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By Harry Littleton, Jr. Scarber, Wanliang Sun

"The Lost Foam Casting (LFC) process is relatively immature when compared to other casting processes. To make the process more robust, computer modeling is needed to reduce lead times and casting defects. In this study, metal filling of lost foam aluminum castings was visualized using state-of-art real time X-Ray technology. Metal filling of the lost foam casting was simulated using a commercial software package. The result of the computer simulation was compared with the results of the real time X-Ray visualization. Some of the computer simulation results agreed with the real time X-Ray visualization, but some did not. Based on other studies using real time X-Ray technology, new input variables, such as foam density variations, were developed for the computer modeling. With the new variables, modeling of the lost foam casting process becomes more realistic. IntroductionMold filling of the Lost Foam Casting (LFC) process is an important step toward making a quality part. In the LFC process, molten metal is poured against a refractory coated foam pattern surrounded by dry sand. At the pouring temperature of aluminum, the foam pattern is decomposed into 30% (wt) gaseous and 70% (wt) liquid foam pyrolysis products. Existence of the gaseous foam decomposition products significantly increases the back pressure on the metal stream compared with the conventional open cavity sand mold. Entrapment of the liquid pyrolysis products during the metal filling generates many casting defects, such as internal pores, fold and blisters. Existence of these casting defects sometimes seriously degrades the quality of the components manufactured by the LFC process. Understanding and controlling the metal filling of the LFC process will certainly make this process more robust."

Jan 1, 2004

By Elton L. Litchfield

The Bureau of Mines determined minimum energies for direct initiation of expanding gaseous detonation waves in acetylene-oxygen mixtures. Composition limit ranges for the initiation of detonation with fixed energies have been compared to data in the literature. Assuming that the stored electrical energy is completely converted to thermal energy, the agreement between the energy of primary explosive initiators and the energy of electrical discharge initiators was good. Minimum energies for initiation of detonation in the most sensitive composition (40 percent C20 plus 60 percent 0 2) were: 0.64 joule at an initial pressure of 1/4 atmosphere, 5.0X10-2 joule at 1/2 atmos¬phere, and 3.7X10-3 joule at I atmosphere. Fuel concentrations in mixtures initiated to detonation by 4.9X102 joules ranged from 10 to 65 percent C2H2 at 1/4 atmosphere, 10 to 67 percent C2H2 at 1/2 atmosphere, and 9 to 68 percent C2H2 at 1 atmosphere.

Jan 1, 1967

By J. B. Watts

"THE Mine Engineering Department at Gaspe Copper is staffed by a chief engineer, assist-ant engineer, research engineer, chief surveyor, four instrument men, chief bonus clerk, draftsman, and a helper pool of four permanent men plus one or two students in the summer. The men are grouped according to function for organization purposes. Routine services provided by the Mine Engineering Department in-elude ail surface and underground surveying and mapping, preparation of working drawings, compilation of engineering reports, recommendations on ventilation, testing the mine air for carbon monoxide, research on mine equipment and its use, and the calculation of incentive bonus. Demand jobs from departments other than the Mine Department .are done on request. Such jobs may be in the plantsite, the Town of Murdochville, or at the shipping points at Gaspe harbour or the St. Lawrence river."

