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By George W. Toth

In a recently completed research project for the U.S. Bureau of Mines - Twin Cities Research Center, a computerized costing procedure for uranium in situ leach mining was developed by the NUS Corporation. This costing procedure, termed a cost model, employs a process engineering approach for estimating total project costs as well as equipment and manpower requirements for uranium in situ leaching operations in either Texas or Wyoming. 80th capital and operating costs are generated by the model along with the overall production cost/pound U3O8 subject to a known rate of return. Conversely, rate of return on equity can be solved for, subject to a given price of yellowcake. During the course of the research effort, this cost model was applied 50 typical site conditions in Texas and Wyoming. Sensitivity rests were conducted for numerous key parameters to determine the cost influence attributable to incremental changes In parameter values. The results of these sensitivity tests along with a description of the in situ leaching cost model have been documented as Bureau of Mines, Mineral Research Contract Report No. J019912, "Cost and Sensitivities Analysis for Uranium In Situ Leach Mining," dated March 1981.

Jan 1, 1981

By Ronnie Melker, Joel Webster, Daniel Belardo, Thomas Hyatt

Quality assurance of drilled shafts is highly dependent upon the construction practices used during the excavation and concrete placement process. When performed under slurry, challenges are presented in maintaining borehole stability and placing a uniform and continuous concrete column. An integral part of the construction operation is to evaluate the as-built integrity to verify the drilled shaft is capable of safely transferring the required loads to the subsurface. One such case is exemplified by a project involving the expansion of an existing public transit rail line. The expansion required the construction of a new bridge bent supported by seven 5-ft diameter drilled shafts of various lengths. The specifications for the project required both Cross-Hole Sonic Logging (CSL) and Thermal Integrity Profiling (TIP) for drilled shaft integrity evaluation. Both of these post-construction non-destructive tests (NDT) are designed to detect anomalies or the presence of regions with reduced concrete quality. However, the implementation and methodologies of performing each of these NDT methods are significantly different. Cross-Hole Sonic Logging uses ultrasonic signals emitted from a transmitter in one access tube to a receiver in another access tube. Thermal Integrity Profiling measures elevated concrete temperatures during the curing process using embedded thermal wires. Both test methods require different access details and therefore have different costs associated to the contractor or owner to perform. This paper provides a detailed comparison of these NDT methods as performed on these seven drilled shafts. The associated financial and time costs to perform each method are compared and the results are discussed in detail. INTRODUCTION Drilled shafts are a fundamental deep foundation element that can be designed to resist large bending moments as well as to transfer large compressive axial loads through weaker upper soils to competent soils or rock below. Often times when the upper soils are very soft, lateral support by casing or slurry, or a combination of both, may be used to hold back the softer soils from collapsing into the hole before concrete has been placed. Due to the blind nature of drilled shaft construction and the qualities issues that may arise during the placement process, it is essential for design engineers to specify non-destructive testing (NDT) to assess the quality of the as-built drilled shaft. CSL and TIP testing are two NDT methods designed to detect areas of lower quality concrete or soil inclusions throughout the length of the drilled shaft and determine its overall quality. Each method uses different technologies and methods to achieve the results required to determine the as-built condition of the drilled shaft.

Jan 1, 2019

By M. K. Buder

Over the years, electric utilities have been faced with declining steam-coal quality, rising coal prices, and increasingly rigid and costly environmental standards. As a result, many utilities are beginning to show an interest in coal cleaning to upgrade coal by improving its heating value and reducing its sulfur content. Among the primary questions that must be answered for these electric utilities is whether the cost of coal cleaning can be offset by the potential cost savings in coal transportation and utilization in power generation. In search of at least a preliminary answer to this question, the Electric Power Research Institute (EPRI) engaged Bechtel to perform a case-by-case economic study to examine the cost effects of using cleaned versus raw coal to fuel new electric power plants. The results of two of the seven cases studied which involved western coals - one from Montana and another from Utah - are reported in this paper. Coal cleaning appeared to be a cost effective option for the Utah coal, but only marginally economical for the Montana coal. These cost results are expected to change to favor coal cleaning to a greater extent with the application of the final revised New Source Performance Standards regulating SO2 emissions just issued by the EPA. The study was based on the regulation originally proposed by the EPA under the 1977 Clean Air Act Amendments.

