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By Steven J. Page

Since 1982 standard calibration materials recommended for respirable crystalline silica analysis by the Mine Safety and Health Administration (MSHA) P7 Infrared Method and the National Institute for Occupational Safety and Health (NIOSH) X-ray Diffraction (XRD) Analytical Method 7500 have undergone minor changes in size distribution. However, a critical assumption has been made that the crystalline silica in ambient mine atmosphere respirable dust samples has also remained essentially unchanged in particle size distribution. The objective of this work, therefore, is to compare recent particle size distributions of underground coal mine dust and the silica component of these dusts with estimated aerodynamic particle size distributions of calibration standard materials MIN-U-SIL 5, Berkeley 5, and SRM 1878 used by two crystalline silica analysis techniques. This work provides resolution to a previously reported discrepancy involving the proper sample dilution for the particle sizing method used. Dust impactor sampling data for various locations in 13 underground coal mines were collected between 1991 and 1998 and analyzed for the respirable mass median aerodynamic diameters. The data suggest that the MSHA P7 Method will underestimate the silica content of the sample by at most 8.5 % in the median size range of 0.9–3.6 µm and that it is unlikely one would obtain any significant error in the MSHA P7 Method analysis when the method uses Berkeley 5, MIN-U-SIL 5, or SRM 1878 as a calibration standard material. The results suggest that the NIOSH Analytical Method 7500 would be more appropriate for a dust sample that is representative of the total (no cyclone classifier) rather than the respirable airborne dust, particularly since the mass fraction in the size range below 4 µm is usually a small percentage of the total airborne dust mass. However, the NIOSH Analytical Method 7500 is likely to underestimate the silica content of an airborne respirable dust sample by only 5–10 %. The results of this study also suggest that any changes that may have occurred in the median respirable size of airborne coal mine dust are not significant enough to cause any appreciable error in the current methods used for respirable crystalline silica analysis.

Aug 3, 2005

By L. H. Michaud

An existing mining company plans to extend their surface coal mining operations in central Pennsylvania. Although the proposed site contains pyritic material, the results of the acid-base accounting for the site show an overall NP:MPA ratio of 4.26:1 for the entire site. The pyritic material is, however, associated with the lower of the three seams to? be mined. If traditional dragline mining methods were used, this would result in the acidic material being placed at the top of the spoil piles, exposing it to air and water, and therefore to the formation of Acid Mine Drainage (AMD). Further concerns are associated with the fact that the alkaline material is not distributed evenly over the mine site. In order to prevent the formation of AMD, the company has proposed to mine the area using Truck/Shovel mining methods. This technique offers a more flexible alternative for cut sequencing and for the selective placement of waste material. Working with the Department of Mineral Engineering at The Pennsylvania State University, a mine plan was developed for this site. This plan results in the layering of the alkaline and acidic spoil in order to achieve an overall net alkalinity in all spoil areas. Mining of this site should therefore not result in the formation of AMD.

Jan 1, 1994

In April, 1961, a large pillar was removed by blasthole mining methods at Inco's Murray mine. More than half a million pounds of explosives were used to break 2,120,000 tons of ore in a single underground blast. The 5, 10 2-inch boreholes, involving 435,815 feet of drilling, were loaded over a three-week period, using conventional explosives and short-period electric caps. From this blast, more than half a million tons of ore have been removed to date (April, 1962) through slusher trenches situated in the footwall of the ore body. T HE Ontario Division of The International Nickel Company of Canada, Limited, operates six underground mines and two open pits in the Sudbury district in the north-central part of the Province. This paper describes the removal of a large pillar at the Murray mine (Figure 1), where 2,120,000 tons of ore were broken in a single underground blast by the blasthole mining method. This is the largest underground blast achieved to date (April, 1962) at Inco mines.