Jan 1, 1959

By D. Kuhlmann-Wilsdorf

A new theory of work hardening is developed which rests on only a few simple principles and is applicable to a wide variety of materials and dislocation structures. It explains, qualitatively, the general characteristics of the three-stage shear stress-shear strain curve of fcc metals, as well as the most important features connected with easy glide. The value of, the work-hardening coefficient in stage 11, is derived quantitatively, from first principles, and is shown to be insensitive to many changes in detailed dislocation behavior. The low rate at which energy is stored during mechanical working in stage 11, the dependence of slip line length on stress, and the so-called Cottrell-Stokes law are explained. Although the theory is primarily developed for fcc metals and alloys, it is largely applicable also to some other materiuls, and in particular apparently to poly crystalline simple steels in theh linear range of hardening. FOR pure fcc metals and alloys, after many different pretreatments and under a great variety of testing conditions, the well-known three-stage work-hardening curve is commonly observed. Moreover, the work-hardening coefficient in stage 11, 011, bears an almost constant relationship to G, the modulus of rigidity, such that K = G/Brr" = 300, within a factor of about 2 either way. The very fact that the three-stage curve and the value of K are so very persistent, can hardly be understood except on the assumption that some quite simple phenomena are responsible. In this light all available theories of work hardening must be judged unsatisfactory: Each of them is based on some quite specific dislocation model, while it is known that metals with similar values of K may have widely different dislocation structures, for example, dislocation tangles in the case of pure fcc metals, and piled-up groups of dislocations in a-brasstype alloys. Thus, while a particular theory might conceivably represent a special case with a reasonable degree of accuracy, it cannot possibly have illuminated the underlying principle. Another severe criticism applying to the presently most widely discussed theories is that they employ empirical or even quite unexplained parameters in order to arrive at a numerical value of K. This indicates that some vital aspect of the work-hardening process has remained unexplained. In the present paper, a new theory of work-hardening is developed which does not suffer from the above shortcomings. It is derived from first principles, and is founded on the realization that the underlying causes for the three-stage work-harden- ing curve and for the persistence of the numerical value of O,, must be few and simple. As a first step it is shown herein that the experimental oklservations on "easy glide" can be explained on the assumption that in this stage dislocations multiply, beginning in a number of restricted area:;, and from there penetrate into crystal regions still substantially free of glide dislocations until a quasi-uniform dislocation distribution has been established. During "easy glide", the resistance to the movement of the foremost dislocations, advancing into still largely a dislocation-free regions, is believed to control the flow stress. This resistance is determined by various factors, among them the reaction of the dislocations against bowing-out into loops, which is due to their line tension. Without such bowing-out no dislocation multiplication would occur, but the corresponding contribution to the flow stress stays practically constant all through stage I. According to the present theory, stage TI begins as soon as there are no more areas left into which newly formed dislocations have not yet penetrated, i.e., as soon :is a quasi-uniform dislocation density has been attaned. Regardless of how, in detail, the dislocations should happen to be arranged at the end of "easy glide", whether in pileups or tangles, on one glide system or on several, the stress necessary to overcome the line tension of the segments must now increase, because the dislocation density increiises while the average free dislocation length decreases. It is argued that all other contributions to the flow stress stay at about the same level as in stage I, and that the said increase in the stress necessary to bow out dislocation segments into loops is responsible for the major part of work hardening in stage 11. Stage TII of the stress-strain curve is explained by largely follouing the ideas of Seeger and his co-workers; however, it is pointed out that the start of stage m may on occasion indicate the onset of "conservative climb" instead of cross slip. Indirect evidence indicates that linear work hardening persists even under the profuse action of cross slip, but at a lower rate. Only when, in addition to cross slip, climb operates extensively, does the work-hardening rat,. drop sharply, ultimately down to zero. BRIEF SURVEY OF EXPERIMENTAL EVIDENCE The stress-strain curve of crystals of many different substances are qualitatively similar. No significant permanent plastic deformation takes place below the so-called "critical resolved shear stress". As the applied stress is raised beyond this level. yielding occurs with little or no work hardening and

Jan 1, 1962

By William Gleason

The numbers tell the story of this one. More than 50,000 people and 1,890 exhibitors took over the Las Vegas Convention Center Sept. 24-26 for the largest mining trade show in the world ? MINExpo. Inside the convention center, some of the largest machines used in mining were on display. Caterpillar, Liebherr, Komatsu and Hitachi displayed ultraclass offroad mining trucks and shovels. Atlas Copco and Schramm were among companies that rolled in mining drills and Bridgestone, Goodyear and Michelin were among the companies showing off the massive tires used by the trucks. Throughout the three halls and an outdoor display area, just about every kind of product, gadget or service used in the mining industry could be found on display somewhere.

Jan 1, 2012

By R. H. Ring

The Automatic Extraction System (AES) was designed and built under a U.S. Department of Energy Contract by the National Mine Service Company in cooperation with the Lee Engineering Division of Consolidation Coal Company. The Pennsylvania State University, Mining Engineering Section, was contracted as an independent auditor of the AES performance. The AES had a rotating cutter drum like a continuous mining machine; but two hydraulically powered roof supports, four roof-bolting drills, a dust scrubber, a ventilation fan and ducting in the roof supports, semi-automatic control, an extensible chain conveyor, and other innovations led to the distinction "extraction system" as opposed to "miner-bolter" or "continuous mining" machine. After design and assembly, the AES was field tested experimentally in an underground coal mine. The in-mine performance audit was preceded by a computer simulation of the potential production for a later comparison with the actual tonnage mined by the AES. This paper describes the results of an in-mine trial and assesses the production potential of the AES.

