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By E. P. Pfleider, D. H. Yardley

Layers of magnetic taconite of above average magnetic iron content extend down dip on the Mesabi Range to depths beyond the reach of open pit mining. A selected layer with thickness of 75 to 100 ft underlies many square miles. Such a selected layer can provide over 50 million tons of recoverable concentratc per square mile, at lower concentration ratios than in the present open pit taconite operations. The system contemplated would consist of large scale underground quarrying by a lane and pillar method. Working heights would be 60 to 100 ft; room widths, 50 to 75 ft; and productivity targets, about 40 tons per man-shift. The most critical problems would he roof control and blasthole drilling. The successful operation of large open pits and concentrating plants for treatment of magnetic taconites indicates that the basic techniques and economic problems of pellet production have been solved. The open pit reserves are large, but they are not necessarily available in sufficient quantity or quality to provide an adequate raw materials base for all potential producers. Furthermore, as the more easily available open pit crudes are used up, the pits will become deeper, stripping costs per ton of concentrate will increase, and recoveries will decrease as leaner crudes are utilized. The time will come when supplementary and replacement reserves of magnetic taconite will be required. One way these can be obtained is by underground mining of the down-dip extension of the Mesabi magnetic taconites, especially those layers averaging well over 25% recoverable iron. A key point in such a proposal is the possibility of realizing substantial savings by beneficiating an ore of 30% magnetic iron at 2.5 to 1 ratio of concentration. This contrasts with current practice in Mesabi commercial taconite plants, where the ore averages 22 to 25% magnetic Fe, with concentration ratios of about 3.0 when producing 61% Fe pellets. Two factors which favor the eventual development of underground taconite mining are: 1) the desirable physical structure and high grade of taconite pellets in general and 2) the improved engineering techniques in underground mining. Experience in large scale room and pillar mines demonstrates that, given strong rock strata immediately above a high grade mining

Jan 1, 1963

By David C. Byrne

PURPOSE: Evaluate practical technological advances in hearing protective devices for use in different mining environments. RESEARCH SUMMARY: Previous NIOSH research shows that miners develop hearing loss much more quickly than nonoccupationally noise-exposed individuals. Also, miners experience a greater severity of hearing loss than would be expected for nonoccupational noise-exposed persons of the same age and sex. As described in the National Occupational Research Agenda, problems created by occupational hearing loss include (1) the reduced quality of life due to social isolation and unrelenting tinnitus, (2) impaired communication with family members, the public, and coworkers,(3) a diminished ability to monitor the work environment (e.g., warning signals, equipment sounds, etc.), (4) lost productivity and increased accidents resulting from impaired communication and isolation, and (5) expenses for workers’ compensation and hearing aids. Although the elimination of hazardous noise is the ideal long-term solution to the noise problem, miners will likely need to rely on personal hearing protection in certain situations for quite some time. However, it is well known that noise attenuation provided by hearing protectors as they are worn during everyday use in the workplace is usually worse than when tested under laboratory conditions. Additionally, workers may refuse to wear their hearing protection for any number of reasons. Therefore, this project was initiated to evaluate practical technological advances in hearing protectors, with an emphasis toward optimizing the balance between effectiveness, comfort, overprotection, and communication ability. Currently, the necessary in-house laboratory facilities and research techniques are being finalized. Individual experiments are planned that involve analysis of new hearing protector technology under both laboratory and field conditions. The laboratory components are to consist of con-trolled studies using human subjects and/or an anthropometrically correct artificial test fixture to evaluate the adequacy/effectiveness of various hearing protective devices. Field studies are planned to determine the actual usefulness and practicality of the hearing protectors as used in diferent mining environments. Recommendations can then be offered to the mining community regarding correct earplug/earmuff usage that maximizes recognition of hazard/warning sounds while adequately protecting the miners’ hearing.

Jan 12, 2000

HE purpose of this chapter is to present a general outline of the basic open-hearth process for the benefit of students, practicing open-hearth operators, and metallurgists who wish to review the subject. The discussion is limited to a description of the sequence and purpose of the operations, illustrated by data and examples. Subsequent chapters deal with more detailed and fundamental studies of the subject. The function of the basic open-hearth furnace is to convert various types of ferrous material into finished steel of given composition and quality. In comparison with other steelmaking methods, the basic open-hearth process is one of the most versa- tile, both from the standpoint of the raw materials that can be used and the types of steel that can be produced. The process can be considered as a sequence involving melting, refining, and deoxidation. With some types of liquid-metal charges, the melting stage is not present, but refining and deoxidation in varying degrees are fundamental features of the process. GENERAL DISCUSSION OF BASIC OPEN-HEARTH PROCESS The first consideration in open-hearth practice is the type of charge on which the operation is based. This consists of three components-fluxing, oxidizing, and metallic materials. Fluxes. In the basic process the fluxing material is primarily lime, added in the form of either limestone (CaC03) or burnt lime (CaO). This is supplemented by magnesia (MgO) and lime from the furnace bottom and banks. The function of these fluxes is to form a slag with the silica (SiO2) , manganese oxide (MnO), iron oxides (FeO and Fe2O3), and phosphorus oxide (P2O5) that are produced by the oxidation of varying amounts of the silicon, manganese, iron, and phosphorus in the charge during melting and refining. The slag, composed of the basic fluxes and the neutral and acid oxides, floats on top of the steel bath when the charge is melted. The composition of this slag depends on the amount of each slag-forming element present in the charge and on the extent of oxidation of these elements during the course

