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By Michael S. Wieland

"Toxic fumes cause fatal andnonfatal incidents in underground mining, where the working environment tends to trap the fumes, hindering the restoration of non-harmful conditions. Workers can underestimate the residual fume toxicity and return to the work site too quickly. The toxic fumes from charge explosions depend upon a number things: formulation ingredients, mixture uniformity, water resistance, hole contaminates, rock hardness, and dust interactions. Transitory hole-to-hole waves can cause partial desensitization, ruin the reaction kinetics, and worsen the fume toxicity.Traditionally the toxic fume hazard for a candidate explosive is resolved from a requisite tally of fume components. The resolution starts with the fume spectrum, a tabulation of specie concentrations, transformed to standard reference conditions and readjusted to remove the air dilution. The resultant influence, which is referred to as relative fume toxicity RFT, is deduced by a formula rule with some chosen constants and the requisite concentrations. The RFT result is normally compared to an RFT criterion that represents the worst-case (maximum) tolerable fume toxicity, as stipulated by regulations or otherwise."

Jan 1, 1997

By G. E. Johnson

E. D. HYMAN*—How much sorting of scrap is done ? G. E. JOHNSON (author's reply)—We do practically no sorting. We charge "run of mine" scrap to the furnace. The unmeltables, mostly iron, are in such demand today that there is no difficulty in disposing of them. It may soon be desirable to sort out from the unmeltables as much of the brass as possible. J. J. BRUGMAN†—We have somewhat similar problems in the secondary aluminum business. What is your method of removing the unmelted material from the furnaces? Why have you such an apparently small space in which to charge your materials? Do you find that you have to seal that opening, or can you have it open and continuously charge at one end and pull out the other ? G. E. JOHNSON—Our means of melting scrap is efficient only to the extent that we use the waste heat from the vaporizing chamber to do the job. It is a batch process. We open the charge door and place the scrap on the hearth by hand shoveling. The door is then closed during the melting down period. After the melting is complete, the opposite door is opened and the unmeltables are raked out. The doors are approximately 3½ X 4½ ft and are not sealed during the process only closed. They are nominally tight. Some metal is oxidized in the process. We have visualized a means of conveying the materials through this melting unit with the metals that are melted trickling out during its travel. T. H. WELDON‡—Mr. Johnson, in line with the last question, is it necessary to seal the furnace between the melting down and the vaporizing unit, or have you got an inverted syphon in the bottom of the chamber? G. E. JOHNSON—That was one of the first things we encountered. We had to have a sealed opening, and it is a molten metal seal. You have indirectly asked me another question, which was: "Do you have to seal up the melting unit?" I would say we should exclude as much air as possible, although we are not too efficient in doing that. We allow the melting unit doors to be open when we charge and when we remove unmeltables. You can readily see that that would lead to the idea of having a controlled atmosphere in the melting unit, and I think this would do a more efficient job of melting the scrap. T. H. WELDON—How often do you charge the furnace? G. E. JOHNSON—We charge the melting unit, and rake out the unmeltables, about every hour. H. R. HANLEY*—Are any provisions made for controlling the rate of oxidation for the production of various size particles for certain characteristics of the zinc oxide product? G. E. JOHNSON—Yes, there are many. You are getting pretty much into the fine points of zinc oxide manufacture. Some of us still think we have something to learn about that. In general, this muffle furnace as I have described it to you produces a rounded particle of zinc oxide which is generally formed by a rapid oxidation of the zinc vapor, followed by rapid cooling. We have gone to the other extreme in some of our experiments. We have changed the furnace to produce a type of zinc oxide, such as we thought was peculiar to American process zinc oxide, by controlling the temperature at the point of oxidation and maintaining that temperature for a much longer period of time than we do when we make the rounded shape. There are other relationships that this furnace readily provides. One of the important factors is the ratio of air to zinc vapors. We can vary that by varying the air supply to the baghouse, or vary the rate at which we are vaporizing the zinc by the simple expedient of regulating the temperature over the carborundum arch. We have a number of variables that permit us to produce all of the grades of French process zinc oxide from lead-free up through the highest grades of seal oxides. There are many controls that we can apply to the operation. What I have said is but a brief condensation. K. MORGAN*—Can Mr. Johnson give us some idea of the fuel consumption of the furnace ? How much oil does he use per ton of zinc distilled? I am also interested to know what sort of heat transmission he gets through the arch? What is the thickness of the tiles used to construct the arch? Some time ago we built a small furnace for a different purpose, using a carborundum arch, and we found that the reflectivity of the molten zinc surface was so great we had to use a very high arch temperature. We found we made an improvement by having a layer of carbon on the surface of the zinc. Has Mr. Johnson had any experience on these points ? Does he make any sort of insolu-bles which he leaves in the furnace which he cannot tap out ? G. E. JOHNSON—I believe your first question was the fuel consumption. If I recall, somewhere in this paper there is a test that I quote. I believe we used 800 gal in a given period of time. Offhand I cannot translate that into tons of metal. I might also state that we have this understanding; that the carborundum arch, as the temperature becomes higher, becomes more efficient in heat transmission. As a matter of fact, I believe it is at about 2600°F or higher that the highest efficiency of heat transmission becomes available. We have calculations that I cannot quote from memory which indicate that the carborundum arch does a really very fine job for this type of furnace. Another point that we had considered was the fact that this furnace could be readily constructed so as to furnish an ideal source of heat for waste heat boilers. If a carborundum arch was used over the melting unit, we would have no zinc vapors in the gases at all— just clean combustion gases from which we would have removed some of the heat. L. P. DAVIDSON†— The insolubles