Jan 1, 1980

By John B. Gustavson

The mineral deposits found in carbonate facies rocks of sedimentary basins include deposits comprised of the country rock itself such as limestone for chemical uses, dolomite for road aggregate, and a variety of other industrial uses of the sedimentary rock. Exploration for new reserves is generally conducted on basis of outcrop investigation, subsurface drilling, and facies change studies while geophysics and geochemistry are not applied. In contrast, rapid development has taken place in the application of both geophysical and geochemical techniques in exploration for stratiform deposits of zinc, lead, barite, and fluorite deposits. This paper will present case histories and show illustrations from such work in the Mid-Continent region. The Metals Division of International Oil & Gas, Inc. has gathered this experience through work over the last three years along the southern margin of the Eastern Interior platform focusing on the states of Arkansas and Tennessee. The exploration work was conducted partially for the Company's own account, partially in joint exploration ventures with mining companies including International Mogul Mines, Ltd., a Canadian company; Bear Creek Mining Company, a subsidiary of Kennecott Copper Corporation; the Mining Division of W. R. Grace and Company, and Boliden-Preussag Exploration, a Swedish-German partnership. We wish to take this opportunity to thank our venture partners for allowing presentation of the various sections of this paper.

Jan 1, 1976

The role of ZnSO4 in activation of sphalerite (ZnS) and deactivation of lead-activated sphalerite has been investigated. Adsorption of metallic ions at the solid-solution interface, electrokinetic results for sphalerite under different chemical conditions, and thermodynamic calculations were used to elucidate the mechanisms involved. It is shown that flotation response of sphalerite in absence of activators is rather weak and that adsorption of ethyl xanthate on the mineral mainly electrostatic in nature. Zinc sulfate is found to activate the mineral but only in a narrow acidic pH range. This is attributed to the increase of zeta potential as well as the possible increase in amount of negatively charged bulk precipitated zinc xanthate. Lead ions activate sphalerite in a wide pH range and the lead activated mineral assumes a flotation behavior similar to that of galena (PbS). However, addition of ZnSO4 depresses the lead-activated sphalerite by replacing lead ions at the mineral/solution interface.

Jan 1, 1999

By M. John Brodie

This report reviews the regulatory environment as it pertains to waste rock management and disposal from mining operations, excluding coal and placer gold operations, in the \irate of Alaska. It is based on the EPA's staff position on an effective program to regulate mining waste as presented in the May, 1990 Strawman II document. The unique conditions which occur in pans of the state, including a fragile environment, high seismicity and permafrost, are identified. The special waste management practices which are necessitated by these conditions are delineated. The report provides a detailed outline of the current technologies that are suitable for designing, constructing, operating and closing a waste rock facility to meet environmental protection objectives. Cost implications to the mining industry of potential regulatory actions, such as the degree of groundwater protection, are examined.

Jan 1, 1992

By Frank A. Peters

A series of evaluations is being made to determine the relative processing costs of producing metallurgical-grade alumina from low-grade aluminous materials. These evaluations include the estimated operating and capital costs as well as recommendations for additional research and development to improve the processes and to lower their costs. The design for each process is based on published work. The three processes considered in this evaluation are designed to leach alumina as aluminum sulfate from calcined clay with sulfuric acid and to remove the silica residue by filtration. The filtrate from leaching is then purified by one of the following methods: Removal of iron in electrolytic cells using mercury cathodes; removal of iron by chemical precipitation through the formation of an iron-manganese complex; or precipitation of the alumina from an aqueous-ethanol solution, leaving iron in solution. Both the operating and capital costs for these processes are higher than those for the Bayer process. For plants producing 1,000 tons of alumina a day, the estimated operating costs per ton of alumina are: Electrolytic iron-removal process........... $81.35 Chemical iron-.removal process............... 82.08 Ethanol purification process............... 86.17 The investment in mercury is not part of the operating cost since this cost is considered equivalent to catalyst cost which would not be shown. The estimated capital investments for 1,000- ton-per- day plants on a Marshall and Stevens chemical equipment index basis of 240.0 are: Electrolytic iron-removal process...... $116,775,500 Chemical iron-removal process............... 99,418,900 Ethanol purification process................ 94,210,200 These investments include the working capital but not the land for each of the processes. Included in the electrolytic iron-removal process is an investment of $4,859,400 for mercury. Investigation of other methods for removing iron, of the decomposition operation, and of the leaching operation could reduce these costs.