Jan 1, 1951

By Noboru Masuko, Sachiko Ono

"The structure and composition of anodic films formed on aluminum in a borate solution have been investigated by using TEM and SIMS with respect to the effect of current density and electrolyte temperature. The film is composed of a boron-containing outer layer and a boron-free inner layer with disc-like crystals present at the interface. The thickness of the film formed at constant current density up to an identical voltage, decreases with increasing current density. The thickness increases remarkably with elevated temperature. The transport number of Al3+ estimated from the thickness ratio .of the cuter to the inner layers is 0.4 when the film is formed at the current density of 20 Am-2 at 60°C. It increases slightly with increasing current density and decreases significantly with increasing temperature. Content of boron in the outer layer increases 1.5 times when current density .increases 10 times. These changes in film structure and composition with electrolysis conditions must have a significant role in the determination of electric properties of anodic alumina films as a capacitor material.IntroductionThe growth of anodic alumina film in ·a neutral electrolyte proceeds at both film/electrolyte and film/substrate interfaces giving a duplex structure. The oxide layer formed at the former interface includes electrolyte anions while a-pure alumina layer is formed at the latter interface. The thickness ratios of each layer respectively represent transport numbers of Al3+ and 0 2 - (1). The film thickness and the incorporated anion content as well as transport numbers of both ions must affect the dielectric properties of the film as a capacitor material. However, the correlation between film structure and electrolytic conditions is not fully understood (2). For the further clarification of anodic film growth mechanism, more detailed characterization of anodic films in relation to electrolytic conditions is advantageous. In the present paper, therefore, the effect of current density and electrolyte temperature on the microstructure and composition of anodic barrier films formed in a neutral borate solution was studied. The films were characterized by direct observation using transmission electron microscopy (TEM) and depth profiles of constituent elements obtained by secondary ion mass spectroscopy (SIMS)."

Jan 1, 2000

By P. Mayer

"Determining the correct mechanism of a boiler tube failure is important for the prevention of future tube failures. Proper corrective measures can be undertaken to alleviate the root cause or causes for a failure only when the correct mechanism is known. A mechanism of boiler tube failure is defined as the process which degrades the tube and produces a failure.An expert system package was developed for use by electric utility power plant personnel during investigation into the cause of a boiler tube failure. It is intended to assist their efforts to recognize and identify a basic failure mechanism that can produce the type of boiler tube failure under investigation. The program guides the user to the most likely failure mechanism by asking simple questions, answers which can be obtained by determining the failure location, type of fuel, and from visual inspection of the failed tube. Expert systems are a form of artificial intelligence which imitate the decision-making process of experts.IntroductionBoiler tube failures (BTFs) continue to be the single largest source of forced outages in fossil fired electricity generating stations. Because of this a concentrated effort is being made to minimize boiler tube failure occurrences. Eliminating boiler tube failures could be worth $5 billion a year to the electric power industry(1).The aggressive service environment of fossil-fired boilers is largely the reason for these failures because the combined effects of stress, temperature, corrosion, erosion and vibration cause deterioration of boiler tube material properties and reduction in wall thickness.Determination of the boiler tube failure mechanism is important for the prevention of future problems. The root cause for a failure can be established and proper ameliorative measures taken only when the failure mechanism is identified correctly. The correct identification of failure mechanism is a complex process which may include many individuals and institutions. Technical specialists in metallurgy, chemistry, combustion, and boiler design frequently cooperate with the station staff in a failure investigation . These individuals rely on personal experience and numerous reports , theories, studies, private communications and investigations concerning the different failure modes and mechanisms."