Jan 1, 1950

By James G. Tilton

Subsidence resulting from mining operations, has been the object of sporadic study for many years. Reports have shown that subsidence has ranged from minor settlement detected only by precise measurements to great slips where buildings are destroyed, roads disrupted, and large fissures opened at the surface. For years, a controversy has raged over the question of whether there is a positive or negative angle-of-draw. None of these questions can be given a flat answer for the variables controlling, subsidence are infinite. Generalizations can be made for various situations and areas but some means of control must be executed.

Jan 1, 1966

By M. F. Dibble

The Fosfago phosphate rock concentrator is located in the state of Goias, Brazil, 300 km south of Brasilia. The ore is from an enriched carbonatite, which was discovered in 1971 during exploration for pyrochlore niobium ore. The flowsheet was developed in Brazil by personnel from Agrico Chemical Company and Consolidated South America. Engineering was provided by Codemin, a subsidiary of Consolidated, Engineering and construction management was pro- vided by both Agrico and Consolidated personnel who formed the Fosfago team. The concentrator was designed for 500,000 tpy of 37-38% P2Oj dried product. The facilities consist of open pit mining, crushing, stockpiling, and concentration. The product is pipelined seven km to a filtering, flash drying, and railroad loadout station. The plant was successfully started in May 1979.

Jan 1, 1980

With the proving of a mineable orebody at Orlando in May 1961, Peko Mines N.L. was faced with the decision of whether to mill the ore on site or to haul it 25 miles to central facilities at Peko. Orlando was a relatively small orebody, so that it was decided to haul, although no reliable assessment of possible ore haulage costs was available.The necessity for a good haul road was recognised from the beginning and a 17 mile, 12 ft wide, single seal bitumen road was constructed, to link the mine to the Stuart Highway and Peko. The initial selection of 20 ton capacity diesel-powered end-tipping ore carts was made, with the best information available at the time.Since then some 130,000 tons of ore have been carted over a period of 3 years, and this paper presents the experiences of this haulage operation and the lessons learned.A second mine has just been brought into operation at Ivanhoe, 19 miles from Peko. For this haul it was decided to use an ore haulage vehicle of the side tipping road train type, which would cart some 70 tons, or more than three times the load of the vehicles initially selected.Only about 5,000 tons had been hauled at the date of this paper, but preliminary figures indicate a cost 01 1· 65d./ton-mile as against 2·47d. for the smaller vehicles.It would appear that the road haulage of ore from small outlying mines to a central plant can be highly economical, if certain basic rules concerning haul roads and vehicle selection are followed. This paper does not try to determine the maximum tonnages and haul distances which would finally make road hauling uneconomical...

