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By D Beahm

Introduction The Energy Metals Corporation (EMC) holds numerous mineral properties in the northern portion of the Great Divide Basin including the JAB Property, one of several potential properties being planned for development as a satellite facility to the greater Antelope Uranium Project. A National Instrument 43-101 Technical Report has been completed for the JAB property. This report estimates total measured and indicated mineral resources on the property totaling 3,555,022 pounds U3O8 contained in 2,400,875 tons at an average grade of 0.073 % U3O8 at a .25 GT cutoff in accordance with CIM Standards on Mineral Resources and Reserves. The JAB Property is located in Sections 13, 14, 15, 16, 21, 22, and 23, Township 26 North, Range 94 West, approximately Latitude 42o 14’ North and Longitude 108o 00’ West (Figure 1, Location Map). This property consists of unpatented mining lode claims and Wyoming State mineral leases comprising approximately 2,100 acres.

Jan 1, 2007

By V. I. Kefeli

Public-private partnership efforts are working to create green landscapes from abandoned industrial sites. Fabricated soils, potentially an essential element of such landscapes, are designed to substitute for topsoil on the areas free from ground cover and plant community. The aluminosilicate matrix and humus form the soil micelle - the primary unit of any soil complex. Humus formation is a process based on carbon-nitrogen interaction. Sources of carbon can be materials rich in easily composted products like cellulose and polyphenols (plant leaves) or aged lignin polymers (saw dust from different woody plants). The pathways of carbon-nitrogen complexes are discussed. Properties of the fabricated soil components were examined by bio-tests and plant models.

Jan 1, 2002

By S. J. Taylor

The siting, grading and fill usage for leach pads are important aspects of leach pad design. A method is presented that optimizes the cut and fill quantities (i.e. minimizes the cost) required to grade a leach pad with parallel cells; given the desired pad characteristics of maximum and minimum pad slopes, maximum cut and fill depths, etc. and the cut and fill costs. The method allows the variation of the cut and fill costs with depth, which allows the accurate modeling of the cost savings practice of using cheaper fill towards the bottom of a thick fill. This optimization could be incorporated into a CAD package, enabling a designer to simply draw a desired pad shape on a site plan, input the pad and heap characteristics (required heap volume, heap height, heap sideslopes, etc.) and then have the computer report both the adequacy of the pad area and the optimum grading for the particular pad shape. Using this, a designer could quickly and accurately assess the optimum, or cheapest, grading of many shapes and positions for a leach pad. This would be especially useful for large pads that need to be sited on rough terrain.

Jan 1, 1996

By A. Bruynesteyn

The results of a 2 phase evaluation program designed to determine the amenability of an oxide ore from the Butte, Montana mine of the Anaconda Company to a sulfuric acid leach is presented. In the phase 1 preliminary evaluation, using roll tests and percolation leaching in small columns, it is shown that the ore is highly extractable at low acid consumption ratios. In the phase 2 confirmatory tests in large columns on l-l.5 tons of material, the amenability of the ore to the leaching process is proven and pregnant solution grade profiles from planned commercial heaps established.

Jan 1, 1978

New Mexico?s active coal mines have reclamation challenges that are being successfully resolved through the application of geomorphic grading methods. Soils and overburden material are often of poor quality, and annual precipitation is low. Revegetation on south and west aspect slopes often failed to persist. The New Mexico Mining and Minerals Division had therefore been encouraging topographic diversity on reclamation to improve water harvesting and reduce the percentage of south aspect slopes. Steeply dipping, multiple coal seams may be mined by open pit methods. Structurally-controlled trellis drainage patterns that exist premine have to be replaced with dendritic drainage patterns on reclamation where the overburden had been pulverized. Sodic spoil material is prone to differential settling and the creation of piping features in areas where runoff ponds, which may lead to downslope instability. In each of these scenarios, the approximation of natural drainage patterns on the reclamation has reduced erosion and sedimentation by creating shorter slopes with correct profiles, and by improving conditions for revegetation establishment.