Jan 1, 1990

By Ruixiang Wang, Jinhui Li, Youjun Xiao, Daoling Xiong, Hao Chen

"Leaching solution prepared from spent battery materials contains large quantities of valuable metals, such as Cu, Ni, Co, Mn and Fe. Therefore, recovery of valuable metals from acidic polymetallic aqueous solutions is of great importance for every plant. In this work, a method is introduced of recovery of Ni, Co and Mn from polymetallic solution prepared by leaching spent battery material. The basic concept includes copper was removed through replacement by iron powder followed by iron precipitation in goethite method. The experimental results show that Cu can be removed 99% at least through replacement by Fe powder, and the removal of Fe can achieve 99% by goethite method. At the same time, the loss of Ni, Co and Mn are about 2%, 3 % and 2%, respectively. The process of differential removal of metals can be selected as the treatment alternative that combines simplicity, efficiency and reliability with low capital and maintenance costs. The purified solution which only contains Ni, Co and Mn can be used to prepare NixCoyMnz ternary system precursor.1 IntroductionNickel and Cobalt are very important metals which are applied in many fields, such as energy material and support material for some new material field. The world demands for such metals is progressively increasing, while raw resources are being depleted and with the constraints raised by the Law and related legislation assigned to protect the environment from hazardous wastes in the world. Nowadays study on reutilization of waste materials has great significance for achieving secondary resources recycling and preventing environmental pollution. There are many hydrometallurgy methods applied for dealing with these wastes, while they usually contain large quantities of valuable metals, such as Cu, Co, Mn and Ni, as well as impurities, such as Fe, Zn, and etc. Therefore, proper treatment of these leaching solutions prior to their discharge is of great importance, so that the contained valuable metals can be recovered.Separation and extraction of metals from leaching solution can use some methods through solvent extraction[l,2], ion exchange[3,4], precipitation with sulfide[ 5] or alkali[ 6] and electrolytic recovery[?]. Among the appropriate treatment alternatives, replacement and differential precipitation is the most attractive treatment that combines simplicity, efficiency and reliability with low capital and maintenance costs."

Jan 1, 2011

By Miaoyong Zhu, Sen Luo

"A two-dimensional CAFD model was developed for simulating the solute diffusion-controlled dentritic growth in the low Peclet number regime. Mass conservation and local equilibrium at the solid/liquid interface were solved to determine the growth velocity of interface with consideration of constitutional and curvature undercooling. The modified decentered square growth proposed by Wang et al was adopted to capture the first eight neighboring cells. The comparison between the classic Lipton-Glicksman-Kurz (LGK) analytical model for steady state tip variables and the simulated values showed that the present model was capable of quantitatively determining the dendritic growth, and the simulation tests demonstrated that the effects of the mesh anisotropy in growth kinetics and crystallographic orientation were eliminated. The numerical simulations were performed to the dendritic growth ofFe-0.6%C binary alloy in undercooled melt, and the results showed that the present model could predict the dendrite morphology with artificial growth orientation in acceptable level.IntroductionDuring the solidification of metals and alloys, the dendrite grows from the substance or the undercooling melt and solute redistributes between the dendrite and the interdendritic liquid. The solidification microstructure is close related to the phase and solute evenness and has a close linkage with the properties of final production.In the past few decades, various kinds of deterministic and stochastic models have been developed to predict the evolution of microstructures in solidification, with the fast development of computer power. Among these methods, Phase Field (PF) model [1, 2] has been known as one of the most adequate deterministic models for directly simulating the dendritic growth. However, there is a vital drawback that the PF method requires considerable computation time and is limited to simulate very small domains with a few dendrites [3]. On the other hand, stochastic models, such as Monte Carlo (MC) method [4, 5] and Cellular Automaton (CA) method [6-8], have been usually applied to predict the solidification microstructure evolution. Comparing with the MC method, the dendritic growth kinetics is explicitly integrated in the CA model, while the MC model doesn't. Therefore, the CA method is more reasonable to predict the dendritic growth during the solidification of melt."

Jan 1, 2013

By Jeffery M. Listak, Ernest A. Curth

"It took only 9 years from the first appearance of roof shields on the U.S. longwall mining scene to the present predominance of shield faces. An apparent consequence is the welcome downward trend in accidents related to failures in longwall ground control. The report addresses load prediction, the effects of shield geometry on support loads, factors contributing to ground control, and related techniques. The Bureau of Mines took an active part in the evolution of longwall mining, including developing one of the first shield faces in 1975, the first lemniscatetype shields in 1976, the Mine Roof Simulator completed in 1980, the first shields in steeply pitching coalbeds in 1981, and the first multilift working of a thick coalbed in 1982. The information offered in this state-of-the-art report will assist in establishing criteria for roof support selection. INTRODUCTIONIn terms of roof support technology, the introduction of shields in the seventies and eighties was a major step forward. Table 1 indicates the predominance of shield faces on the U.S. longwall mining scene and its apparent consequence, a welcome downward trend of ground-control related accidents.The rapid progress of longwall technology began in 1975 and is highlighted by a development leading from the caliper shield to the modern lemniscate type. Shields have been gaining preference over chocks since their advent to longwall mining in 1975. Safety and productivity factors favoring shields over chocks are--1. A sheltered working space requiring minimum cleanup work.2. Structural stability that allows advancing without delay even with brushing roof contact and, by controlling all lateral loads, removing the requirement for cumbersome leg restraint and restoration devices.3. Wide hydraulic range of mining height.The development of roof shields in the Federal Republic of Germany preceded and paralleled their rapid adoption by U.S. miners. In 1983, 87 pct of the total production in Germany came from shield faces (1).2 In Great Britain, the National Coal Board introduced the first shield support system in 1977 to initiate Advanced Mining Technology (ATM) and Heavy Duty Mechanization (HDM) programs; as of mid-1982, 44 shield faces were operating, or 7.5 pct of a total of 581 faces CJ_)."