Jan 1, 1963

By Lewis Dr. Wade

It is really an honor and pleasure to be here at the 231d International Conference on Ground Control in Mining. My name is Lew Wade. Until recently I served as the Associate Director of Mining for NIOSH. That's a position that I recently moved out of and that position is now filled by Dr. Jeffery Kohler. I now serve as the Sr. Science Advisor to the NIOSH Director. The topic of my talk today will be coal mine ground control, a look back at the last 30 years, and then a brief look forward. I will start by talking about coal mine ground control statistics and with just a brief look, those statistics clearly show that you people are making a difference and there have been tremendous improvements. I'll then look at reasons for those improvements and those will include people working together (this conference being a prefect example of that), as well as technology development. In the area of technology development, I will give you my top-ten list of technology developments in the area of coal mine ground control. Then I'll end with a look towards the future and point to some possible future developments. Then, I will be so bold as to issue some challenges to the participants of this conference. Again, if you take the broad look back at underground coal mine roof falls and ground fatalities, you see this tremendous improvement that we have all come to appreciate. We have all been a part of improving the safety conditions for our underground miners and the number of fatalities is orders of magnitude less than what it was at the turn of the previous century. We have every reason to be pleased and proud of what we have been able to accomplish. Now, the question is why. What are the reasons for those improvements? Just look around this room. You see evidence of what I think are one of the principle reasons: miners, mine operators, government workers and researchers, equipment suppliers, people in research and academia working together. Coming together and sharing at a Conference like this over the last 23 years.

Jan 1, 2005

By H. A. Thorne

This paper describing the mining practices at the Chino mines at Santa Rita, Grant County, N. Mex., is one of a series being prepared by the Bureau of Mines on mining practices, methods, and costs in the various mining districts of the United States. The Nevada Consolidated Copper Co., Chino mine's, operates an open-cut copper mine at Santa Rita, Central mining district, Grant County, N. Mex. This mine is located about 16 miles east and slightly north of Silver City, the county seat. The ore is-transported in standard railroad cars by the Atchison, Topeka & Santa Fe Railway Co., approximately 10 miles to Hurley where it is milled. The concentrated are shipped approximately l'70 miles over the same railroad to the smelter of the American Smelting and Refining Co., at El Paso, Tex. (fig. 1)

Jan 1, 1931

At a special inaugural ceremony in Blue Hill, Maine on August 3rd, Governor John H. Reed detonated the first blast to signify the beginning of development work on a new copper-zinc mining venture that seems destined to focus increased interest in the mineral potential of that State. The property which is the first new base metal mine to be discovered in the northern Appalachians in 30 years, is owned by Black Hawk Mining Co., a subsidiary of Denison Mines Ltd. of Toronto. When it goes on-stream in 1966, it will join Ducktown, Tenn., as the only active copper producing region, east of the Mississippi River. According to official reports submitted by the Company to Robert G. Doyle, Maine's State Geologist, Black Hawk has found ore reserves totalling 4.5 million tons. Assays show 2.8 million tons of 2.0'; copper and 0.4 oz. of silver per ton; 1.2 million tons of 16.75% zinc and 0.58% copper; and 0.5 million tons of 6.75% lead with 4.25 oz. of silver per toil and some zinc.

Jan 9, 1964

By Rogis J

The basic parameters causing outbursts are rock pressure, gas characteristics and the physico-mechanical properties of the coal. The three sets of factors are to some degree interdependent. At Collinsville, the gas characteristics coupled with the physico-mechanical properties of the coal appear to be the major factors contributing to outbursts. The relative importance of rock pressure is uncertain. The following geological parameters appear to be necessary before outbursts can occur. 1. A coal rank greater than 1.2% vitrinite reflectance. 2. Thrust faults with throws greater than 3.Om or strike- slip faults with coal gouge wider than 0.3m. 3. Seam gas emission values greater than 1.1 cm3 /g. Igneous intrusions are the likely source of the carbon dioxide seam gas. Outbursts have occurred at depths ranging from 215m to 265m. They occur at these relatively shallow depths due to: 1. the high rank of coal at the present topographic surface and the high coalification gradient; 2. the relatively high degree of tectonic deformation of the strata; 3. the seam gas characteristics; and 4. the extent of igneous intrusions.

Jan 1, 1980

By W. S. Dyer

The Minto coal basin lies in the central part of the province of New Brunswick, near the head of Grand Lake. The village of Minto which is the centre of mining activity of the district, lies about 35 miles east of Fredericton and somewhat greater distances to the north and west of the cities of St. John and Moncton respectively. Although coal has been mined in the basin for a number of years, it was not until recently that the production became sufficient to warrant a thorough geological investigation. The summer of 1922 was spent, in accordance with in structions from the Director of the Geological Survey, in making a detailed study of the geology and coal resources of the Minto coal basin, which is covered by the Minto and Chipman map sheets, each embracing an area of 200 square miles. The present paper is a short outline of the principal results obtained in the field. The writer wishes to acknowledge the assistance cheerfully given him by the managers and officials of the various companies operating in Minto. He wishes especially to thank Mr. W. B. Evans, Manager of the Rothwell Coal Company, for assistance in the collection of fossils and for the generous supply of information relative to the early history of the coal field.