Jan 1, 2010

By J. Herbst, H. B. Treweek, V. R. Degner, G. Poling

Current mineral industry trends toward handling the high tonnages of lower grade ore and tailings retreat lead to a continuing need for large volume flotation cells. The benefits which accompany the larger cell flotation circuit include reduced capital cost, less mill floor space, reduced operational complexity, and less maintenance, provided the metallurgical performance and solids suspension capability of the cell are not compromised. This study reviews development considerations pertinent to large mechanical air induction type flotation cells from the standpoint of cell internal hydrodynamic characteristics and metallurgical performance. While the importance of process related factors such as chemical usage, pH control, solids size and loading, and circuit arrangement cannot be denied, this study will emphasize the relationship of cell size (volume), mechanism geometry (impeller diameter), and mechanism operating conditions (rotor speed and submergence), to air transfer, liquid circulation, liquid "rise velocity" and mechanism power draw. Specific hydraulic performance maps for large capacity 300 and 500 cu ft flotation cells will be reviewed in terms of the relationship between the operational variables and the air and liquid flow conditions within the cell. Metallurgical results, obtained under plant operating conditions for 300 and 500 cu ft cells, are included to illustrate the metallurgical performance similarity of flotation cells which follow hydrodynamic scale-up criteria set forth in this investigation.

Jan 1, 1976

By W. D. Monaghan

Flame cutting and welding is one of the major causes of fires and explosions in underground U.S. coal mines. On May 20, 2006, a flame cutting operation at the Darby Mine No. 1 led to an explosion that resulted in five fatalities. On January 22, 2003, another explosion caused by flame cutting at the McElroy Mine resulted in three fatalities and three serious injuries. On June 22, 1999, a flame cutting and welding operation at the Loveridge Mine caused a fire that resulted in sealing of the mine for an extended period. Finally, on March 19, 1992, an explosion at the Blacksville No. 1 Mine that was caused by welding resulted in four fatalities. The National Institute for Occupational Safety and Health (NIOSH) conducted a study on fires and explosions in underground U.S. coal mines that were caused by flame cutting and welding operations to determine the root causes of these types of fires and explosions. The methodology included interviewing mining personnel who perform flame cutting or welding operations in underground U.S. coal mines and visiting mines to observe these operations, and MSHA reports of investigations and accidents statistics were analyzed. The results were used to identify and compare differences between flame cutting and welding practices and techniques in small and large mines, Eastern and Western mines, low seam and high seam mines, room and pillar and longwall mines, and between experienced miners and new miners. As a result of this study, best practices and interventions were developed to reduce the number of fires and explosions caused by flame cutting and welding. This paper provides a summary of this work. Disclaimer: The finding and conclusion in this report are those of the authors and do not necessarily represent the views of the National Institute for Occupation Safety and Health.