Jan 1, 1985

By Robert Linton

This paper deals with work that has been carried on for over three years by the management and staff of the North Butte Mining Co. in an effort to standardize mining methods, to eliminate lost motion in the operating organization, and to improve the efficiency of the individual workman. Naturally, much of this work is an adaptation of ideas suggested by other mining operations or by practice in manufacturing and other productive industries. The studies, however, have developed some features that it is believed are new. Both new and adapted features are given in detail. The normal capacity of the North Butte mines is about 2000 tons per day. Hoisting is done principally through the Granite Mountain shaft, which is 3740 ft. (1140 m.) deep and equipped with an electric hoist capable of hoisting 200 tons per hour from a depth of 4000 ft. (1219 m.). The lowest stoping level is the 3400-ft. The Granite Mountain shaft is connected with the Speculator shaft, about 850 ft. (259 m.) south, and with the Gem, about 1150 ft. (350 m.) east, which are 3000 (914.4 m.) and 2000 ft. (609.6 m.) deep, respectively. These two shafts are equipped with reversible fans for ventilation. The Speculator has a steam hoist with a capacity of 150 tons per hour from the 3000-ft. level. The Granite Mountain shaft is concrete lined from the 1600-ft. to below the 3000-ft. level and gunited with cement from the 1600-ft. level to the surface. Sections of the Speculator shaft with a total combined length of 500 ft. (152.4 m.) are also concreted. There are approximately 36 miles (58 km.) of drifts and crosscuts in the mine, in which there are 15.2 miles (24.5 km.) of track. At normal capacity, about 75 stoping faces, 20 raises, and 30 sill development faces are working. Stoping is usually by the standard square-set method. Where the size of the oreshoot and character of the walls permit, the rill (or cut-and-fill) system is employed. Scope and Limitations of Standardization Standardizing mining operations presents problems quite different from the problems of the manufacturer. Methods of a successful

Jan 1, 1922

By John J. Sammarco

Computer-related accidents have caused injuries and fatalities in numerous applications. Normal Accident Theory (NAT) explains that these accidents are inevitable because of system complexity. Complex systems, such as computer-based systems, are highly interconnected, highly interactive, and tightly coupled. We do not have a scientific methodology to identify and quantify these complexities; specifically, NAT has not been operationalized for computer-based systems. Our research addressed this by operationalizing NAT for the system requirements of safety-related computer systems. It was theorized that there are two types of system complexity: external and internal. External complexity was characterized by three variables: system predictability, observability, and usability -the dependent variables; internal complexity was characterized by modeling system requirements with Software Cost Reduction dependency graphs, then quantifying model attributes using 15 graphtheoretical metrics the independent variables. Dependent variable data were obtained by having 32 subjects run simulations of our research test vehicle: the light control system (LCS). The LCS simulation tests used a cross-over design. Subject perceptions of these simulations were obtained by using a questionnaire. Canonical correlation analysis and structure correlations were used to test hypotheses 1 to 3 - the dependent variables predictability, observability, and usability do not correlate with the NAT complexity metrics. Five of 15 metrics proposed for NAT complexity correlated with the dependent data. These 5 metrics had structure correlations exceeding 0.25, standard errors < 0.10, and a 95% confidence interval. Therefore, the null hypotheses were rejected. A Wilcoxon signed ranks test was used to test hypotheses 4 to 6 - increasing NAT complexity increases system predictability, observability, and usability. The results showed that the dependent variables decreased as complexity increased. Therefore, null hypotheses 4 to 6 were rejected. Lastly, this work is a step forward to operationalize NAT for safety-related computer systems; however, limitations exist. Opportunities addressing these limitations and advancing NAT were identified. Lastly, the major contribution of this work is fundamental to scientific research - to gain knowledge through the discovery of relationship between the variables of interest. Specifically, NAT has been advanced by defining and quantifying complexity measures, and showing their inverse relationship to system predictability, observability, and usability.

