation systems alone cannot meet the requirement for WI, control, radiation safety engineers and ventilation engineers begin to look at alternatives and auxiliary methods. Since the radon which produces the toxic daughter products originates in the rock of the mine walls, and since, in most United States mines, that rock is a porous sandstone through which air can move under the effect of atmospheric pressure gradients, and through which radon can diffuse relatively freely, one way to help control the growth of WL would be to hinder the escape of radon from that reservior of porous rock. An appealing; method for hindering that natural flux of radon-polluted air from the walls of the mine has long been apparent; namely, to apply a low-permeability coating over the surface of the rock, thus sealing the radon in place and, in theory at least, preventing its escape into the mine air. Our 1970 report to the U.S. Federal Radiation Council on the subject of cost impacts of proposed changes in the occupational standards for exposure of underground uranium miners to airborne radon daughters noted the possibility of using polymeric wall sealants as a means of controlling radon-pollution of mine air. Since that time a number of reports have appeared in the technical literature advocating this technique for restraining the escape of radon from building materials, mill tailings, and other materials containing 226Ra, in addition to the surfaces of underground mine workings. During this period, some controversy has occurred over the question of the probable effectiveness of wall sealants in limiting the escape of radon from the rock. Our 1970 report speculated that flaws (cracks and "pinholes") in the coating might be all but unavoidable in practice, and that even a conservative estimate of the frequency of such flaws would lead to a prediction of ineffectiveness. Hammon et al, in a laboratory evaluation of radon sealants conducted by Lawrence Livermore Laboratory of the University of California in 1975 on behalf of U.S. Bureau of Mines, concluded that a wide variety of polymeric coatings would provide "nearly 100% effectiveness" in restrain¬ing escape of radon from mine wall surfaces if applied in "thicknesses between 5 and 10 mil" (125-250 [y]pm). John Franklin and co-workers at the U.S. Bureau of Mines laboratories in Spokane, Washington, have carried the experiments with polymeric sealants through additional laboratory tests and into actual mine environments, reporting that selected sealants could provide attenuation of radon flux by a factor of four (75-80% reduction). Robert Bates and John Edwards of USBM developed a computer-assisted mathematical/physical model that predicts a relatively small effect of flaws in a low-permeability coating on the radon flux from a sandstone-type matrix. FIELD TEST Since all actual experimental work with wall sealants showed some beneficial effect on radon attenuation (even if not as exciting as the "nearly 100%" predicted by Hammon), USBM was encouraged to extend its evaluation to an actual operating uranium mine, and awarded a contract for that work to Arthur D. Little, Inc. in September 1979. We were fortunate in obtaining the voluntary cooperation of Atlas Minerals Division of Atlas Corp., who operate a mill and several underground mines in and around Moab, Utah. Atlas made available for our use a small T-shaped drift in their Pandora Mine in LaSal, Utah, and provided space for instrumentation and recordkeeping by our field crew in a surface building near the mine entry. Atlas also provided electricity and water to the test site, together with assistance in establishing necessary ventilation, removing rubble from the site, conducting periodic WL surveys and furnishing auxiliary man-power for the heavy hard work of coating the walls with gunite prior to application of the polymeric sealant. The generous coopera-