Jan 1, 1963

By R. T. Johansen

The wettability of petroleum reservoir materials by water and crude oil has been recognized as an important factor in determining waterflood behavior. The problem of simulating and studying reservoir conditions and characteristics, to produce laboratory results congruous with field experience, has been examined by many methods. (1, 2, 3, 4).4/ The most generally accepted approach to this problem is determining wettability of the porous medium by imbibition methods; or, after solvent extraction, predetermining the wettability of the samples by coating the surface of the medium to render it preferentially oil-or water-wet (3, 5). The emphasis of wettability studies has been placed on the characteristics of the reservoir materials with little consideration being given to the fluids themselves. Therefore, the wetting tendencies of crude oils have been investigated by a recently developed method (6). Imbibition and the more general phenomena of capillary pressure are the most direct ways to measure wettability. The simplest type of capillary pressure observations involves the use of an interfacial-tension capillarimeter, which does not produce as separate values either interfacial tension or contact angle. Actually, these quantities do not have independent significance in reservoir engineering, and the property of wettability they are intended to evaluate is given, immediately by capillarimetric measurements.

Jan 1, 1961

By A. R. Hasankhoei

Water recycling is of prime importance in mineral processing due to consumption of large amount of water. This issue is vital in areas with a low rainfall such as Shahre-Babak (Kerman, Iran). Deep cone thickeners, unlike conventional thickeners, have a low diameter to height ratio where large amounts of flocculants are used to increase particles settling rates. The high bed heights (over 10 m) produce enough pressure to create a paste type underflow indicating maximum water recovery. Flocculent preparation, conditioning, and distribution along with the bed height are among the important factors affecting the performance. Four deep cones are used to recover water from the tailings of the plant. Plant data indicated that the thickener never reached the target underflow percent solids (63%) and the average underflow concentrations were between 49 and 57%. By modification and optimization of flocculent preparation and distribution methods the underflow percent solids increased by 3-4%. This translated to an increase of water recovery from 67 to 73%. This was equivalent of 53000 m3 saving in monthly water consumption. Keywords: deep cone thickener, water recovery, flocculant, copper tailings

Sep 1, 2012

By Webster S. Anderson

The design, development, and testing of hydraulic props to provide a lightweight, high strength, recoverable unit is described. Modifications in design incorporating threaded components resulted in the development of a lightweight hydraulic prop fabricated from aluminum alloy tubing. This new design facilitates interchange and replacement of component parts, and the physical properties of the material are not changed as they are when welding is required. The new prop is 65,000 pounds and has a maximum load capacity of a lower weight-to-load capacity ratio than the hydraulic props now available.

Jan 1, 1967

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By W. O. Philbrook, K. Natesan

A mathematical model was derived for the temperature change in a particle undergoing heterogeneous chemical reaction and heal exchange with a stream of fluid. The model has been applied specifically to the exothermic oxidation of spherical particles of zinc sulfide. Very good agreement was obtained between the theoretical prediclion and experimental observation. It was found that at no stage of the reaction was the particle subjected to a state of thermal instability. In studies of gas-solid reactions, it is generally assumed that the gaseous and the solid phases are at the same temperature throughout the reaction period. This is not strictly true since most of the reactions are either exothermic or endothermic in nature. This paper deals only with heterogeneous gas-solid reactions of noncatalytic type. In the case of exothermic reactions (e.g., roasting of sulfides), there may result an increase in temperature of the particle due to the release of the heat of reaction, and this increase in temperature may enhance the reaction rate and thus cause further release of heat. Such a "spiralling" effect may lead to a thermally unstable situation within the particle. On the other hand, in the case of endothermic reactions (e.g., calcination of limestone), there may occur a decrease in temperature of the particle as the reaction proceeds, leading to a virtual "quenching" of the reaction front. A general discussion of autothermic processes with emphasis on properties and reactor design is given by Van Heerden. 1 Levesque and cubicciotti 2 observed self heating of the sample in the oxidation of thorium. Anous et al. 3 observed self cooling of the sample in their dehydration studies on chrome alum. Cannon and Denbigh, 4 and Ong et al.5 reported self heating of the sample in their studies of the kinetics of oxidation of zinc sulfide. A typical plot of heat generation and heat loss rates with reaction temperature is shown in Fig. 1.'14 The Sigmoid curve represents the heat generation rate while the straight lines represent the heat loss rate. The main objectives of this analysis are to estimate the temperature increase of a single particle undergoing an exothermic reaction and to study the possibility of a thermally unstable situation within the particle. Also, a study of the dependence of the tempera- ture increase on different process variables has been attempted. A "she ll-by-shell" method has been proposed to analyze the problem. The shells may be made as small as necessary, and equations may be set up for heat transfer by various mechanisms. Such equations will have to be solved for each shell, with the boundary conditions for ith shell being obtained from the (i - 1)th shell, using a high speed digital computer. The oxidation of spherical particles of zinc sulfide is considered for this analysis, but the treatment is general enough to be applicable to other reactions and particle shapes. The oxidation reaction can be written as AH =« -115 kcal/mole ZnS [1] Such a reaction has been found to proceed in a topo-chemical manner; i.e., as the reaction proceeds, a progressively thicker outer shell of zinc oxide is formed, while the inner core of unreacted sulfide decreases. It has been found experimentally, both in the present work and in the previous investigations,4-6 that the particle retains its original dimensions. Therefore, at any given time, the particle may be

