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By Don K. Smith

The northern zone of the Indus Basin comprises a vast area of essentially level land in West Pakistan which is irrigated by means of a complex system of canals that distribute more water to more land than any other irrigation system in the world. In spite of the development of this extensive irrigation system, production of food and fiber has not met the requirements of the country. The major reasons that crop production has not reached anticipated levels are the lack of adequate irrigation supplies and effective subsurface drainage. As a result, yields per acre are among the lowest in the world for areas where irrigation is practiced, and many thousands of acres of valuable crop land have been removed from production or are seriously affected by salinity and waterlogging. To rectify the problems of irrigation supply and drainage the Government of Pakistan has undertaken a program of Salinity Control and Reclamation which is being administered by the Water and Power Development Authority. The Authority retained Tipton and Kalmbach, Inc. , of Denver,, Colorado as Engineers for the program in the northern zone of the Indus Basin. -This program consists of the construction of about 30, 000 irrigation wells to provide drainage for waterlogged lands and full irrigation supplies, including leaching requirements, for; all cultivated lands. The program will ultimately ensure the productivity of 20 million acres and will involve an annual pumpage from ground water storage of nearly 40 million acre-feet of water. The value of the crops produced when the program is in full operation will be 4 to 6 times the value of the crops that could be produced without the benefits of this program.

Jan 1, 1967

By T. Y. Yan

An effective elution technique has been developed for recovering uranium from ion exchange resins loaded in the carbonate leaching process. The procedure employs direct elution with acid. In comparison with conventional base elution, the procedure yields eluate of higher uranium concentration, results in a more complete elution, keeps the performance of the ion exchange resin at the fresh stage, reduces acid consumption, eliminates carbonate consumption and reduces waste for disposal. The technique has been successfully demonstrated in the field and patented. It lends itself to commercialization.

Jan 1, 1988

By Alan A. Campoli

"Pumpable crib support for longwall gates and ventilation bleeder entries is becoming more prevalent in United States coal mines (Figure 1). The pumpable crib selection and spacing design has been greatly refined by the full-scale testing conducted at the National Institute for Occupational Safety and Health (NIOSH) Mine Roof Simulator. The peak load on 24-, 27-, 30-, and 36- inch diameter J-Cribs is, on average, 300, 400, 450, and 650 kips, respectively. Standard J-Cribs have useful deformation of approximately 8 inches. The maximum useful deformation can be extended to 18–20 inches with an exterior reinforcement layer or an interior double wall, for all four J-Crib diameters. J-Crib diameter is primarily limited by a maximum 4.5:1 height-to-diameter aspect ratio. Spacing is controlled by immediate roof stability, anticipated pillar deformation, and roof bolting practice. Approximately 500,000 J-Cribs have been successfully placed in 18 different United States longwall coal mines. The selection and spacing of the J-Crib utilized to date is summarized for each mine.IntroductionLongwall productivity and safety depend on ground stability in gate and ventilation bleeder entries. Numerous support technologies have been developed and employed in United States longwall mines since the 1990s to replace traditional wooden cribbing. Many, if not all, of these technologies have been evaluated at the NIOSH Pittsburgh Research Laboratory since the mid-1990s (Mucho, et al., 1999). The NIOSH Mine Roof Simulator has a 3,000-kip vertical load and 24-inch displacement capacity, which has facilitated in-depth analysis of complex multimaterial support technologies like the pumpable crib. The extreme cost and human exposure risk associated with modern longwall mining prevent the trial and error development that would have been required absent the Mine Roof Simulator research. As a direct result of this research, pumpable cribs are becoming a major component of the standing support industry, providing the peak capacity and extended deformation characteristics necessary for successful gate and bleeder support. Recent research by Zhang, et al. (2012) and others have defined the required performance characteristics."

Jan 1, 2015

By John Rudolf

Metropolitan Washington Airport Authority is currently undertaking the first Phase of Metro?s Extension to Northern Virginia. A wide range of tunnel options was considered for the Tysons Corner segment that involves about 20,000 feet of track and four stations. After consideration of many options that involved EPBM single track tunnels, NATM tunnels and large bore tunnel concepts for the entire Tysons alignment, the design now being progressed into construction involves two 520 meter (1,700 feet) long soft ground NATM tunnels with several hundred feet of cut-and-cover construction. The paper discusses the process that led to selection of the short NATM tunnels as the most feasible of the options considered for Tysons.