Jan 1, 1966

By N Ross, N Goldup

In recent years the heavy mineral mining industry has started to look more closely at the feasibility of mining moderate to high slimes deposits. During the initial stages of orebody evaluation the primary purpose of the pilot plant is to produce heavy mineral concentrate for metallurgical evaluation. Hence, generally little consideration is given to the disposal of the waste material (slimes and tailings). This paper suggests additional, simple, low-cost, tests that can easily be incorporated into an initial pilot plant operation. With this data a more definitive evaluation of available waste disposal options can be made during the early stages of the project. Thus, permitting better project focus and utilisation of resources. These tests include free settling of slimes, natural slimes drying and consolidation (undrained and drained), tailings stacking angles, and particle size distributions of input and waste streams, etc. The objective and method of each test is summarised, together with its relevance to possible waste disposal options. The results of this preliminary testwork can later be used to expand upon and define further tests for the second phase of pilot plant operations, which will enable a more definitive evaluation of a proposed waste disposal solution.

Jan 1, 2001

By P. C. Kapur

The hierarchical relationship between three broad kinds of grinding models for a spatially homogeneous, linear, time-invariant batch grinding system are first established. Limiting and asymptotic arguments for arriving at approximate but simpler models are employed in all three cases. These parsimonious models -provide alternate and, in some respects, more incisive and convenient tools for the analyses of grinding processes and simulation of comminution circuits. In the discrete-size, discrete- time class of models, a reduced mill matrix is derived, which is seemingly adequate for those grinding circuits whose identification otherwise is difficult because of limitations on the quantity of data and its quality. In discrete-size, continuous-time types of models, parameter lumping has been employed for obtaining a G-H solution which is quite accurate and exceptionally useful for graphical display of data, model verification, and parameter estimation and simulation. In continuous-size, continuous-time models, a similarity transformation is used for solving the integrodifferential equation of grinding. This solution provides a unified framework for a number of apparently diverse empirical grinding phenomena such as self-preserving behavior of ground particle size distributions and energy laws.

Jan 1, 1987

The objective of this study was two-fold: 1. To perform a world literature search pertinent to the analysis, selection and design of hoisting systems and their components and to codify the search into a literature compendium. The search resulted in a technical data file of over 1500 hard copy items. The compendium index was also suitable for establishing a comprehensive computer-loaded data base on the subject. 2. To develop a comprehensive system of guidelines to cover salient aspects of analysis, selection and design of electrical, mechanical and structural components of a hoisting system with a view to establish parameters for maximum safety and efficiency of operation, installation and revision of the system. 3. (a) In-shaft equipment design and specification criteria are considered as they influence guideline preparation. (b) A comprehensive review of hoist electric drives and controls as applied to mine hoists is set forth with special reference to duty cycle calculations and their effect on electrical component selection.

Jan 1, 1988

By William Gleason

While a fleet of 100 t (110 st) haul trucks were busy removing overburden from the open pit mine in the distance, a host of dignitaries, including congressmen and California state representatives, gathered for Molycorp?s official groundbreaking ceremony at the Mountain Pass Mine on April 20. The mine, recognized as arguably one of the richest rare earths deposits in the world, is at the center of Molycorps? strategic ?Mine to Magnets? plan. It aims to secure America?s future supply of rare-earth-dependent technologies by developing a strong domestic rare earth industry. Molycorp identifies rare-earth-dependent technologies as those that cannot be produced without rare earth materials. The list of these technologies is long and growing. It encompasses everything from personal electronics and batteries for hybrid cars to permanent magnet generator wind turbines and national defense applications.