By H. S. Bell

It seems unnecessary to dwell upon the advantages of the modern pipe still as compared with the older type of distillation equipment used by oil refiners. The relatively low installation cost, coupled with remark-able operating efficiencies of pipe stills, precludes the consideration of other types except in rare instances where some peculiarity of product or plant may have an influence. The question of the actual heat required to distil petroleum is quite involved. We have extensive data on the latent heats of the various hydrocarbons under different conditions of pressure. We also have information regarding the specific heats of various oils and the effect of temperature thereon. Because of the complexity of petroleum and of each fraction, it seems impossible to analyze theoretically the partisl-pressure effect of the great number of hydrocarbons upon one another during distillation. If a distillation curve and percentage analysis of any crude or product are available, it is possible to construct a curve with ordinates representing per cent. distilled and B. t. u. per pound or per gallon, or other desired unit. To do so the crude is split into fractions, of 5 or 10 per cent. The boiling point and gravity of each fraction being determined, it is possible to evaluate the heat required. As each fraction is removed, the residue must be raised to a higher temperature and latent heat supplied for the successive fraction. As the calculations progress, the latent heats decrease while the specific heats increase. Such a curve is illustrated by Fig. 1 for a typical 39.9° Be. North Texas crude. The chart is illustrative of batch distillation. Theoretically, if the oil were heated by a single flash in a pipe still, the partial-pressure effect of the light products would result in the removal of the same amount of vapors at a lower temperature. Practically, we find that because of the necessarily extensive fractionating equipment required when separating many products all in vapor phase, the oil has to enter the separator at temperatures corresponding to or even above those required in the batch still. Furthermore, in single-flash distillation the vapors evolved in the heater must be superheated to the outlet temperature. Very little is

Jan 1, 1928

By Paul K. Sims

The Central City district is one of several adjacent mining areas in the Front Range, Colorado, that contain precious metal-bearing sulfide veins of Laramide (Late Cretaceous and early Tertiary) age. The districts are in the Front Range segment of the Colorado mineral belt, and are characterized by porphyritic igneous rocks of intermediate composition that have intruded Precambrian crystalline rocks in the core of the Front Range. The veins and minor stockworks that constitute the ore deposits in the district were formed largely as fracture fillings. The veins range from simple fissures to complex, branching lodes, and commonly range from less than 0.5 to about 1 m in width. The principal values are in gold and silver. The common metallic minerals are pyrite, sphalerite, chalcopyrite, tennantite, and galena; minor minerals are enargite, pitchblende, and gold tellurides. Quartz, cryptocrystalline silica, carbonates, fluorite, and barite compose the gangue. The ore deposits have a well-defined concentric zonal arrangement. A core of pyrite veins (central zone) is surrounded by a peripheral zone of galena-sphalerite vein with gold and silver. An intermediate zone contains transitional veins that have minerals characteristic of both the central and peripheral zones. The minerals were deposited in three stages during a single period of mineralization: uranium; pyrite; base metals, gold and silver. The wall rocks were altered primarily during pyrite-stage mineralization. Sericite envelopes developed adjacent to fractures, and were succeeded outward by zones of argillized rock. The valuable ore minerals were deposited in ore shoots localized by several types of structural controls, principally vein intersections, the physical nature and structure of the wall rocks, and irregularities along the fractures themselves. During supergene alteration, gold was concentrated in the upper, oxidized parts of the veins, commonly to depths of 50 m. The precious metal-bearing sulfide ore deposits are closely associated in space and time with the Laramide porphyritic igneous rocks and are believed to have been derived from hydrothermal fluids related to the magmatic activity. The ore deposits have mesothermal aspects and apparently formed in the temperature range 200°-320°C.

Jan 1, 2013

By J. O. Brittain, E. P. Lautenschlager, F. Felberbauer

This investigation was undertaken to evaluate the effect of a dispersed phase in a iron upon the strain hardening and the stress dependency of dislocation velocity as inferred from the strain-rate sensitivitv. Experiments were made in a-iron specimens with controlled amounts of carbon and/or A12O3. They were tested at various temperatures between 195&apos; and 373 &apos;K under a variety of conditions oj-concentration, strain rate, and quench aging. The resulting yield strength, flow strength, and strain hardening increased with addition of Al2O3. The parameter m, measured from changes in strain rate, was xsed to describe the stress dependency of the dislocation velocity. m was found to increase with additions of A12O3, increasing carbon in solution, and with increasing temperature. When applied to a recent model for yielding based upon dislocation multiplication and velocity characteristics, the values of m alone did not successfully predict yielding for the materials of this investigalion, but had to be adjusted with a consideration of the number of unlocked dislocations nucleated heterogeneously at discontinuities or inclusions. The change of yield strength with Al2O3 Interpar-ticle spacing appeared to obey a theory of Orowan. EVER since the concept of dislocations has been introduced to explain plastic deformation in crystalline materials, the origin of dislocations and the source of the number of dislocations required to account for plastic deformation have been persistent questions. Ruling out the possibility of homogeneous nucleation,&apos; Gilman2 drew attention to the various cases of heterogeneous nucleation as the source of dislocations, whereas Frank and Read3 introduced the Frank-Read source. However, Gil-man found that the Frank-Read sources did not play a dominant role in the deformation of LiF crystals, but that small precipitates are often, if not always, associated with the nucleation of dislocations. In LiF crystals, however, most of the dislocations arose through the multiplication of moving dislocations in a mechanism described by Koehler4 and Orowan,5 Thus, relatively few precipitates could account for a large number of dislocations. A similar observation of dislocation nucleation at inclusions of precipitates was made by Stein and LOW&apos; on Si-Fe crystals. Their crystals contained