Jan 1, 2008

By W. L. Hollinshead, Alwin Poltinger

GENERAL Tunneling by hand digging is as old as the history of humans. Mechanized tunneling did not begin to replace manual methods until around the beginning of the 19th century(Brunel patent, 1818). It brought with it a higher degree of safety for the workers and a much higher rate of boring. But for any kind of tunneling, an accurate and feasible method of survey must be accomplished during mining. That means, that for each tunnel project the most efficient survey method must be devised. The method to be defined depends on the spatial curve, diameter, speed of tunneling method, purpose of tunnel and contractual accuracy. The final accuracy of a finished tunnel is heavily dependent upon three critical elements: 1. Quality of the above ground survey 2. Tunnel traverse and directional transfer from above ground 3. Guidance for the mining job. Combining The Survey And The Guidance System Since the guidance system (GS) is one of the critical factors in the determination of the tunnel final accuracy, it is necessary to analyze how the directional control and tunnel traverse measurements can be effectively integrated into a GS. Any guidance system, whether manual or high technology, must continually detect and compensate for movement of its theodoliteor laser station standpoint. Movement beyond a preset allowable amount must generate a warning or error message to the user. Additionally, optical refraction can interfere with the results. Refraction will have an effect on any kind of optical GS. Therefore, careful selection of target and standpoint locations can minimize, but not completely eliminate, the influence of refraction. Some GS technologies, such as Gyro-based systems, are immune to the effects of refraction, but have other, methodical errors which will be described later in this paper. GUIDANCE Overview Why is a guidance system necessary? A GS, whether manual or automatic, compares the design direction and slope of a tunnel with the current position of the mining

Jan 1, 1997

By Paul D. Merica

One of the most useful, and at the same time least commonly used, methods of thermal analysis for the determination of transformations in metals and alloys consists in the recording of the time intervals required for successive increments of temperature change during heating or cooling, the temperature of the furnace containing the specimen being altered at a uniform rate. The curve obtained by plotting these time intervals as a function of the mean temperature of the specimen durnig the interval is called the inverse-rate curve. It is probably due to the fact that no simple and convenient method has apparently been available for the measurement of the successive time intervals that this method has not been so generally used as, for example, the differential method, for which several types of automatic or semi-automatic apparatus have been designed. Whenever this method has been used, the intervals have usually been measured with the use of a chronograph; its operation as practised at the Bureau of Standards1 is as follows: The temperature of the specimen is measured by a thermocouple and a dial potentiometer. The operator sets the potentiometer at successive values of the electromotive force differing by equal increments, usually 0.02 millivolt, and records the exact instant on the chronograph, by pressing a contact key, at which the galvanometer coil passes through its null position. Two-second intervals are also recorded on the chronograph record, and the number of seconds elapsing between successive signals is afterward counted from the record and plotted as a function of the electromotive force or of the temperature. This is an admirable method and by it most minute thermal arrests may be detected, but it requires a good chronograph, which is generally difficult to obtain, and the time of one operator during the recording and, subsequently, to read the intervals from the record; the latter operation often takes from one to two hours. There is here suggested a simple, convenient, and, it is believed, equally accurate method of recording the successive time intervals, by which the

Jan 1, 1921

By Minoru Kawamura, Yukio Kato, Yoshikazu Hirosaki, Satoru Suzuki

A wireless blasting system for tunnel construction which consists of a loop antenna, oscillator and blaster, has been developed. The basic technology of this method had been studied in the 1970s to develop a wireless initiation system for undersea blasting and has been successfully used as a Remote Controlled (RCB) system in the Honshu/ Shikoku Bridge construction project in Japan.

Jan 1, 1998

By Ian Sherrell, Michael Pirie, Riddhika Jain

"Conventionally, direct reduced iron (DRI) mixed with char from rotary furnace discharge is cooled down before separating the char via magnetic separator. The DRI must then be reheated for further downstream processing (i.e. smelting). Outotec has developed the PyroMagTM specifically to conserve thermal energy during this separation process. DRI is cleaned by feeding hot particles directly after reduction onto a rotating surface just above a stationary magnet assembly. This allows hot feeding of the clean DRI to the downstream equipment which renders the entire process more energy efficient.Continuous testing of the PyroMagTM was performed to verify design and separation. Because the use of DRI/char was not practical for test use, the pilot plant testing was conducted using a magnetite/sand mixture heated up to and beyond the Curie temperature of magnetite. The separation was efficient for both cold and hot processing with recoveries varying from 98.6% to 99.74%. As expected, the recoveries decreased to as low as 83.99% as the temperature was increased beyond the curie temperature of magnetite (585°C). Long term separation was thermally efficient and sustainable.INTRODUCTIONMagnetic drum separators are widely used in the mineral processing industry. These low to high intensity magnetic separators are used to concentrate various magnetic material, such as iron ore, and less magnetic materials, such as ilmenite and garnet from silica and other contaminantsThere are multiple classes of magnetism based on the effect that each material displays while in and out of a magnetic field. In a simplified fashion, which is more appropriate for mineral processing, the magnetic susceptibility of a mineral is used to define it as being ferromagnetic, paramagnetic or nonmagnetic. In this respect, minerals with varying degree of magnetic susceptibility can be selectively separated with different types of magnetic separation equipment in different stages of a process flow sheet. The lower strength magnets are employed early in the process to remove highly magnetic fractions. Later stages utilize greater magnetic fields for capture of less magnetically susceptible material."