Jan 1, 2011

Mansfield MERRIMAN,Lehigh University, South Bethlehem, Pa. (communication to the Secretary*):—The formulas established by Mr. Campbell require the use of tables in order to take into account the influence of manganese on the tensile strength of steel. On examining these tables I find that the quantities given therein may be expressed by simple algebraic forms, so that formulas for tensile strength may be written which can be used without the help of tables. Referring to the case where the carbon is determined by

Jan 1, 1906

By A. F. Witt, A. M. Gaudin, A. K. Biswas

Finite contact angles in the rutile-water-organic liquid system are traceable to contamination or are due to addition of a surface-active agent. Finite contact angles are invariably accompanied by hysteresis.1-7 Besides surfactant concentration and pH, the nature of the organic liquid is found to be a factor influencing the magnitude of contact angle hysteresis. In a recent paper the authors have reviewed principles and applications of measurement methods for contact angles and their hysteresis.22 In this paper, these methods are applied to a system consisting of rutile, an organic liquid (benzene, decane, or dode-cane) and water. Surfactant (dodecylamine hydro-chloride) was used at various concentrations and various pH values. MATERIALS A synthetic rutile crystal supplied by the Linde Co. was used. It is a cylinder 1/2 in. diam, with faces cut perpendicular to the c-axis. The crystal is stated to be alumina-free and to be 99.9% pure. Rutile is a high surface energy solid, its calculated average field strength being 2.7 x 105 e.s.u.,9 its dielectric constant10 is 114. The zero point of charge of rutile was measured11 and found to be at pH 6.0. The atomic distances and bond angles of the tetragonal crystal of rutile have been precisely determined.l2 The organic liquids were specially purified. Reagent grade benzene was repeatedly shaken with concentrated sulfuric acid to remove thiophene, then neutralized with sodium hydroxide solution, rinsed with distilled water, dried overnight over anhydrous calcium chloride and filtered. The refractive index of this chemically purified benzene was 1.4977 (at 25°C) - very close to that of pure benzene (n25 1.4979).13 Both n-decane and n-dodecane were washed with aqueous sodium hydroxide, concentrated sulfuric acid (repeatedly, until the acid layer was colorless) and again with alkali, to free the organic hydrocarbon from fatty acids and unsaturated hydrocarbons. The organic phases were subsequently washed with distilled water and dried over anhydrous sodium sulfate. The filtered hydrocarbons were then distilled under vacuum (9-10 mm pressure). Decane, thus purified, showed traces of unsaturated and of carboxylic compounds. On gas chromatographic analysis the purified n-dodecane showed the presence of two impurities. However, the relative retention times of these components did not change substantially when measured on a polar or a nonpolar column. This indicates that the compounds were probably in the same polarity class, i.e. the components were nonpolar like dodecane. Conductivity water has always been one of the liquids. Conductivity of the water used was 10-6 ohm-' cm". Dodecyl ammonium chloride was prepared from pure amine (supplied by Armour and Co., Chicago), by Li in the MIT Mineral Engineering Laboratory.14 The critical micelle concentration at 25°C is known, 0.0138 mol per liter15 or 3.06 gpl. All our experiments have been done at concentrations lower than the critical micelle concentration. The glass vessels, glass drop holder and the rutile crystal were all cleaned with hot chromic acid followed by distilled water, concentrated hydrochloric acid, and again distilled water, and conductivity water. During the experiments, a nitrogen atmosphere was maintained. For this purpose, commercially purified nitrogen was further purified by passing it over hot copper wire at 450°C, then through columns of acidic chromous chloride solution, concentrated alkali solution and distilled water. APPARATUS AND METHODS Hysteresis was produced by the "captive drop" or constant volume drop method as well as by the "expanding-contracting drop" or variable volume drop method. The image of the drop while it is expanding, contracting or moving is projected on a screen and photographed. The developed film is projected through an enlarger and the contact angles measured with a protractor.