Jan 1, 1979

By Gordon M. Pugh, David G. Rasmussen

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

Jan 1, 1982

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Sublevel long-hole stoping is a large tonnage stoping method that may be used when there is a fairly com¬petent straight vein and the walls will stand with little support. It is a method that lends itself to a high degree of mechanization. Two variations of the method are illustrated in Figs. 1 and 2. Fig. 1 illustrates development typical for a 3-m (10-ft) wide vein. Initial development is by a lower haulageway, which may be either track or trackless. A two-compartment raise is driven in the footwall of the vein. At about 12 m (40 ft) above the main head¬ing, a footwall scram is driven and pockets through the vein are cut on 7.6-m (25-ft) centers. Each of the sub¬level blocks selected in this example is 45.7 m (150 ft) long by 12 m (40 ft) high. Stoping starts at the slot raises at stope endlines and retreats to the timbered raise. This retreat sequence starts on the lowest sublevel and the opening cut is blasted into the slot raise. This retreating system is used in order to provide a good safe back over the miners working in any given sublevel. A crawler-mounted down-the-hole or conventional pneumatic drill puts the holes down from one sublevel to the next on any desired drill pattern. Gravity flow brings the broken ore down to the scram level. The illustration shows a large slusher mounted at the timbered raise and slushing from the pockets back to the raise chute. Development of the ore block in Fig. 2 is typical for a 5.5-m (18-ft) wide vein. Initial development is by a trackless heading driven in the footwall of the vein. Pockets are then put into and through the vein on 7.6-m (25-ft) centers. For ease in using trackless rubber tired mucking equipment, these pockets are crosscut to the vein at approximately 1.04 rad (60°) rather than per¬pendicular. A two-compartment timbered raise is driven down from one level to the next. Endline slot raises are again put through from level to level. Each of the sub¬level blocks selected in this example is 40-m (130 ft) long by 33.5 m (110 ft) high. Stoping starts at the slot raises and retreats back to the timbered raise. Long holes are drilled from one sub¬level down to the next by starting between the lowest pair of sublevels and then working up the ore block. Mucking out of the pocket drawpoints is accomplished by rubber-tired load-haul-dump (LHD) units which then dump into a centrally located orepass.