Jan 1, 2016

By W. Morgan NeSmith

"The DFI ACIP Pile Committee, in conjunction with the Florida DOT, completed an ACIP Pile installation monitoring and performance test program in late 2016 to advance the inclusion of ACIP piles in future specifications for bridges by state agencies. Piles of different diameters were installed for compression, tension and lateral testing, and one pile was extracted for visual inspection. This paper presents the pile installation, non-destructive testing and load test results of the program.INTRODUCTION The Deep Foundations Institute (DFI) Augered Cast-In-Place (ACIP) Pile Committee performed a ACIP pile installation and test program in the fall of 2016 to demonstrate a fully monitored installation of 18-in and 24-in diameter ACIP piles. The program included automated and manual installation monitoring, Thermal Integrity Profiling, compression, tension and lateral load testing (including monitoring of strain gages embedded along the compression pile shaft), and post-testing extraction of an installed pile for visual inspection.PROJECT DETAILS AND FIELD EXPLORATION A general program location map along with an aerial view of the property in Okahumpka FL is shown in Figure 1. Two Cone Penetration Tests (CPT) were performed at initial reaction pile locations R-6 and R-8 (Figures 2 and 3) The results from CPT R-6 indicated very loose soil conditions at depths below 55-ft, indicating potential relic sinkhole conditions. The test layout was subsequently rotated and moved south of CPT R-8.TEST PILE INSTALLATION DETAILS Seven piles were installed to either 40-ft or 60-ft below grade. Thermal Integrity Profile (TIP) wires were installed in all piles but varied to include one wire each attached to a steel center bar or one wire attached to the steel center bar AND four additional wires (each) attached the steel reinforcing cage installed in these piles. Additionally, the 24-in lateral and compression piles had four steel or PVC casings installed next to the TIP wires for testing by thermal probe and/or wire inserted into the casing."

Jan 1, 2017

By Garg R. K

Himalayan ranges on the northern border of India are source of many perennial rivers and suitable sites for large dams have been located therein. However, Himalayas are young mountains and comprise of comparatively weaker rocks having adverse geological features like shear seams, thrust zones, minor and major faults and intense jointing. Tehri dam project under construction, is located in lower Himalayas. It envisages construction of a 260.5 m high earth and rock fill dam and an underground power station of 2000 MW (Eight machines of 250 MW each, four of which will be reversible) installed capacity. This dam will be the highest and the underground complex will be the largest to be constructed in India in the Himalayas. The under- ground complex comprises of 25 km of tunnels varying in diameter from 6 m to 11 m and seven large cavities of various sizes, the biggest being 22 m in width, 194 m in length and 65 m in height. Rock formation in the power house complex belong to phyllites having various joint sets spaced from 0.5 m to 3 m. Due to the poor geological conditions and closely spaced net work of large cavities, the planning, analysis and design of the cavities is a challenging job. Many aspects have to be considered in determining the sizes, locations and orientations of the various cavities taking into account the properties and jointing pattern of rock mass, functional requirements and economics. Correct evaluation of loads and forces is necessary for economical design of supporting system for the various cavities. Rock loads on the roof of cavities and their supporting system will be assessed with the help of the latest methods. The stability of the rock mass around the cavity will be evaluated by elasto- plastic method using finite or boundary element techniques. A scheme of field investigations, laboratory tests and instrumentation for deter- mining rock properties in preconstruction stage and for monitoring behaviour of the underground cavities during and after construction is envisaged.