Jan 1, 1964

Experimental petrology may be used to accurately constrain PTX conditions in order to calibrate and understand geologic processes. This method may be used to answer postulated questions, by recreating known geologic conditions and observing changes in the experimental charges. The purpose of this project was to determine the eutectic composition and temperature in the ternary system PbS-FeS-ZnS ¦ (Ag2S), at ambient and elevated pressures and elucidate if the sulfides of the Broken Hill orebody may have been partially molten during highgrade (granulite facies) metamorphism. The study determined the eutectic temperature and composition in the system PbS-FeS-ZnS to be 800¦2¦C and 62.5% PbS, 30.5%FeS, 7% ZnS respectively, at one atmosphere, and the eutectic temperature at 27 kbars was found to be 960¦C¦10¦C, with a eutectic composition enriched in ZnS relative to the one atmosphere eutectic. the addition of lwt.% Ag2S to the system depresses the eutectic temperature to 772¦2¦C at 1 atmosphere and below 810¦5¦C at 5 kbars. The conditions of peak metamotphism for the cental and southern areas of the Broken Hill Block, New South Wales, Australia , lie above the solidus of the PbS-FeS-ZnS-Ag2S system, strongly suggesting that partial melting occurred in the orebody.

Jan 1, 1998

By S. L. Bessinger

Introduction Extensive testing in the past ten years has shown that where a uniform natural gamma background is present in the strata bordering a seam, the thickness of the boundary coal left in place after mining can be determined by measuring the attenuation of that radiation (Nelson and Bessinger, 1989). Measurements made by the authors in underground mines in Pennsylvania, West Virginia, Ohio, Illinois and Kentucky and by others in Wyoming and New Mexico have shown the presence of such a gamma background (Nelson. 1989). Natural gamma coal-thickness sensors of several configurations have been tested in mines owned and operated by the Consolidation Coal Company (Consol) in Pennsylvania and West Virginia (Nelson and Bessinger, 1988). This paper describes the installation of a natural gamma coal-thickness sensor on an operating continuous miner. Previous tests had shown that the NGB-1000 coal-thickness sensor manufactured by American Mining Electronics, Inc., of Huntsville. AL, is an accurate, mine-worthy instrument. This large gamma detector consists of a sensing head and a control panel. The sensing head contains thallium-doped, sodium iodide scintillating crystal, which is coupled to a photomultiplier tube. The control panel contains the electronic components required for calibration, count conversion and display to the operator. Methods Conditions at a Consol mine in northern West Virginia require that 10 to 15 cm (4 to 6 in.) of coal be left at the roof boundary of continuous miner development sections. This roof coal is required because the shale of the immediate roof is friable and unstable. In the past, operators have used a dirt band that is usually visible near the top of the seam as a guide in maintaining the proper cutting horizon. However, this is not always reliable. Earlier observation showed that the actual thickness of the coal left on the roof varied widely; further, it was noted that occasional, accidental excursions into the immediate roof required supplementary roof control measures, such as installation of planks or center bolts. Thus, it was concluded that operators needed a better source of guidance for control of the cutting horizon, and a roof-coal thickness sensor was scheduled for installation. The NGB-1000 sensor was installed on a Joy 12CM10 continuous miner in June 1988. The sensing head was mounted on the cutter boom of the miner, and the control panel was mounted in the operator's cab. Power for the sensor was initially derived from an intrinsically safe battery power supply. Initial measurements with the sensor showed that the calibration was the same as that used in earlier tests at two other mines, indicating the uniformity of the natural gamma background above the Pittsburgh seam. Operating personnel were initially skeptical of the instrument's accuracy, and were hesitant to use its readings as a guide in maintaining a proper cutting horizon. Because gamma attenuation, the instrument's operating principle, is somewhat abstract, attempts to demonstrate the instrument's accuracy by explaining that principle were generally ineffective. It was found, however, that an operator could usually be convinced of the usefulness of the instrument by placing a large piece of coal of fairly uniform thickness over the instrument's sensing head and allowing the operator to see that the instrument reading increased by an amount very near his estimate of the thickness of the piece. The mine was provided with seven battery power supplies and a charging station. The charging station was kept in the lampman's office, and the mechanic on each shift was instructed that he was responsible for two battery power supplies each day: a freshly charged one to be taken in at the beginning of his shift and a depleted one to be brought out at the end. This system worked well for a few weeks, but eventually some battery power supplies were left in use so long that their batteries were discharged too deeply to allow recharging. In addition, transport and recharging of the batteries represented an additional task for the mechanics, who were already very busy. Consequently, a request was filed with MSHA to allow the sensor to be powered through intrinsic safety barriers by an electronic power supply connected to machine power. The permit was granted, and the sensor was connected to machine power. After the sensor was connected to machine power, the only operating problem experienced was occasional failure of cables. A supply of the required cables was made and delivered to the mine so damaged cables could be quickly replaced. Much of the cable damage could be eliminated by slight modifications to the miner during a rebuild, so that cables could be installed in more protected locations. After the sensor had been in operation for about two months, a survey was made to determine its effect on continuous miner operations. In previous research, coal thickness measurements made in 88 locations by the natural gamma method were compared to measurements made in the same locations by observing drill cuttings and by inspections of drill holes with a borescope. That research showed that the gamma method is at least as accurate as the other two methods (Nelson and Bessinger, 1989) and is also much easier to use. The object of the survey described here was not to assess the accuracy of the natural gamma measurements. but rather to determine the effectiveness of the sensor output as a guide for the operator in maintaining control of the cutting horizon. Thus a smaller, hand-held gamma detector