Jan 1, 1964

By J. E. Hawley

"The property of Francoeur Gold Mines, Limited, lies directly west of the Armfield mine in Beauchastel township, Abitibi district, and 12 miles west and slightly south of Noranda. The gold deposits are of particular interest because of their replacement character and close association with fault zones. The property was originally staked in 1923 following the discovery oi a mineralized zone near the eastern boundary. This became No. 1 orebody. Later discoveries in 1936 of No. 2 and No. 3 orebodies led to underground development and actual mining in July, 1938. A fourth orebody, No. 8, was found in 1939.The following is the production record up to December 31st, 1947:Beginning in September, 1944, ore was shipped to Noranda as siliceous flux, but late in 1946 milling of low-silica ore from No. 1 orebody was undertaken until operations were temporarily suspended in 1947.The four main ore deposits, Nos. 1, 2, 3, and 8, consist of siliceous ferrocarbonate and albite replacement bodies mineralized with fine pyrite and specularite in schistose fault zones in Keewatin volcanics. Barren to low grade quartz veins though present in some are uncommon, but small stringers of quartz are usually present wherever here and there visible gold is found. Nos. 1, 2, and 3 deposits lie in a major fault zone that extends completely across the property from west to east (over 7,000 feet) and dips N.40° to 50°, and are separated by intervals of 1,800 to 2,000 feet respectively. The fourth lies in a smaller fault zone striking east and dipping on the average 65°S.No. 1 orebody has been explored by mining to a depth of 500 feet, No. 2 to 400 feet, and No. 3 to slightly less than 300 feet . The greater part of No. 8 lies above a depth of 200 feet, but a raise has recently .been driven through from the west end of the 5th level of No. 1 to the 2nd level on No. 8 zone. Most of the ore of No. 8 zone above the 2nd level and of Nos. 2 and 3 above the 3rd levels has been removed."

Jan 1, 1954

By J. D. Zhu, T. Ohmichi, H. Yasuda, S. Kashiwai, I. Ohnaka

"Since the casting process affects the properties of castings, its computer simulation has been greatly developed in the last decades. This paper presents our recent works, especially on mold filling simulation, because most casting defects such as porosities and inclusions that greatly affect the properties originate during mold filling. The first topic is the mold filling simulation with consideration of cavity pressure change during mold filling, including direct observation of flow phenomena with X-ray and pressure measurement. The second .one is the mold filling simulation in the lost foam process, in which gasification of polystyrene pattern and gas outflow through the coating and mold have been considered. The simulation can predict not only the mold filling sequences but also periodical vibration of the gas layer between the melt and pattern. The third one is the comparison of simulation of Al melt flow in a metallic runner and direct observation through a transparent mold, showing that the entrapped gas or inclusions is expelled ahead through the melt and its simulation needs very fine meshes. These examples show that the comparison of simulation and observation is very important to develop the numerical simulation further.IntroductionThe mechanical properties of castings are greatly affected by casting defects such as porosities and inclusions, and solidification structure. Although recently the structure simulation has greatly progressed [1-4 ], the application in plants is very limited, because the structure is mostly determined by chemical compositions and cooling rate, enabling us to estimate it from the structure of past similar castings without simulation.On the other hand, the casting defects often vary with each casting or time. Among the defects, porosities and inclusions are main causes to deteriorate the properties and mostly originate during mold filling. Therefore, a lot of efforts to simulate the mold filling and estimate those defects have been made in the last decades [l,5,6]. These efforts have been mostly to develop numerical techniques to solve the Navier-Stokes equation. However, the flow phenomena in molds have not been well understood, because the direct observation of the flow phenomena in mold is not easy. This means the accuracy of the simulation is still not clear."