Jan 1, 1970

By R. Schaller, Ch. Karcher, A. Thess

"Electron beam evaporation of liquid metals is an innovative technology increasingly used in industry to produce very thin coatings of high purity. In this process the surface of a metal ingot, placed in a water-cooled copper crucible, is heated by bombarding it with a high-energy electron beam gun. The material melts, forming a free surface and eventually starting to evaporate. The rising vapor cloud condensates as a thin film on a moving substrate. The strong energy input at the free surface gives rise to vigorous thermos capillary and buoyancy driven convective motion within the melt, leading to unwelcome heat losses. This paper aims to show how these losses can be reduced by optimizing the geometry of the crucible and by using external magnetic fields. We present both experimental and numerical results.IntroductionElectron beam evaporation of liquid metals [l] is an innovative PVD-technology which is excellently suited for the industrial production of high-quality coatings. In this vacuum process a high-energy electron beam heats up the surface of a metal ingot. The ingot melts and forms a free surface. A schematic view of the arrangement is shown in Figure 1. When the temperature of the free surface exceeds the actual (pressure-dependent) saturation temperature of the ambient gas, the liquid starts to vaporize. The rising vapor cloud finally condensates on a moving substrate located at some distance above. To guarantee a superior quality of the coating the melt is typically confined in a water-cooled copper crucible. The intensive cooling is necessary to avoid chemical reactions between the melt and the crucible walls, that is, crucible materials and their reaction products are practically excluded from evaporation."

Jan 1, 1999

By H. Wu, V. Izaxon

Intensive ventilation is required during the excavation of a mine opening and the subsequent exploitation of the mineral resources. The exchange of heat between the ventilation air and the rock mass has a major influence on the thermal regime of a mine, especially in the initial segment of the airway. In frozen ground, heat flow into the rock mass from the ventilation air can result in the eventual thawing of the permafrost, which can result in roof-rock instability. This paper presents a mathematical analysis of the dispersion process of ventilation-induced heat by thermosyphons placed on the surface of the mine. The principal uses of thermosyphons in the Arctic are to reduce the ground temperature to prevent potentially unstable permafrost from thawing and to increase the bearing strength of the material. The mathematical model of the physical problem, generally known as "the Stephan problem," takes into account the varying thermo-physical properties of air and the frozen and thawed rock mass, the thermal properties of the freeing fluid in the thermosyphons and the phase change of water. Finite-difference techniques were employed to determine the temperature isotherms around the thermosyphons. It was found that two rows of thermosyphons on a 10- x 10-m grid provide an acceptable design spacing for reducing the ground temperature around the mine airway.

Jan 1, 1997

By F. N. Kissell, R. J. Bielicki, J. H. Perkins

A sorption apparatus for measuring the equilibrium sorption isotherm and the rate of diffusion of methane from fine-sized coal was designed and constructed. Preliminary work that used Pittsburgh and Pocahontas No. 3 coal shows that the diffusion coefficient varies with pressure and that the fracture spacing in these two coals is substantially different.