Jan 1, 1987

By Robert Marks, Enrico Dal Negro, Richard Schulkins

"In July 2012 construction started on the Grosvenor project, a five million tonne per annum underground longwall mine planned for Moranbah in Central Queensland. Grosvenor is located immediately to the south of the existing Moranbah North mine which has been in operation since 1998. EPBM technology is being used for the first time in Australia for excavation of a coal drift. The TBM was used to build two drifts on the project, one for the coal conveyor which will transport to surface, and another for people and equipment to access the underground once the mine’s operational (as shown in figure 1). The TBM will pass beneath a steel archway roof that has been installed at the drift’s entrance and begin drilling into the ground to build the seven metre diameter tunnel, descending at an angle of one in eight until it reaches the depth of the coal seam approximately 160 metres below.INTRODUCTION AND BACKGROUND The first TBM used in a coal mine was in 1962, a Wohlmyer, in Westerholt, Germany. A further 75 km of tunnels were excavated in coalmines within Germany between 1971 to 1983. This paper will review the Grosvenor Mine project (Grosvenor) owned by Anglo American Metallurgical Coal (Met Coal) includes the construction of a new underground coal mine and surface facilities on the northern boundary of the town of Moranbah in the Central Queensland Bowen Basin coal precinct. The contract to undertake the underground development works was awarded to Redpath in 2012. The scope of the Grosvenor Project was to construct temporary office and workshop facilities, install 2 portal arches, backfill boxcuts and infrastructure, excavate by EPBM method 967 m of 1:6 gradient Conveyor Drift and 1,258 m of 1:8 gradient Transport Drift, development of 2,520 m of pit bottom area roadways and development of 10,550 m of the first longwall tailgate. In the following figures 2 and 3, is possible to see the aerial view of the job-site."

Jan 1, 2016

By N. De Nevers

A calculation method has been developed for carbonated water flooding. This method takes into account the effects of oil viscosity reduction and oil swelling, due to carbon dioxide transferred to the reservoir oil from the carbonated flood water. It shows the effect of changes in carbonation pressure and carbonated water slug size. The method is based on a Buckley-Leverett-type linear flow model. The mathematical approach is similar to that developed by Welge, et al, for enriched gas drives. However, it contains a very important advance over the previous papers in its solution to the problem of injection of a "slug" of carbonated water rather than continuous injection of fluid of constant composition. The method also shows the chromatographic nature of the transport of Co 2 in a carbonated water flood. Sample calculations indicate that viscosity reduction is the most important source of incremental oil recovery. Swelling contributes to a lesser extent. Raising the carbonation pressure increases recovery, but this effect diminishes as pressure is increased. increasing the slug size also increases recovery, but again this effect diminishes as slug size is increased. INTRODUCTION In recent years there has been considerable interest in carbonated water flooding as a secondary recovery method. Several carbonated waterflood field tests are now in progress.1'3 Yet there is no published method for predicting the results of a carbonated water flood. This paper presents a method for predicting the oil recovery for ; carbonated water flood. In a carbonated water flood. a carbonated water slug is injected into the reservoir. Then this slug is driven forward by injecting plain water behind it. The field operator must choose the carbonated water slug size and the CO2:water ratio in the slug. The calculation method presented here shows the effects of both slug size and CO2:water ratio on oil recovery. The calculation assumes that the addition of carbon dioxide to flood water increases the oil recovery by two mechanisms: 1. Carbon dioxide transfers from the flood water to the oil, thereby reducing the viscosity of the oil. At the lower viscosity, the oil flows more readily and is more easily swept out by flood water. 2. The carbon dioxide which transfers to the oil phase causes the oil phase to swell. Assume that the residual oil saturation (volume per cent of pore space) is the same whether the oil is carbonated or not. Then the same volume will contain less stock tank oil if it is carbonated, due to oil swelling. The plain water driving the carbonated water slug forward will strip the CO2 out of this residual oil., causing this oil to shrink. This swelling-followed-by-shrinking leaves behind a lower residual oil saturation than would have been left behind by a plain water flood. Since some oils swell to almost twice their original volume at reasonable CO2 pressures, this effect may be large. Two additional benefits sometimes claimed for carbonated water flooding are: 1. Adding CO2 to flood water greatly increases the water injectivity in some cases.1 (Injectivity = BPD of water injected/ft of sand x differential bottom-hole pressure, psi.) CO2 forms a weak acid in water; injecting carbonated water is, therefore, somewhat like acid treating. If a formation responds to acid treating, it will probably have a higher injectivity for carbonated water than for plain water. Although the increased injectivity increases the profitability of a carbonated water flood over a plain water flood, it does not change the production curve based on total fluid injected. Therefore, the method presented here does not show any effect of increased injectivity. 2. It is claimed in one patent4 that carbon dioxide will react with some constituents of the oil to form detergents. These detergents are expected to increase the oil recovery. There is little data available to support this idea. The calculation method presented here takes no credit for extra oil recovery due to the formation of any such detergents.