Jan 1, 1982

By Carson Rife, Michael L. Clark

INTRODUCTION Zortman Mining Inc. is a wholly owned subsidiary of Pegasus Gold Corporation of Spokane, Washington. Pegasus has been producing gold and silver from the Zortman and Landusky mines since 1979. The Zortman and Landusky mines are two distinct open pit mines located approximately 6 kilometers apart in the Little Rockies Mining District of north central Montana. The two mines use large-tonnage, low- grade heap leaching techniques to efficiently process the gold and silver ore. PROJECT DESCRIPTION The Little Rockies rise abruptly 914 metres from the surrounding plains to about 1829 metres above sea level. The topography is quite steep, making efficient leach pad construction a major challenge. The climate is semi-arid with moderate summers and occasionally severe winters. Freezing temperatures usually arrive in late October or early November, influencing the effectiveness of leaching operations. GEOLOGY AND MINERALOGY The geology of the Little Rocky Mountains consists of steeply dipping sedimentary ridges forming an elliptical dome surrounding a core of intruding syenite porphyry. Within the syenite porphyry are well fractured and mineralized shear zones where the existing reserves and open pits are located. The ore is presently divided into oxidized, mixed, and sulfide. Only the oxides and mixed zones that are of ore grade are placed on the leach pads. The syenite porphyry host rock enables effective leaching of gold and silver to depths of 61 metres, having a major impact on leach pad construction costs. MINING All mining since 1979 has been by contractor. This enabled initial mining to commence without capital investment in equipment. Through competitive negotiations, Pegasus has been able to keep contract mining costs at levels very close to what would be experienced if Pegasus had purchased equipment initially. Contracts are negotiated yearly and are based on a unit price and haul distance for ore and waste. Haul distances have ranged from as little as 457 metres to as much as 3,050 metres. At present, Pegasus is developing a 27,000,000-tonne leach pad at the Landusky project that is adjacent to the minine area. This design feature will reduce the mining cost by 20 percent over the life of the mine. In 1987, over 5,900,000 tonnes of ore and 3,200,000 tonnes of waste will be mined at the Landusky project. Through 1986, 13,763,000 tonnes of ore have been mined at a waste-to-ore strip ratio of 0.60:l at the Zortman project, and 13,005,000 tonnes of ore at a waste-to-ore strip ratio of 0.79:l have been mined at the Landusky project. Current mining reserves are 32,769,000 tonnes of ore at a waste- to-ore strip ratio of 0.59:l. Through efficiencies of large-scale heap leaching, Pegasus has been able to lower the mining cutoff from 0.34 grams per tonne gold to 0.27 grams per tonne gold. This has reduced the mining cost per tonne of ore h percent through a reduction in strip ratio of 14 percent. Beginning in 1984, all pits were designed based on computer-generated "floating cone" pits. The geologic models were developed by Pegasus staff with a series of economic models evaluated. The resulting computer-generated pits were used as a guideline in designing the mine plans, including haul roads for final calculation of tons and grade. Input to the floating cone models consisted of mining cost, processing cost, cutoff grade (0.34 grams gold per tonne), recovery (67 percent) and gold price. Due to the gold being deposited primarily on the fractures of the syenite host rock, no crushing of ore is needed prior to leaching. This characteristic of the ore is a key in being able to profitably mine and process gold ore at an average grade of 0.69 grams of gold per tonne. Strict blasting control is also a key in success of the leaching operation. A system of splitting ore and waste in blast patterns has been devised to reduce dilution. Also, drill hole spacings are varied from 3m x 3m in breccia ore to 4m x 4m in syenite ore to 5m x 5m in waste. Drill holes

Jan 1, 1987

By Lee A. Goetz

Cost control is a lot like motherhood - everyone proclaims its need and its worthiness, but is it really appreciated and used to successfully complete a project on time and within budget? Good Cost Control is a yardstick - requiring fixed definition but this does not mean rigidity for the project. It provides the means for telling: 1. Where you are going 2. When you will complete and 3. What it will finally cost. Cost Control does not fix the price of dreams nor the cost of projects loosely defined. It does not control costs in itself but provides the tools whereby engineering decisions can be made which will ultimately affect the project cost It does not even guarantee that a project will be completed within the original funds - but only, if this reflects management desires.

Jan 1, 1970

By F. H. Mertens

The costs of an open-pit operation can be classified according to the following hierarchy : -overall cost of production (per ton of ore) -costs of basic operations and functions -costs of auxiliary operations -costs of machines and equipment -maintenance costs. Production cost is a measure of the open-pit mine performance and should be the final target of the management control. It is also a measure of the share of the mine in the processing stages.

Jan 1, 1990

By G Lane

This paper discusses the factors that contribute to the cost effective design of a concentrator. Concentrator design and layout outcomes are functions of the team (engineer’s and owner’s) participating in each particular project. The benchmark for relatively modest projects was set in the 1980s and 1990s during the ‘gold boom’ when numerous cost effective plants were designed and constructed on a lump sum basis in a very competitive market. A number of factors contribute to lack of transference of the lessons learnt in gold plant design to concentrator design including established paradigms in the design and layout of concentrators, lack of experience in cost effective design, operator’s preferences for flowsheet and layout and simple lack of appreciation of the impact of plant footprint on materials quantities and resultant capital costs.Experiences with recent copper concentrator projects (both small and large) are used as case studies.