Jan 1, 1992

By E. R. Borcherdt

IT is notable that the first large-scale mine operation equipped entirely with detachable bits was the Badger State mine of the Anaconda Copper Mining Co. in Butte, Montana, just 30 years ago. This mine in 1922 was producing approximately 1200 tons of ore per day. Much of the data presented in C. L. Berrien's article' describing the development and installation of the Hawkesworth detachable drill bit were obtained from these operations. As in any pioneering effort, no precedent existed and many difficult problems required solution, so that the changeover to detachable bits at all Butte hill mines was not completed for 6 years. There was widespread disbelief as to the probable efficiency of the new installation. Some attempts were made in 1931 by the owners of the Hawkesworth patents to interest Ontario gold mine operators in the bit. These efforts were not successful, but they undoubtedly stimulated thinking which resulted in the invention and patenting of several well-known Canadian detachable bits, one of which is now a widely used throwaway bit. The success of the Butte installation also led to the development of the threaded type of bit connections by several well-known manufacturers, and in 1935 these bits were introduced to the mining industry on a national scale. The original Hawkesworth bit was not provided with a water hole but, depended upon water passing through the clearance opening between the tongue in the bit and the groove in the rod to flush cuttings from the drill hole, see Fig. 1. In December 1935 it was found that this method of introducing drilling water to the bit face resulted in high dust counts. To correct this a water hole was drilled on the central axis of the bit, passing through the tongue. Unfortunately, quenching water would rise through the small water hole, spot-hardening the tongue to cause breakage, never completely eliminated. In the fall of 1936 large-scale tests indicated that savings would be effected by use of a threaded type of bit, which was therefore adopted as standard for all Butte mines. This type of bit was used until 1947, when it was superseded by a one-use slip-on type. Since the first use of the Hawkesworth bit every detachable bit of importance has been investigated, and where advantages which might reduce costs or increase efficiency were indicated, substantial tests of the bit were carried on in the Butte mines. When tests demonstrated the advisability of changing from one kind of detachable bit to another the change was made at one level or in one area each day until the new rod and bit equipment was used throughout the mine. This involved a minimum of cost and disruption of drilling. Intelligent selection of a detachable bit to obtain optimum results requires careful consideration to achieve a balance between the three principal types of equipment used in the drilling process: 1—drill bits, 2—drill steel, and 3—drilling machines. Optimum results imply maximum output and minimum cost per unit of output. Since every rock type differs in drillability and it is generally impractical to provide equipment for more than one or two types of rock which may occur in one operation, selection of equipment must encompass average drilling conditions. However, on exceptional occasions several widely differing conditions may make it mandatory to provide equipment best suited to each condition. The choice of rock-drilling equipment is a most controversial subject and one that is further complicated by unreliable and frequently misleading performance claims. Small operators without the means for making accurate evaluations of equipment frequently suffer from these over-enthusiastic claims. It is apparent from experience in rock drilling throughout the world that rock drillability is not alike in any two places, and that selection of proper equipment can only be made after conducting thorough trials of various types of equipment. Some recent drilling tests in tactite and hornstone at the Darwin, California mine of the Anaconda Co. present some interesting clues on rock drillability. Microscopic examination of thin sections of these rocks reveals that mineral composition and rock texture are equally important in governing drillability. The Darwin hornstone is at times so abrasive that the carbide bit cutting edges become flattened to 3/32 in. in 2 to 4 ft of drilling, and some carbide bits were dulled to this point after 9 to 10 in. of drilling. This wear was determined to be the proper point for resharpening to eliminate carbide insert breakage or breakage of the steel rod when drilling with 1½ to 1?-in. bits, with a drifter of 2 3/4-in. diam and 90 to 100 psi air pressure, see Supplement A. Before considering the merits of various bit designs it may be well to review the mechanics of drilling rock with percussion drills. A sharp bit cuts by penetration and chipping. The amount of penetration governs the amount of chipping and depends upon the contact area of the cutting edge, the foot-