Jan 1, 1965

By Robert W. Bartlett

Pretreatment of sulfidic refractory gold ore by biooxidation in coarse-crushed ore heaps is a promising new technology, especially for low-grade ores that cannot afford pressure leaching or roasting. An algebraic equation for determining the economically optimum biooxidation of the sulfide minerals in the ore is provided that is based on the empirical biooxidation rate constant for the ore, the project's financial discount rate, the ore grades of both refractory and non-refractory gold, and the gold price. Then, the economically optimum duration of bioheap pretreatment, t$, is determined from its logarithmic relation to the optimum fractional extent of ore biooxidation, F$, through the ore's biooxidation rate constant. In doing so, it is assumed that the ore is adequately crushed and amenable to bioheap pretreatment and that the cumulative oxidation of the ore with time is known from ore testing so that the rate constant can be determined.

Jan 1, 1995

By U. Paschedag

Today's modern longwalls have to produce high outputs with maximum availability and running time with the additional requirement of ever increasing panel width and length high-performance face conveyors require intelligent solutions for the integration of drive systems and chain tensioning devices into the overall longwall automation. The fully automated chain tensioning system allows the operation of an AFC with optimum chain tension at all times regardless of actual load on the conveyor. For the heavy-load start-up, the accurate load-sharing and an efficient overload protection of modern AFC's WESTFALIA and Dodge jointly developed the integrated WB/CST drive system.

Jan 1, 1995

By H. S. Caldwell

The objective of this investigation by the Bureau of Mines was to develop and test for laboratory use a retort embodying many features required for practical commercial-scale applications of vacuum distillation of metals. A gas-fired vacuum retort was developed comprising a hollow, cylindrical graphite liner encased in a thin shell of stainless steel. The liner prevents the shell from collapsing under vacuum and protects it from corrosion by metal vapors and the molten metal charge. The vacuum-tight shell protects the liner from oxidation and flame erosion. The retort was used to distill zinc from dross and die-casting scrap. More than 99 percent of the zinc was recovered as a condensate containing 99.8 to 99.9 percent zinc. (Atmospheric distillation recovers 95 to 98 percent at a purity of 99.5 percent zinc.) Distillation rates of 50 pounds of zinc per hour were attained at a temperature of 65° C. with pressures ranging between 50 and 100 microns. Distillation at this low temperature is not practicable at atmospheric pressure. Also, the retort is suitable for distilling other volatile metals such as cadmium and magnesium.

Jan 1, 1961

By P. Valicek, M. Farren, G. Harrison, G. van den Berg, P. van Dorssen, P. Kramers

Full mechanization of the stoping operation using rockbreaking methods other than blasting has been a target of research for decades because of the considerable payoffs in terms of increased face advance rates, productivity and safety. However, the unavailability of economically viable rock breaking tools has prevented progress so far. This has changed with the invention of activated cutting tools, which achieve significantly reduced operating forces, power consumption and wear compared to non-activated tools and thereby permitted the development and evaluation of the first mining system using this technology. Targets for the mining system were a face advance of 100 metres per month on a double shift basis for a 30 metre long panel and eventually fully automated man-free stoping. The purpose of this paper is to describe the design, development and evaluation of the mining system. This was a major collaborative effort between Anglo Platinum’s Rustenburg Platinum Mines Rustenburg Section — Townlands Business Area, Anglo Platinum’s Research and Development Department within the Mining Technology Division, the inventor of the technology Bechem GmbH, the equipment manufacturers Long-Airdox Ltd and Zollern GmbH, and CSIR Miningtek who managed the project. A structured process was followed during the development to minimize the risks involved in using such leading edge technology. The system was rigorously evaluated to ascertain the economic viability of the method and to collect data for the design of a production system.