Jan 1, 2000

By Bryon W. Cheever

The greatest objection to be brought against the present methods for estimating phosphorus in iron and steel, is the time consumed in the operation. The following method, originated and perfected by Mr. E. D. Campbell, assistant in the chemical laboratory of the University of Michigan, will, 1 think, if carefully followed out, give results that are sufficiently accurate for technical purposes, and will give them in less time than they can be obtained by any other process. The process consists in comparing the color obtained by the action of stannous chloride upon a solution of the phospho-molybdate precipitate obtained from the solution of iron or steel, with that obtained from a standard solution of molybdie acid which represents a given amount of phosphorus. The operation is as follows : Take of ordinary steel 1 grm. and treat as in any of the wellknown processes until the phosphorus is precipitated as ammoniumphospho-molybdate. When precipitated, the phospho-molybdate precipitate is filtered out, thoroughly washed with an acid solution of ammonium nitrate, and, finally, twice with water, to remove most of the ammonium nitrate. The precipitate is then dissolved in about 1 c. c. of 10 per cent. potassium hydrate, which is let fall drop by drop on the filter till the precipitate is dissolved ; the filter is now washed with about 25 or 30 c. c. of water, and the potassium hydrate solution, which now consists of 25-32 c. c. of solution, is boiled until the peculiar frothy appearance, which first appears on boiling, is entirely gone. The solution is then cooled until perfectly cold, and about 10 or 15 c. c. of strong hydrochloric acid added ; if the solution is kept cold this will form no precipitate, but a light yellow or colorless solution. The solution is now ready for the addition of stannous chloride, and the subsequent comparison of color produced with that obtained from the standard molybdie acid. The comparison is made in tubes that are calibrated like carbon-tubes, but which are larger, about 5/8-inch bore, and graduated to 100 c. c.; this makes them about 18 inches long. To make the comparison of color, first transfer the now acid solu

Jan 1, 1886

By M. W. Hayward

THE object of this paper is to record and tabulate the data and field observations obtained by the writers and their associates during 10 years of intensive study of lead-zinc deposits in the Cretaceous limestone areas in Northern Mexico, particularly in mines owned by the Cia Minera de Peñoles, S. A. At the El Paso meeting of the Institute in October, 1930, J. G. Barry made the suggestion that a committee be appointed to correlate the ideas and theories which have been put forth by the numerous engineers and geologists who, during recent years, have made a study of the ore deposits in the area under consideration. The first step in a work of this nature would be to record all of the observed facts, and these facts would have to be accepted by those who attempted to formulate a theory of ore deposition. It is with this idea in mind that the writers have described as accurately as possible the actual, observed facts pertaining to the occurrence of the lead-zinc ores in dolomites or dolomitic horizons. By most writers the term "dolomite" is loosely used to designate magnesium-bearing limestones, regardless of the percentage of MgO present. Since no true dolomites exist in the areas to be described, we shall also loosely use the terms dolomite or dolomitic limestone to desig-nate those beds having a magnesia content of 5 per cent MgO or more. For a number of years the writers have been impressed with the fact that in areas where dolomite beds occur a large percentage of the lead ore is found in these beds, in spite of the fact that such beds may be overlain or underlain by pure limestone, which one would assume to be just as favorable for replacement as the dolomite, if not more so on account of its being more soluble in acid solutions than the dolomite. It is true that in some limestone areas, such as the Erupcion-Ahumada districts, Chihuahua, and Providencia, Zacatecas, where no dolomite beds have been found, large tonnages of lead ores have been mined. We venture to predict, however, that had there been dolomitic beds in these areas, the

Jan 1, 1932

By Frank D. Shepherd

THE Gunnar Gold Mines property is on the west shore of Beresford lake, in the Rice Lake mining division, southeastern Manitoba. Gold was first discovered in place in the Rice-Beresford Lakes area in 1911 on the Gabrielle claim, now included in the San Antonio Mines property at Rice lake. Following an eastward trend of favourable prospecting ground, promising showings were next staked in the vicinity of Long lake during the years 1915 and 1916. At that time, William Walton and associates, among them Gilbert LaBine, staked the Eldorado vein, north of Long lake, and Eldorado Gold Mines, Limited, was formed to develop the property a Company that, later, was to become widely known through successful operation of its radium-silver mine at Great Bear lake, Northwest Territories. Continuing eastward from Long lake, Walton and his partners, Gunnar Berg, Leo Seaberg, and Wm. Quesnel, staked claims during the years 1920-24 in the vicinity of Beresford lake which, later, were to become incorporated into the Gunnar Gold Mines property. The presence of gold-bearing quartz on these claims had been known from the time they were staked, but an opportunity to consolidate them into a group was not forth-coming until early in the year 1933. Although promising gold showings had been disclosed during the performance of assessment work on individual claims, no continuous mineralization having mine-making possibilities was discovered until after the consolidation of the entire group. From this stage onward, progress was rapid. Active exploration was begun on August 1st, 1933, and late in October James Houston reported favourably on the property. On October 27th, Gunnar Gold Mines, Limited, was incorporated in the Province of Ontario to develop and operate the property. The Company is also registered in Manitoba. The authorized capitalization is 3,000,000 shares of $1.00 par value. Gilbert LaBine was appointed president of the Company, with James Houston as mine manager. Following completion of a diamond-drilling programme, planned in 1933, it was decided to install a mine plant capable of carrying development to a depth of 1,000 feet. A transmission line, 5 miles in length, was built to the Great Falls-Central Manitoba line of the Manitoba Power Company, Limited.