Jan 1, 2009

By A. S. Winters

INTRODUCTION Homestake Mining Co. owns and operates the Bull¬dog Mountain mine near Creede, CO. Lead-silver con¬centrates are produced from ores extracted from narrow veins which dip from 1.04 to 1.39 rad (60 to 80°). An¬nual mill throughput is approximately 90 700 t (100,000 st). Ore shoots average 2 m (7 ft) in width and the ore is often weak, muddy, and poorly consolidated. Walls can be either strong or highly fractured along the strike of the veins. Cut-and-fill methods were initially selected to mine the deposit and considerable overhand mining occurred during the initial stages of the mine. Because of un¬stable ground conditions, mine management recognized that other methods would have to be employed. Undercut-and-fill mining was adapted to extract the loose ore from narrow veins and is presently the primary method used. Approximately 80% of the annual pro¬duction is extracted by this system. Loading ramps are positioned along the strike of the vein on 90-m (300-ft) intervals to allow for 45-m (150-ft) long cuts in each direction from the extraction raise. This interval was selected for efficient slusher operation and stope cycling. Loading ramps rather than chutes were selected for ore movement into 4.5-t (5-st) Granby-type cars because of extremely sticky ore conditions. Mined-out areas are filled with deslimed mill sands. Sand is pumped from the mill into two mine storage dams where the water is decanted. Each dam has a 544-t (600-st) capacity. Sand is pumped into the mine at a 30 to 40% slurry through 2195 m (7200 ft) of 51-mm (2-in.) thick wall pipe. When called for, sand is reslurried within the dams to a density of 55 to 65% and pumped to the particular mine opening requiring fill. Dry cement is added to the slurry prior to pumping. The capacity of the mine fill system is 20 t/h (22 stph). DEVELOPMENT Vertical development below the main adit level oc¬curs on 61-m (200-ft) intervals from an internal winze. Main haulage drifts or laterals are 2.7 x 2.7 m (9 x 9 ft) in section and driven parallel to the vein structures. Ground conditions generally dictate that better progress can be realized in the footwall than in the hanging wall. A distance of approximately 21 m (70 ft) is maintained between vein and drift openings to accommodate a smooth 18-m (60-ft) radius curve for crosscuts which are driven on 91-m (300-ft) intervals. Approximately 18 m (60 ft) of tail room is driven past the vein inter¬section to accommodate train loading. Ore and waste are hauled using diesel or electric driven locomotives operating over 18-kg (40-Ib) rail [762-mm (30-in.) gage]. Loading ramps are established above each crosscut¬ vein intersection. Scram drifts are then driven in the vein to determine the width and grade of the ore. If a prior decision to mine the block had been reached, then the scram is driven only far enough [about 12 m (40 ft)] to establish a drawpoint and scraper tail room. In most cases, the drawpoint is a 2.1 x 2.1-m (7 x 7-ft) cutout driven about 2.4 m (8 ft) into the footwall. On the level above, a raise bore station is established in the footwall to one side of the crosscut-vein inter¬section. The station is of sufficient size to accommo¬date a Robbins 41R raise boring machine. Support for the station is accomplished with 1.2 and 1.8-m (4 and 6-ft) roof bolts and wire screening. Once completed, a 152-mm (6-in.) concrete slab is poured and the raise drill set in position to center the 1.5-m (5-ft) diam hole approximately 1.8 m (6 ft) from the vein footwall. A 229-mm (9-in.) pilot hole is then drilled to the scram cutout below and reamed back to full diameter (see Fig. 1). STOPING Initial mining in a typical undercut-and-fill stope be¬gins from the top of the borehole. The initial cut can be either immediately beneath the track level or the old scram drift depending upon the particular situation. Cuts vary but are generally 2.4 to 3 m (8 to 10 ft) high. Support is accomplished by square-set timber, overhead caps, or roof bolts and screening depending upon ground conditions. Once the initial cut is removed, underhand timber and screen is installed and cemented sand fill placed. The stope is then ready for undercut mining. Prepara¬tion work begins with sinking 4.5 to 6 m (15 to 20 ft) on the borehole plus cutting a slot to the vein. As sinking progresses, hanging sets are installed to provide man¬way and service facilities. Hanging sets 1.7 m (5 ft 8 in.) long and 1.6 x 2.2 m (5 ft 4 in. x 7 ft 4 in.) in section are installed with steel hangers similar to shaft sets. Framed 203 x 203-mm (8 x 8-in.) timbers are made up in advance and shipped to the stope during sinking operations. When manway and service facilities are completed to the floor of cut to be mined, generally 3.65 to 4.3 m (12 to 14 ft) below the previous sand fill, the ore is removed to the hanging wall plus a round each way along the vein. Levelers or long stringers are then installed over the raise opening to support the work deck, slusher, and grizzly opening. All waste and ore from the sinking process is fed by gravity to the raise (see Fig. 2). Once the stope is equipped, mining pro¬gresses outward in each direction from the grizzly open¬ing. Stope equipment consists of 907 kg (2000 lb) pull air tuggers, timber skips, 11-kW (15-hp) ventilation fans, 15 to 22-kW (20 to 30-hp) electric slushers, 762 to 1270-mm (30 to 50-in.) scrapers, 67-mm (25/a-in. ) jackleg drills, as well as other drilling and blasting supplies. In narrow stopes [generally less than 2 m (7 ft) wide], no support is required for the underhand caps of the previous fill. Where wide sections in the vein occur, segment braces are installed beneath the caps to insure that the sand support cannot slip away from the hanging wall. Sections as great as 6 m (20 ft) wide and 4.5 m