Jan 1, 1924

By Keith A. Heasley

In the past, numerous displacement-discontinuity (DD) models of stratified rock masses have been proposed using a homogeneous isotropic elastic overburden. In this paper, it is postulated that a laminated overburden model might be more accurate for describing the displacements and stress in these stratified deposits. In order to investigate the utility of using a laminated overburden in a DD model, the fundamental mathematical basis of the laminated model is presented and graphically compared with the fundamental behavior of homogeneous isotropic elastic overburden and with field data. Specifically, the stresses and displacements surrounding an idealized longwall panel as determined from the laminated overburden model are presented and compared with results from the homogeneous isotropic overburden and with measured abutment stress data. Also, the remote displacements and surface subsidence as calculated by the laminated overburden model are compared with homogeneous isotropic calculations and with measured subsidence data. The results of this investigation show that the laminated overburden is more supple, apt to propagate displacements and stress further, and better able to fit observed data than the classic homogeneous isotropic overburden. Ultimately, it is suggested that the laminated model has the potential to increase the accuracy of displacement and stress calculations for a variety of mining situations.

Jan 1, 1996

By E. M. Weenink

A proposed alternative to the conventional zinc extraction processes is the flash reduction of zinc calcines. As part of an overall study of this process, the kinetics of single zinc oxide pellets were investigated. Pellets, compacted from Outokumpu zinc calcine (60.9% zinc, 9.6% iron), were suspended in a flow of CO/CO, mixtures and reduced at temperatures between 1000 and 1200°C and atmospheric pressure. The pellet weight loss with time was recorded and analyzed. It was found that the reduction of calcine pellets (0.9 cm height and 1.36 cm diameter) was controlled by chemical reaction. This paper includes an analysis of the experimental results of two previous investigations of ZnO reduction.

Jan 1, 1992

By Sherwin F. Kelly

SINCE we used Chief Buehler's name last year to give our annual report a semblance of respectability, we can follow the good precedent thus established by telling of the work his Missouri State organization has been carrying on, as reported by H. S. McQueen. Considerable value is set upon their magrletorneter working which has covered over 13,000 sq. mi.), for the indications it gives of the major structural features of the region. Numerous geological problems have been tackled Ly resistivity techniques, among which are the locating and outlining of filled sink deposits of sulphides, of iron ore, and of fire clay; tracing of faults and solution channels; mapping of the preglacial structural pattern; measurement of depth of overburden; and

Jan 1, 1937

By Kymberley Worrell

The Cripple Creek and Victor Gold Mine was acquired by Newmont in August 2015. The high grade flotation mill project was commissioned in February 2014. The mill to date has been throughput limited due to mechanical and design issues. In order to create the most value from a throughput limited plant a campaign to provide timely ore characterization was undertaken. A diagnostic leach methodology was developed to determine gold associations by ore type. The resulting associations were then used to characterize feed blend; gold losses and optimize gold production. The diagnostic methodology is presented along with discussion on flotation implications.

By Jr. Langhans, J. S. McPartland

"The presence of zinc in cyanide solutions is generally thought to not have an adverse effect on precious metal recovery or recovery rate by heap leach processing. However, analytical interferences of zinc in solution can cause problems in accurately determining and controlling cyanide concentration in the leach solution, leading to increased cyanide consumption. Analytical methods and the chemistry of cyanide in the presence of zinc are reviewed. The effects of solution pH and zinc concentration on the free cyanide determination were evaluated in detail. Use of the standard silver nitrate titration and a modified titration method for controlling cyanide concentration during heap leaching were investigated experimentally. The practical implications of cyanide heap leaching in the presence of cyanide soluble zinc minerals are discussed.IntroductionThe presence of zinc in precious metal cyanidation circuits is not uncommon. Zinc can be introduced into cyanide leach solutions through dissolution of various zinc minerals1>, or through the use of a Merrill Crowe (zinc precipitation) circuit. Zinc minerals typically occur infrequently and in low concentration in gold ores, but are more abundant in silver ores. Merrill Crowe systems are most commonly employed for recovery of precious metals from solution for relatively high silver ores. It is desirable to control the amount of dissolved zinc in a cyanidation circuit for several reasons. Excessive zinc build-up in cyanide leach solution can, under certain circumstances, inhibit precious metal dissolution and adversely affect solution recovery systems. Additionally, high zinc concentrations can lead to excessive in-process cyanide inventories and potentially, to higher cyanide consumption"

Jan 1, 1999