Jan 1, 1939

By Yongjie Bu, Wei Sun, Zhiyong Gao, Yuehua Hu, Hongbiao Tao, Runqing Liu, Central South University

"The flotation behavior of O-isopropyl-N-ethyl thionocarbamate (IPETC) and pH for chalcopyrite and galena was studied through pure mineral test, adsorption quantity determination, solution chemistry calculation and Materials Studio simulation. The results show that the recovery differential value of chalcopyrite and galena is about 20% when pH of the pulp 9.5 and IPETC is 7×10-4 mol/L, the flotability differences of the two minerals is significant. There is competitive adsorption between OH? and IPETC on mineral surfaces, it lead to galena flotability is lower than chalcopyrite flotability. IPETC can remove the hydration layer on mineral surface and adsorb on active sites for collector adsorption, minerals have a better hydrophobicity.INTRODUCTIONCopper and lead are two of the most widely used metals in industrial and living, their applications include building construction, various metal alloys, heat and electricity, machine manufacturing (Zhang et al., 2014). Chalcopyrite and galena are the major source of copper and lead, normally they associated with each other. With the growing use of these resources, high grade and dissemination large particle sizes minerals are less and less, the use of low grade and complex dissemination minerals have to be paid more attention (Chen, Ye, & Li, 2010; Liu, Sun, Hu, & Wang, 2009). As the special properties and complex composition of copper-lead sulfide ore, flotation separation of chalcopyrite from galena effective is a research emphasis and difficulty point in mineral processing (Bruckard, Sparrow, & Woodcock, 2011).In the separation process of chalcopyrite and galena, it is rather important to choose effective collector (Huang, Wang, & Liu, 2014). Enlarging and using the two mineral flotability will be a good basis and start in flotation (Reyes-Bozo et al., 2014). Qin et al. (2013) made a great deal of research on flotation separation of copper/lead concentrate, collector O-isopropyl-N-ethyl thionocarbamate (IPETC), butyl xanthate (BX) and ammonium dibutyl dithiophosphate (ADTP) were investigated, BX and ADTP floated both chalcopyrite and galena well, specific adsorption appeared in the surface of mineral, the problem was the collectors were not selective enough, however IPETC showed better selective."

Jan 1, 2016

By Perry L. Christian

"Mana/ta Coal Ltd. 's Costello Mine, located seven kilometres east of Estevan, Saskatchewan, has been operating at the present location since 1960. During the past twenty-three years, annual production of lignite coal has averaged 590 000 tonnes. The coal is used for power generation and for industrial and domestic use. Markets, operating techniques, facilities in place and reclamation are discussed. Introduction The Costello Mine (formerly the Klimax Mine) is located in southeastern Saskatchewan, approximately 210 kilometres southeast of Regina and 5 kilometres from the United States' border, as shown in Figure 1. The present mine, 7 kilometres east of the city of Estevan, has been operating since 1960. The history, however, of the Costello Mine and of the Estevan Coalfield dates from the initial settling of the area in the 1880s. Mining initially consisted of burrowing into the side of the Souris River Valley or other hillsides where the coal seam was exposed. Eventually, slopes were put down in the Estevan/Bienfait area and vigorous underground activity existed. In the late 1920s surface strip mining was introduced. The use of strip mining progressed to the point where it completely replaced underground mining in the 1950s and it is the only method in use today. During the past 23 years annual production has averaged 590 000 tonnes. The lowest production was in 1960 when the mine was starting up at 167 000 tonnes and the highest production was in 1977 when production was 1 509 000 tonnes. The Costello Mine controls approximately 11 400 hectares of mineral rights in the Estevan area. This is made up of 4 900 hectares of freehold coal and 6 500 hectares of leases. The area forms a large, irregular shaped configuration. Most holdings are contiguous and adjacent. The mine also owns some 8 400 hectares of surface rights. The lands under the Costello Mine's surface rights control, which are not being used in the mining operation, are leased to local farmers and ranchers. This ensures that the land is used to its full potential, both before and after mining."

Jan 1, 1985