Jan 1, 1982

By D. A. Dahlstrom

Since the initiation of continuous vacuum filtration over a half a century ago, control of the performance of these units has been largely in the hands of the plant operators, Obviously there has been the introduction of variable speed drives, adjustable speed feed slurry agitation, vacuum control, etc. However, control of , filtration rate, cake moisture content, dissolved solids recovery by cake washing and dryness of solids where thermal drying is practiced is primarily up to the operators judgment and the settings he imposes on the controls he regulates, - filter cycle time, submergence, vacuum level, filter medium condition, wash water- rate, and other such factors. While this may seem to indicate that good operators are a paramount necessity, and certainly always desirable, a considerably more vital factor is a properly designed filter and filter station for the particular application. With improper design, highly skilled operator control will still leave much to be desired in the way of process performance. On the other hand, with correct design and application, the operator can easily and quickly control normal plant surges and quality variations within the flow sheet. Due to the growing complexity and inter dependency between the various processing steps in the modern flow sheet as well as the rising costs of capital investment, labor, maintenance, and losses involved with off quality product or insufficient recovery, automatic or inherent performance control of many of the unit operations within the metallurgical mill must be achieved. Continuous filtration should and can be susceptible to such control. However, as every instrument engineer knows, before control and automation can be achieved, a basic understanding of the operation is mandatory. This is primarily due to three factors. First, as in so many control applications, it may not be possible or feasible to measure the factor that it is desired to control. Thus, secondary variables that influence the major factor are substituted for control purposes. Secondly, it is always desirable to take advantage of any ?self-regulation? within the operation being performed on the equipment. In so doing, not only will automatic control be facilitated, but instrumentation is minimized while experiencing most efficient operation. Finally, automatic control is not a cure-all for improper process and equipment design and application. Hanging instruments on such an operation can be compared to constructing a sizable building on sand, Results will be disappointing.

Jan 1, 1960

The Kinross DeLamar Mine has been operated for more than 17 years producing precious metals dore' from ores, principally silver bearing with associated gold co-products. In late 1995 a new orebody was commissioned, the Stone Cabin orebody, and the site began a phased shift from silver dominance to gold dominance in terms of precious metals revenue. The Knelson Concentrator was selected to remove gold from a portion of the cyclone underflow stream in the grinding circuit. This paper describes the development of the flow sheet, specific mineralogical impacts on Knelson product and downstream materials, and impacts on mine/mill grade reconciliation.

Jan 1, 1997

By Kotaro Ohga, Gota DEGUCHI, Kiyoshi HGUCHI

The situation of Japanes coal mine is in the severe enviroment, because the policy on coal mining has changed. Every year one or two coal mines will be foced to close. To continue the production of coal, we have to reduce the cost of production. Therefore, the mining by longwall method which is more than 250m was attempted at Taiheiyou Coal Mine to reduce the cost and this mining method have had successful results. To get successful results by this mining method, the controlling and prediction of gas emission from the longwall panel will be most important things. In this paper, the comparison of the predicted gas emission from the mining panel based on the measured gas content by using new device with the observed results is described.

Jan 1, 1993

By Carlos A. Jaramillo, Seng Hing Ngu

"The Murum Hydroelectric Project is a 944-MW development consisting of a 145 meter high RCC dam, two 2660 meter long, 9 meter excavated diameter upper pressure tunnels, surge shafts, two 950 meter long 7 even meter OD penstock tunnels and a surface powerhouse. The project is located in the eastern section of the State of Sarawak, Malaysia, about 200 kilometer from Bintulu, in the Rajang Fold-Thrust Belt, an Upper Cretaceous to Upper Eocene fold and thrust belt that extends under most of Sarawak. The Belaga Formation consists of a thick sequence of deep marine sediments currently appearing as steeply dipping strata of thin to thick bedded, fine to medium grained sandstone interbedded with argillaceous rocks, resulting on a “Flysch” type geologic environment. The project pressure tunnels were excavated perpendicularly to these strata resulting on frequent change on rock conditions and behavior, and required rock support. This paper describes the characterization of the rock mass, design of rock support, and incidents during construction. Particularly interesting were convergence issues in Tunnel 1 resulting from Tunnel 2 excavation. INTRODUCTION The Murum Hydroelectric Project is located in the eastern section of the State of Sarawak, Malaysia, about 200 kilometers from Bintulu, and upstream of the Bakun Hydroelectric Project. The project consists of a 145 meter high RCC dam, a 50 meter high intake tower, two nine meter OD and 2660 meter long upper pressure tunnels, two nine meter OD and 175 meter long vertical shafts, two 25 meter diameter and 80 meter deep surge tanks, two seven meter OD and 950 meter long penstock tunnels and four seven meter OD and 50 meter long branch tunnels, two 8 meter diameter to 3.4 meter diameter and 1400 meter long penstocks installed within the shafts, penstock tunnels and branch tunnels ending in a powerhouse with 944 MW installed capacity. LAYOUT OF THE WATERWAY TUNNELS The two waterways tunnels are basically parallel as shown in Figure 1, with each tunnel supplying water to two units in the powerhouse. The tunnels start separated 25 meters center to center, but 250 meter from the intake the separation increases to 35 meters. The tunnels are lined with reinforced concrete until they reach the surge shafts and steel lined from there to the powerhouse. The tunnel profile is shown in Figure 2."