Sep 1, 2001

By A. Shakoor, A. M. Hirt

The compressive strength of coal, and its variation within and between seams, was determined for four Pennsylvania coal seams (Pittsburgh, Sewickley, Upper Kittanning and Lower Kittanning). Large coal blocks were obtained from 12 coal mines (three mines per seam) and cut into 50-mm (2-in.) cubes using a low-speed saw. The cubes were tested for compressive strength, with the results used to compute actual pillar strength. Using the tributary area approach, the pillar stress was determined. The pillar stress and strength data were used to calculate the factor of safety against pillar crushing for each mine. The cube strength was also correlated with various physical and chemical coal properties (density, percent absorption, percent ash, percent sulfur and percent carbon) to see if any strong relationships exist. The coal's mean compressive strength varies from 9.1 MPa (1326 psi) in the Lower Kittanning seam to 37.8 MPa (5480 psi) in the Pittsburgh seam. Between different seams the pillar strength varies from 29.5 MPa to 48.1 MPa (4282 psi - 6981 psi). Although coal strength appears to be weakly related to carbon content, no relationship was found between compressive strength and other physical or chemical properties. Since the values for the safety factor against pillar crushing generally exceed the recommended value of two, the existing pillars are mostly overdesigned. This implies that an accurate assessment of coal compressive strength could result in the use of smaller pillars and, thus, increased coal production.

Jan 1, 1993

The great philosopher, Marshall McLuhan, is said to have stated that the use of forecasting as a means of guiding future actions is akin to driving an automobile through observations made in the rear-view mirror! In effect, in forecasting it is assumed that there will be an orderly progression from the past into the future, and that the model will be sufficiently unvarying for use to gauge both the directions and the rates of change. The concepts of forecasting have served man well in disciplines in which the relationship between causes and effects can be clearly demonstrated; such as between rainfall and crop yield, for example. These are of no value, however, when conditions change rapidly and unpredictably, or when there is no model at all. Even worse, under such circumstances our preconceived notions may cause us to make some horrendous errors in judgement. Such are the risks at present in attempting to make an assessment of the educational needs of the minerals industry for the coming century! The model, and the conceptions that we have had of the linkages between cause and effect are in disarray. Minerals markets are chaotic, and there is almost no ability to forecast trends in metal values. Linkages between supply and demand seem tenuous. Internationally, producers are struggling for survival. Mine closures, temporary shutdowns, and reopenings are frequent occurrences. Decision-making under uncertainty is probably one of the most certain things about the world mining industry Today!

Jan 1, 1986

A comprehensive understanding of the rocks associated with coal seams, and also of the coal itself, is essential in assessing the likely behaviour of the materials before, during and after the mining process. The nature of the rock materials involved in coal mining can give rise to risks from adverse geomechanical properties, create problems in coal quality control and produce difficulties with environmental impacts, all of which may affect the profitability, impact and safety of the operation. Examples of issues associated with rock composition include slaking and degradation leading to roof, floor, pit wall or spoil pile instability, ignition of methane in mine atmospheres by frictional processes, liberation of quartz-rich dusts into ventilation systems, production of coal with unfavourable preparation or marketing characteristics, and escape of suspended solids or potentially toxic leachates from spoil piles and other types of emplacements. Methods for evaluating rock composition include microscopic examination, chemical analysis and X-ray diffraction, with the latter being especially useful for fine-grained rocks or for clay-bearing materials. Other techniques, such as geomechanical testing and down-hole geophysical logging, are also useful, but give less direct information about the actual composition of the materials involved. Indeed, understanding of rock composition, through microscope or quantitative X-ray diffraction studies, provides a basis for checking information from other sources, explaining different aspects of rock behaviour, and an essential link in relating geophysical or geotechnical data to broader-based geological models as part of the risk assessment process.

Jan 1, 2003

By C. Douglas Lee

This case study describes the work done at the University of Arizona on the development of a computer mine model for an irregularly shaped and spotty ore deposit. The purpose of the work was to develop techniques for modeling any highly irregular deposit, for which the regular 3D fixed block model is not well suited. The deposit contained complex geology consisting of sedimentary beds that had been faulted, thrusted, and intruded by igneous rocks (see Fig. 1). The physical dimensions of the modeled ore body were 1067 m (3500 ft) along the N-S direction and 914 m (3000 ft) along the E-W direction. Topographic relief in the modeled area was approximately 274 m (900 ft). The highest grade of ore was found near the surface with mineralization extending down to approximately 152 m (500 ft) below. Using a fixed bench height of 12 m (40 ft), the model required 25 levels with the highest elevation being at 1286 m (4220 ft). Fig. 2 shows a cross section of the developed computer model of the ore deposit. The particular modeling approach selected for this deposit was the 3D irregular block model.* This meant that only the dimension of bench height was fixed at 12 m (40 ft). Irregular size blocks in the form of polygons were adopted in the model to describe the ore zones and geologic contacts. The following sections of this chapter describe some of the techniques used in modeling.

Jan 1, 1979