Jan 1, 1981

By K. Mulukhov

Introduction The depth of open-pit mines continues to increase. The typical designed depth for large quarries is now in the 300 to 600 m (1,000 to 2,000 ft) range, with some quarries even reaching depths of 700 to 800 m (2,300 to 2,600 ft). The cost of handling material from such deep mines can be 60% to 70% of the total cost. The capacity of using a cyclic means of transport, such as trucks or skip hoists, continues to drop as the depth of the mine increases. In addition to the problem of maintaining relatively low slope angles for haul roads, increases in environment pollution and increased fuel consumption make the use of trucks ineffective below 100 to 150 m (300 to 500 ft). In handling large quantities of bulky materials, i.e., more than 15 to 20 Mt/a (16 to 22 million stpy), under fixed terminal points, belt conveyors are preferable in almost in all cases. An exception is when the lump size of the load is too high for the conveyor to handle. Belt conveyors in mines can handle loads having a maximum lump size of no more than 250 to 300 mm (10 to 12 in.). The use of a movable crusher prior to the belt conveyor in an open-pit mine would make the process very expensive. Such an operation could not compete with truck haulage on a cost basis. Belt-car conveyor The main factor that restricts the lump size for conventional belt conveyors is collision impacts between the load-carrying belt and idlers. These impacts are completely eliminated in the belt-car conveyor, which was previously described by Mulukhov (1977). The unique feature of the conveyor is that the load-carrying belt is supported by moving cars. Each car consists of a troughed cross strap that supports the belt with the load and two rollers, on which the cars move along upper and lower rails. Cars are connected by two endless chains revolving over sprockets. The system uses conventional drive pulleys, tail pulleys and return idlers. There are no fasteners between the carrying belt and the car cross straps. The conveyor belt simply drives the cars along by the force of friction between the belt and the straps. A belt-car conveyor was first installed in 1970 at the Karatau mining operation in Kazakhstan. The system was an elevating conveyor with an inclination of 20° and was located on the final part of the truck haul road. The conveyor transferred blasting phosphate ore from the quarry to the crushing plant. A bunker was mounted between the trucks and the conveyor. The maximum lump size was typically 1,200 mm (48 in.). However, there were some lumps as large as 1,500mm (60 in.), which was greater than even the belt width (Fig. 1). Because the prototype belt-car conveyor was intended only for industrial trails, it had a relatively short length of about 50 m (160 ft). Nevertheless, the conveyor yielded a substantial profit over the cost of truck haulage. Naturally, the first question one might ask is why such a system is not now widely used. The reason was the further development of a huge transporting complex for the whole quarry at the base of the belt-car conveyor. After successful industrial trials from 1971 through 1973, which demonstrated the systems engineering feasibility, the State Planning Committee of the USSR included the development of the belt-car conveyor in its plan for the development of the national economy. However, mistakes were made. First, a single organization was not in charge of the development. Second, a plant that was not specialized in belt conveyors was put in charge of the design and manufacturing of the conveyor system. In addition, the crusher was installed before the stacker, which negated the major advantages of the belt-car conveyor. For loading the conveyor, a special blade feeder with pendulum suspension was developed, but it was found unreliable. The main conclusion was that, in-

Jan 1, 2003

By Paul E. Senseny

Data from triaxial compression tests per- formed at confining pressures between 0 and 100 MPa are used to evaluate parameters in a contitutive model that is based on microcracking. The tensorial formulation is reduced to equations that are specific to the simple Loading imposed in triaxial compression tests. Procedures for evaluating the model parameters are described, dependence of parameter values on confining pressure is discussed, and prediction of failure is assessed.

Jan 1, 1986

A general form of pore pressure in a saturated porous medium, which is subject to external forces it has the pressurization of a borehole in hydraulic fracturing, is derived. The pore pressure consists of two parts: the preexisting initial pore pressure and the pressure induced by fluid flow due to the pressure difference caused by pressurizing the borehole The initial pore pressure associates with the effective stresses in the stress, fields of crustal stresses and a thick hollow cylinder subjected to internal and external pressures. The pure pressure weaned by fluid flow is implied in Biot's (1941) constitutive equation for three-dimensional consolidation and proved to be the exact effective pore pressure (Nur and Byerlee, 1971). Using this general form of pore pressure, a general poroelastic model for hydraulic-fracturing breakdown (fracture-initiation) pressure was derived. This model was validated by laboratory experimental results on Indiana limestone. Three sets of tests were conducted under the conditions of preexisting pore pressures of 6, 21 and 28 MPa, respectively. The equal horizontal (confining) pressures applied in the tests ranged from 6 to 60 MPa. Hainison and Fairhurst's (1967) model appear to be applicable to rocks having microcracks or flows. Experimental results of fracture reopening pressures on Indiana limestone supported this assertion. The test results also demonstrated that, when the pressure difference between the confining and the initial pore pressure grows larger than 14 to 21 MPa, the breakdown pressures agree well with the general poroelastic model developed in this paper. This indicates that the developed model will be superior when hydrofacturing stress measurements are conducted at great depth.

Jan 1, 1997

By Duane Jorgensen

Two types of factors direct the search for gypsum deposits-geologic factors and economic factors. The prime control and major economic factor is the market for the possible products. The major geologic factor is the control of the depositional environment, which determines whether or not a deposit exists within a given market area. After discovery of a deposit, the distinction between economic and geologic factors is not as clear as before, and the geologic characteristics of the deposit become the economic factors in evaluating the deposit for production and processing. The geologic characteristics are concerned with 1) the gypsum rock, 2) the host or country rock, and 3) geologic fluids, such as gas and water. Prospecting includes both original discovery and the continuing information-gathering period. During this time the above three characteristics of the deposit are determined, usually from core drilling, from phenomena observed during drilling, and from the core and hole. Important information and means of gathering it include sample analyses, sample volume, impurities and their distribution, water levels, aquifer tests, permeability measurements, rock strength tests, rock and overburden thicknesses, rock types, and direct observations of water and gas. Data gathered or inferred from all of these determine the mining or quarrying geometry and environment: which in turn determine the equipment, development, production, and processing costs. Impurities in gypsum deposits are very important in manufacturing processes. The impurities and their effects vary widely from inert minerals with little effect to reactive minerals with strong and mostly detrimental effects.

Jan 1, 1978