Aug 8, 2011

By Phil Wilkinson, Sam Whitaker

"The installation of large (XL) diameter monopiles offshore typically requires expensive and specialist plant, equipment and vessels. Due to the increasing size (weight, diameter and length) of the foundations, in order to mitigate installation risk, the use of large diameter drills is becoming an increasing requirement.Project developers often try to avoid drilling operations due to the time associated with the drilling process. This restricts projects in terms of location and it also overlooks the additional engineering benefits drilling can bring to monopile foundation design e.g. reduced foundation material and better fatigue design life.By comparing the lessons learned of three milestone overwater drilling projects; methods of maximising drill speed and decreasing drilling process time is established. Particular emphasis is put on pile internal relief drilling in a process known as Drive-Drill-Drive (DDD).The milestone overwater drilling projects are North Hoyle Offshore Wind Farm (OWF) (2003); Flamanville EPR cooling water outlet shaft (2008); and Gwynt-y-Mor OWF (2014).By evaluating the effect of foundation pile diameter on drilling process times, example foundation installation costs are derived and used to show that incorporating drilling as part of a process may not be cost prohibitive to projectsContinued advances in drill technology, and drilling rates of penetration, combined with an integrated approach to foundation design, challenge whether hammering only is always the optimal solution for XL monopile installation."

Jan 1, 2017

By Richard D. Luark, Fernanda S. Madrona

Deep foundation construction in highly developed urban settings pose unique challenges of working with deep excavations in a relatively confined area while preserving the integrity of nearby structures. For construction of the below grade parking for two, 40-story residential high-rise towers, shoring walls were required to support the maximum 70-foot-deep vertical excavation in downtown Seattle, Washington. The design of the excavation support system consisted of soil nails combined with closely spaced vertical elements to allow the top row of soil nails to be installed up to 10 feet below the ground surface avoiding conflicts with adjacent underground utilities. To limit deflections next to a seven-story building, the upper rows of soil nails were post tensioned. An instrumentation program was implemented to monitor the deflections in the shoring walls by means of vertical inclinometers and optical survey monitoring. The shoring wall performance, designed to limit deflection to less than 1-inch to meet the City of Seattle requirements, was validated by the instrumentation program. This paper provides a discussion of the design and construction of the shoring walls as well as the monitoring data collected from the instrumentation system. PROJECT INFORMATION The project consisted of two residential high-rise towers, with five levels of below grade parking, located at 707 Terry Avenue Seattle, Washington. The project site covers a half city block, approximately 120 feet by 240 feet, and is bounded by Terry Avenue to the east, Columbia Street to the north, Cherry Street to the south, and an alley to the West. Prior to the development, site was occupied by surface parking. A 7-story concrete building with two stories of below grade parking, owned by the Archdiocese of Seattle is located just west of the 16-foot-wide alley. Numerous below grade utilities consisting of electrical, telecommunications, gas, water, sanitary sewer, steam, and storm drains, surround the project. The perimeter below grade parking walls extended within inches of the property boundaries. Topographic relief across the site was on the order of 8 feet. Figure 1 illustrates the site plan view.

Jan 1, 2019

By D. W. Horsfall

The different types of processes used in fine coal beneficiation in the Republic of South Africa are listed, together with data giving an indication of their separating efficiencies The two classes of coal, fines and ultra-fines, present in the -0,s m fraction of the run of mine coal are discussed, with special reference to their beneficiation characteristics. Using those properties, together with the data on plant efficiency, case studies are worked out to indicate the advantages and disadvantages of the different processing routes available. The paper closes with an indication of areas where research is required to enable some of the processes to be applied to a greater extent than is presently the case.