Jan 1, 1972

By Lendall P. Warriner, Blandford C. Burgess

Jasper County lies just north of the geographical center of Georgia, bounded on the west and north by the Ocmulgee River. The county seat, Monticello, is approximately 65 miles east-southeast of Atlanta and 40 miles north of Macon. It is served by the Macon-Athens branch of the Central of Georgia Railway Co. and by five paved highways converging from adjoining county seats. Farming and lumbering are the principal local occupations. About half the county is covered with pine and mixed hardwoods; a large part of the forested land is owned by the Soil Conservation Service of the Department of Agriculture in its Piedmont Land Utilization Project. The topography is one of gently rolling field and forest with occasional hills rising 100 ft or: more above the creek beds. Though annual rainfall is in the neighborhood of 60 in., most of it falls in torrential downpours of short duration which tend rapidly to erode the red soil and convert the country roads to greasy tracks of red mud.

Jan 1, 1949

In accordance with the usual custom, manuscripts to be presented at the next meeting of the Institute must be in the hands of the, Secretary before July 1, 1917. Manuscripts to be properly distributed in advance of the meeting must be in the hands of the Secretary before June 1, 1917. At the last meeting of the Institute 45 papers were presented, and 42 of them were actively discussed. This experience makes us believe that advance distribution is the best possible means of getting a good discussion of papers and of securing for them the interest of the members of the profession. It will be impossible, however, to give the papers proper distribution if they bare not received before June 1, 1917.

Jan 4, 1917

By F Wu, H Wang, W Chen

A pressurisation method with multiparallel fans has been adopted at HongLiuLin Coalmine to control the significant external air leakage through the goaf area of coalface 15207. Potential damaging effects of large external air leakage on a coalmine’s ventilation system are analysed according to the current air leakage status caused by the exploitation of shallow-covered coal seams. The use of pressure balanced ventilation to prevent external air leakage through goaf areas is introduced. The calculation of ventilation network of HongLiuLin Coalmine shows that the air resistance of an external air leakage branch through a goaf area and the air pressure drops from the intake slope to the working face can be adopted to determine the values of pressure to be increased by regulating fans and airflow controlling parameters of regulators. By investigating airflow requirements of the pressure-balanced area as well as the performances of the auxiliary fans currently used in the coalmine, the number of pressurisation fans needed is fixed. The expected result is achieved after the pressurisation method has been carried out, which can be helpful for other coalmines to handle similar problems.CITATION:Wu, F, Chen, W and Wang, H, 2015. Multiparallel fan method to prevent significant external air leakages through a goaf area of a shallow covered coalmine, in Proceedings The Australian Mine Ventilation Conference, pp 267–270 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 31, 2015

By J. H. Puffer

The copper mineralization of the Mesozoic Passaic Formation of New Jersey is typical of copper deposits in eastern North American Mesozoic sediments. Most such deposits are located at the base of the first of several basalt flows that are interbedded with continental, red-bed sediments. The copper mineralization is similar in most respects to that of the Keweenawan Basalts of the Lake Superior area. The first basaltic flow units of the eastern North American Mesozoic basins, and the Keweenawan Basalts, are continental quartz tholeiites of similar chemical composition and copper contents. The copper minerals and associated zeolite facies metamorphic mineral assemblages of these two rock units are also approximately the same. Stoiber and Davidson (1959) suggested that Keweenawan copper was mobilized and concentrated during metamorphism. The evidence pertaining to the Mesozoic copper deposits of New Jersey is consistent with such an interpretation and also with DeVore's (1955) suggestion that ore fluids commonly arise as a result of the mobility of metals under metamorphic conditions. There is additional evidence that final metamorphic concentration of Mesozoic copper was preceded by an initial stage of hydrothermal precipitation related to fluids emitted by basaltic magma.

Jan 1, 1987

By M. E. Poad

In 1986, a cooperative effort by the Homestake Mining company, the U.S. Bureau of Mines and the University of Utah was initiated to determine shaft stability while recovering high grade ore in a nearby pillar. Premining finite element simulations indicated the shaft would remain in elastic ground. Six years have elapsed and, despite some initial concern during the first lifts, it appears that the bulk of the original reserve will be safely and profitably recovered. An integrated application of rock mechanics tools from laboratory testing through numerical modelling to mine measurements and observations contributed substantially to the operational success of the project.