Jan 1, 2016

By G. Robert Ganis

The folded Appalachians are a physiographic province of eastern North America extending from Nova Scotia to Alabama. The province strikes through the state of Pennsylvania (see Fig. 1) comprising roughly 20 percent of the state and it is here that the focus of this paper will attend. The province is a thick sedimentary sequence which was compressed into a folded mountain belt by the Appalachian Orogeny which terminated at the end of the Permian. The stratigraphy contains both carbonate and non-carbonate rocks as a reflection of the depositional environment at any specific time and place during the Paleozoic geosynclinal accumulation. Although a great deal of variety and difference exists in the geology from one end to another, a common thread of structural and stratigraphic style exists. Carbonate rocks, limestones and dolomites, comprise a substantial percentage of the total Appalachian stratigraphy and are interbedded with clastic materials such as sandstones, siltstones, and shales. The carbonate rocks are extensively mined and collectively represent a vast economic resource.

Jan 1, 1983

By Peter J. Alderson, M. P. Evans

The application of conveyor belting in bulk materials handling is crucial to the success of most mining enterprise selection of the cover rubber grade and thickness has a direct bearing on the cost per ton conveyed. Current and future demands placed on belt covers by higher speeds, increase capacities and greater lifts means today's cover will not be adequate for tomorrow's applications. The simple selections of RMA Grade I and I1 and MSHA above and below ground need to be expanded upon. This paper describes various new approaches for selecting conveyor belt covers for specific applications. It also presents examples of material and compounding techniques available to the belt manufacturer to improve the performance of their covers.

Jan 1, 1996

By William A. Dupree, Mark J. Schultz, William J. Francart

Maintaining intake escapeways clear of contaminants from mine fires has been an issue for many years. Removing fire sources from these entries reduces the risk of contamination greatly. In recent years, concerns have been raised related to contamination of the intake escapeways by leakage when fires occur in adjacent airways. This paper discusses measured leakage rates for multiple stoppings at various pressure drops across stopping lines in a number of actual coal mines. Also addressed are methods for reducing the contamination risks for the intake escapeways, considering the ventilation system employed. The Effects of Air Leakage The effects of air leakage through ventilation controls has been well documented. While normal leakage is expected through stoppings, overcasts and other structures, abnormally high leakage results in an inefficient ventilation system and can increase operational costs for a mine. In addition, excessive leakage results in less control over mine ventilation airsplits. This may compromise the isolation of mine escapeways, and allow adjacent entries to be rapidly contaminated by products of combustion in the event of a mine fire. A study was conducted at ten mines in which air pressure and quantity measurements were taken near the active sections. The method used to construct the ventilation controls to separate airways was also noted. Stopping Construction Table 1 shows the type of stopping construction commonly used in the 10 mines included in this study. Main entry stoppings were normally built with mortared joints (wet stacked) with sealant applied to the high pressure side of the stopping. In most cases the entire face of the stoppings were sealed. In one case only the joints between blocks were sealed but the perimeters were sealed in all cases. Some mines also sealed the low pressure side of the stopping. Solid blocks were normally used for main entry construction and hollow blocks were sometimes used for longwall panel development. [ ]

Jan 1, 1993

By C. N. Speltz

The cost of transportation of industrial minerals is frequently higher than the cost of the product. Where a minc, industrial mineral plant. or customer is served by a single railroad, that single railroad, because of lack of competition, may increase rates and/or provide inadequate service so that the economics of that mine, plant or customer are adversely affected, or so that the level of service precludes satisfactory performance. Transloading of industrial minerals from rail-to-truck or truck-to-rail in order to improve economics and sales of product is not new. However, transloading to reduce freight costs or improve inadequate service caused by a lack of competition in the rail industry, is a relatively new concept brought on by the recent mergers of many of the main line railroads, both in the eastern and western United States. For example, with completion of the Burlington Northern/Santa Fe and the Union Pacific/Southern Pacific mergers, the western United States has only two mainline railroads. The industrial mineral producer, in order to reduce the cost of transportation, or increase the level of service, must evaluate potential transloading opportunities in light of higher prices and reduced service brought on by reduced railroad competition.