Jan 1, 1988

By John Chadwick

In the drive for faster development advance rates, railbound systems can be advantageous, in the right application. Skanska BS, the contractor to mine-owner VSK Mining, is driving an adit access deep into the Volouské Urchy Mountain in Slovakia to access the Gemerská Poloma talc deposit, some 70 km (43 miles) west of Ko?ice, the economic and cultural center of eastern Slovakia. Skanska began driving the Stolna Elisabeth drift in February 2007, extending 25 kg/m (50 lb/yd) rails (with 750 mm or 30 in. between sleepers) into the mountain to provide the main haulage to bring talc ore and waste out of the new mine once it gets into production. It will also be used for personnel transport, and is now almost complete. As a second egress and for ventilation, a 2.5-m- (8.2-ft-) diameter shaft of about 300-m (984-ft) depth will be sunk from surface down into the orebody. Stolna Elisabeth is a 4.2-km- (2.6-mile-) long drift mined at a cross section of some 11 m2 (118 sq ft). The finished profile is 3.2-m- (10.5-ft-) wide by 3.6-m- (11.8-ft-) high. It has been developed on a very slight incline so that water runs down it to the portal. The portal is 450 m (1,476 ft) above sea level and the end of the tunnel is 475 m (1,558 ft) above sea level. Much of the rock through which it has been developed is magnesite with a compressive strength of 150 MPA. The drilling jumbo is an Atlas Copco twin-boom Boomer 282 Raildrill drilling 42- to 45-mm- (1.6- to 1.7-in.-) diameter holes to a maximum length of 2.5 m (8.2 ft), depending on ground conditions. This is a drill jumbo that Skanska has used successfully on a number of projects during its life. The condition of the rock through which this drift has been driven is extremely variable. Five classifications of A1 and A2 and B1, B2 and [ ] B3, where A1 is the most competent rock and B3 is the worst, have been established. The amount of support used, of which more later is determined by the classification of ground conditions. At the time of the author?s visit, the project it was about 40 days ahead of schedule, thanks to high advance rates. For instance 6 m/day (20 ft/d) was the plan for the advance to be achieved in B1 rock, but sometimes the crews have achieved 8 m/day (26 ft/d) ? which is the average for the whole project ? in those conditions. In A1 the best advance rate achieved is 12 m/day (39 ft/d). The loading unit is a Häggloader 9HR-B on crawler tracks with a pony track to run on the rails. This allows the unit to be towed up and down the drift along the rails. To load from the face, the Häggloader lowers the foot of its onboard conveyor system to the floor and the backhoe pushes material on to it. This digging arm-conveyor combination produces uninterrupted, optimal volume loading of the haulage unit, without spill ? a technique offering load capacities as high as 180 m3/h (6,356 cu ft/hour). Further-more, as an electrically powered machine, it offers great

Jan 1, 2009

By V. C. Rogers, K. K. Nielson

INTRODUCTION The potential health impact from uranium milling operations is mainly associated with long-term releases of radioactive contaminants from the mill tailings.(1,2) The major mechanisms for mitigating these potential releases focus on increasing the tailings containment with the addition of migration barriers such as thick earthen covers and clay liners. Some limited investigation has also focused on reducing the radionuclide source terms. This alternative approach has some desirable features, but stringent cost requirements are placed upon source removal methods in order for them to be economically favorable. A cost effectiveness evaluation is presented herein, in which costs for containment methods are used to establish maximum cost guidelines for the source removal methods. The 230U decay chain, shown in Figure 1 contains the radioisotopes of major concern, 230Th, 226Ra, 222Rn and its short-lived alpha-emitting daughters, When the uranium is removed in the milling process, these isotopes, which are normally not removed remain in the tailings solids. The general pathways to the environment from uranium mill tailings are depicted in Figure 2. The dominant pathways are radon to the atmosphere and radium and thorium to a groundwater aquifer. The magnitude of the contaminants released into these pathways can be characterized by the following parameters: 1. The 230Th and 226Ra content. 2. The emanating power for radon. The fraction of radon generated that is free to migrate. 3. The leach rates. Source reduction methods seek to reduce the magnitude of any or all of these parameters. In general the dose rates are proportional to the source terms. For radon production the source term is proportional to the product of the radium concentration, R, and the emanating power, E, so that the risk is reduced by the same factor that reduces RE. For example, if the radium content is reduced by 95 percent, but E is increased by a factor of two, then the dose rates are reduced by an order of magnitude. The source term for the groundwater pathway is not as easy to identify. The peak in the radium concentrations in the accessable water body can also

Jan 1, 1981

By A Misra

In pursuit of solving the development rate problem the coal industry has investigated various development systems. While all the systems may improve the development rate they may not necessarily he cost effective. This paper carries out an economic analysis by developing cost models for present and future development systems for retreating longwall faces with particular reference to Australian longwalls. A critical cost analysis of various development systems is presented by taking into consideration all the costs associated with development systems. The costs have been divided into equipment, face and panel services, belt and structure, wages, roof and rib support, ventilation, consumables, power and maintenance. The annualised cost of various equipment has been obtained on the basis of depreciation rates prescribed by the Australian Taxation Office.