Jan 1, 1994

By Eric M. Klein

As part of the Woodrow Wilson Replacement Bridge project several damaged wetland and environmental areas were to be mitigated. One of the mitigation projects consisted of constructing a Denil fish ladder around Peirce Mill Dam in Rock Creek National Park, Washington, D.C. The dam is a historic structure operated by the National Park Service. The fish ladder is located between the east abutment of the dam and Beach Drive, a major collector road. Construction of the fish ladder required removal of a portion of the east abutment and an excavation that extended about 22-ft below the pavement of Beach Ave. It was not possible to close Beach Drive except for very brief intervals or at night; therefore, a support of excavation system that would minimize disruption to the public was required. The high groundwater required that the groundwater be cut off prior to excavation; the random rubble and soil fill, and shallow but erratic rock surface precluded driving sheet piling. A jet grout system was selected because the grout probe could penetrate the rubble fill, stabilize a variety of soil conditions, and cut off the ground water prior to excavation. It was not possible to construct a stable wall without tiebacks or soil nails. Given the schedule and tight working conditions it was decided that soil nails through the jet grout wall facing would be the most cost effective method, as it did not reduce the working space as rakers would and would be more cost effective than prestressed tiebacks.

Jan 1, 2006

By Frank F. Aplan, Edward C. Dowling, Richard R. Klimpel

Several series of time-instantaneous samples were taken on a cell-by-cell basis in the flotation circuits of the Pinto Valley, AZ copper concentrator. These samples were used to calculate mass flow rates throughout the circuit using standard, deterministic metallurgical formulae and also a variety of more statistically oriented techniques. Because of the magnitude of the errors associated with the individual cell samples, the standard metallurgical formulae are not sufficient. Consequently, it is necessary to utilize some type of statistical technique capable of distributing the individual cell sample errors in such a way so as to provide a physically useful mass balance for the entire circuit. This paper explores the use of linear and least squares programming as one technique to providing credible mass balance calculations. For linear programming this was done by distributing the error using two different criteria involving specific linear programming formulations: least absolute sum minimization (which assumes an underlying double exponential distribution error structure) and minimax (which minimizes the maximum absolute deviation and assumes an underlying rectangular distribution error structure). These results were then compared to the more standard formulation of minimization of the sum of least squares which assumes an underlying normal distribution error structure. The least absolute sum approach was particularly attractive due to its simplicity of use and the ease in which repetitive calculations can be done with no user intervention.

Jan 1, 1990

By A. Rojo, N. J. Ramirez, J. E. Salgado

"This work aims to characterize the plastics used in a handgrip located in the car interior body in order to improve its design process and be able to predict its behavior when the part is subjected to impact test involving human casualties.Through stress analyses at high speeds and a video camera set up (front and side view), the reduction of area of specimen was obtained from image analysis, allowing to get the stress-strain real curve of the material.Continuing the characterization process a set of impact test was developed, where standard lzod configurations were used in unnotched specimens. From these tests were gotten displacement curves and absorption energy estimations.The proposed testing procedure allowed the acquisition of force and displacement data through time at various strain rates which are indispensable for the characterization of dynamic events. Using a specimen with a special configuration has shown a more detailed dynamic behavior resulting into more accurate numerical analyses within the design process.IntroductionIn the design process, one of the many analyses which take place is the impact test of the designed part. Nowadays there are many software capable to do these kind of analysis and each industry use the one they think it is convenient for them, but also the selected software has mathematic models that help to find the solution and the simulation of the case of analysis. These are not the only considerations that a company has to take into account, other one is the type of information and data necessary to feed the software. For this study we are going to focus on the acquisition of these data.The mail reason of this work is that generally the supplier gives the card of the material with all the characteristics and information requested by the buyer, but all of that information is acquired by standard methods, but it is inconvenient when the materials are polymers."

Jan 1, 2011