Jan 1, 1998

Introduction The science of geometallurgy requires that ore characterization be performed holistically, that is, with consideration of ore quality on plant performance and product quality. Geometallurgy, by definition, requires knowledge of geology, mineralogy, extractive metallurgy and pyrometallurgy. Examples of the ore characterization challenges for the geometallurgist in the production of iron ore pellets from low-grade Proterozoic iron formation are presented here. Ore Characterization Ore characterization, the quantifying of mineralogy, grain size and the texture of ores, is an often overlooked tool in improving process efficiency. Some mineral related problems are not unique to the production of iron ore pellets; e.g. the problems caused by expanding clays in flotation. The negative impact of carbonate minerals in the ore on pellet productivity and pellet quality are unique. The long and complex geologic history of Precambrian iron formations contribute to modifications of grain size and the introduction of complex textures.

Jan 1, 2007

By Wallace W. Roepke, James C. Church

Introduction The history of mining machine development has been and is one of "bigger is better." While this may not always be true, it is currently the only approach open to machine designers without specific in situ cutting force information. The US Bureau of Mines has tried to modify this approach by supporting research, both in the laboratory and in field tests, that demonstrates that bigger is only better if the equipment is used correctly. In-house research has shown conclusively that deeper cutting reduces both dust and energy (Roepke, Lindroth, and Myren, 1976). This approach has been applied to a large Lee-Norse (LN456) continuous mining machine, which confirmed that slow deep cutting is beneficial and does not have a negative effect on production (Black and Rounds, 1977; Black et al., 1978). Specific in situ engineering data needed on coal cutting forces have not, however, been obtained by field tests with the LN456. Although the machine was instrumented to obtain data on one seam, it would be impractical to move the machine from mine to mine. It became apparent that a portable, easy to use device was needed for obtaining coal cutting forces in situ. Such a device could provide the engineering data needed by manufacturers, the site-specific cutting information needed by operators, and a field research tool to establish correlation with laboratory cutting tests. Design and Development According to a set of design guidelines established by the Bureau, the proposed device should be able to measure cutting forces directly with standard coal cutting bits, be transportable by two miners, and be self-contained. These guidelines were given to Ingersoll-Rand Research Inc. (IRRI). The contractor performed a rigorous review of 16 existing in situ portable tester designs, none of which could use a standard coal cutting bit to measure cutting force directly. These included a wide range of devices, all of which are discussed in the final report (Banerjee and Wisner, 1982). A series of design tradeoffs were established to satisfy initial design guidelines. The final deci¬sion was heavily weighted by an industry survey to determine the preferred concept from four man¬ufacturers and four operators. To assure ease of use with maximum information, all respondents wanted a linear cutting device that could work in seams down to 1 m (40 in.). The final design choice was between a linear and a rotary tester. The rotary tester would cut on an arc in the same manner as a rotary drum machine. It had several faults, however, that could not be resolved within the scope of this contract. The worst was related to arc cutting (Roepke, Lindroth, and Myren, 1976). Since any bit cutting on arc enters the coal at zero depth, goes to maximum depth only at the horizontal centerline, and exits at zero, maximum cutting force is obtained only at the instant the bit crosses the horizontal centerline. This centerline peak response would require that many replications be made with an arc-type cutter to get reliable force information. The arc-type tester would require extensive face preparation to establish the cutting path and extensive wasted clearing cuts to advance the device for additional layers of cuts. The linear tester could be mounted directly on any face or rib with only minor preparation. By its nature, a linear cut at a constant depth provides multiple peak force data over the length of the cut, thus reducing the required number of replications. The final review of these two major design concepts plus indus-

Jan 9, 1983

By R. Sandefur

Recent advances in personal computer technology allow movies of various types to be made and displayed on PC and Macintosh computers. An interesting application of this movie technology is animation of computerized geological models. These movies, which can be most easily constructed along rows, columns, and benches of block models, allow an ore body to be more easily and quickly visualized than the traditional method of flipping back and forth through various sections.

Jan 1, 1994

By Ludger Suarez-Burgoa

At the Porce III Underground Powerhouse Complex in Colombia, the observational method (Peck, 1969), was used to provide information to an underground wedge and a 2D Finite Element Method numerical models, both used to estimate the mechanical behavior of the hoisting rock mass. The observation process consisted in collecting fault events in the field and in interpreting Multiple Point Borehole Extensometers and Surface Convergence Displacement Points readings. The calibrated models were employed to predict the behavior of the pillars formed between the machine and transformer caverns. The observational method is an important activity during excavation of large openings and provides arguments to understand the rock mass mechanical behavior.

Jan 1, 2008