Jan 11, 1996

By A. J. MacKellar

The economic viability of any project is generally assessed on an evaluation of the total capital constructed costs along with operating costs, owner's costs and market conditions. This paper will be directed to the role that Cost Engineering plays in the development of the capital costs and the cost control techniques applied thereto.

Jan 1, 1975

By O. Schumacher

Guidelines for publication of information defining the amount and grade of resource material available in a deposit are well established in Australasia?s JORC Guidelines, Canada?s National Instrument 43-101 and similar documents in South Africa, Peru, Chile and elsewhere (Table 1). These guidelines provide standards to apply regarding the geologic and geostatistical basis for describing the level of confidence to be assumed for an estimate of the amount and grade of material. All of them provide for a strict distinction between material that has been demonstrated to be economically mineable and that which has not, and all of them provide terms such as ?resource? or ?exploration information? to be applied to data regarding material that has not been demonstrated to be economically mineable. It is not until an economic study has been completed that we are allowed to apply the powerful terms ?reserves? and ?ore? to our deposit material. Indeed, the application of these eye-catching labels immediately tells the reader that an economic study has been completed by a qualified person, and that the results of the study were positive. While extensive guidelines are established for describing the level of confidence in published geologic data, and the nature of the economic study that should be accomplished before using the terms ?reserves? or ?ore?, surprisingly little guidance is provided regarding the process to use, and the level of detail to attain, for estimating capital and operating costs for the study. The intent of this paper is to lay the groundwork for development of suitable guidelines for the cost estimating process. While the codes do not provide specific instructions for the cost estimate, they do provide numerous hints and clues to help determine the level of detail expected and the type of procedure to apply in preparing the estimate. This paper will first review the various codes to gain an insight into the thinking of the drafters of the codes regarding their unstated expectations for the cost estimating process. Based on this review, a series of guidelines are proposed for this purpose.

Feb 27, 2013

By William Grady

FACTORS entering into the market value of coal are its grade, and the cost of labor, material, and capital. Reduction in these costs cannot be expected in the future, and it therefore follows that greater economy in their use must be accomplished if it is desired to hold the present markets; more tons of product per unit must be obtained in a given time from both men and material, and means devised for increasing the percentage of lump and domestic sizes. This paper is presented for the purpose of inviting discussion and stimulating thought in regard to the influence of methods of procedure and plans of mining, upon the quality of coal, the economical use of labor, material, and capital, and the cost of production. Even the better methods of the. present day leave much to be desired, and a review of the reasons and necessities for their use may be of value in pointing the way to more economical mining. The writer does not attempt to discuss untried methods of mining, but rather to state clearly and concisely a summary of the results obtained under the several methods of procedure that have been adopted. In some instances a high degree of concentration has been effected, and comparisons of the results obtained with the results of lesser degrees of concentration form an interesting study. It is believed that much may be accomplished in improving the quality, and in the more economical use of labor, material, and capital, by concentration, and it is to concentration in particular that your study is invited. The writer, who .has occasion to visit many mines, most of which ire under different management, has been inclined to the opinion that improvement has been retarded in many ways and that mining has not as yet been entirely freed from the early-day practices which were forced upon it.

Jan 5, 1915

By Arthur F. Johnson

THE principal costs of making a pound of aluminum are for 9 kw-hr of electricity and for 1.9 Ib of the oxide (A1203) called alumina. A pound of alumina is made by digesting 2 or 3 lb of bauxite in hot caustic solutions, settling and filtering out the oxides of silicon, iron, and titanium and precipitating and calcining the A1203 as a pure, white powder. The cost of foreign bauxite delivered at an alumina plant in the U.S. may equal or exceed the cost of its chemical treatment to make alumina. Some of the principal factors affecting costs in the utilization of foreign bauxites are: transportation from mine to market, mineralogical and physical composition of the bauxite, thickness of the bauxite beds and their clay and forest overburden, amen- ability of the bauxite to beneficiation and to caustic extraction of the alumina, and lastly size of bauxite deposits and rate of mining employed.

Jan 6, 1954