Jan 1, 1954

1. Data from the Mine Safety and Health Administration (MSHA) indicate that a total of 44,015 employees worked in underground mining locations in 2003, corresponding to 44,778 full-time equivalent (FTE) workers. • Ninety percent of underground employees (n = 39,735) were mine operator employees. • Contractors comprised 10% of all underground mining employees (n = 4,280) and comprised only 5.8% of all underground FTEs. • The majority of underground employees worked for coal operators (73.0%; n = 32,130).

Feb 1, 2005

By Johnson T, Huynh D

The State Electricity Commission of Victoria is conducting a program of research and development into alternatives to the use of conventional boilers for the generation of electricity from brown coal. The objectives of the program are to reduce the cost of electricity generation and also to improve the thermal efficiency so as to reduce the emission of greenhouse gases and other pollutants. The program presently includes the following four technologies: Hydro-thermal Drying, Direct Coal Fired Gas Turbine, Integrated Gasification Combined Cycle, and Steam Fluidised Bed Drying. The Hydro-thermal Drying process produces a combustible coal-water-slurry fuel which is suitable for use in a coal fired gas turbine. The coal fired gas turbine has the potential to produce electricity at a lower cost than conventional power stations and when combined with a steam cycle, it will also use significantly less coal. The Integrated Gasification Combined Cycle process gasifies the coal, cleans the gas, and then bums it in a gas turbine/combined cycle. This process is being developed in Europe and the USA in order to meet their stringent emission regulations. It offers Victoria cleaner more efficient and cheaper electricity.

Jan 1, 1992

Reduce the dust exposure of workers in large processing buildings at mineral processing operations, using a total mill ventilation system.

Jan 1, 1994

By P. K. Koh

With identical annealing heal treatment the development of major annealed texture component seems to depend primarily upon the degree of cold reduction. Cube texture was evident on annealing- 1/2-, 1,/4- , and 1/8-mil tapes at 760°C, but it dissap-)eared on 927°C annealing of 1/8-mil tape cold reduced 96.9 pct. Some cube texturee remained as tile strongest 927 °C annealed texture component in 1/4- and 1/2-mil tapes cold reduced 93.8 and 87.5 pet, respectively. The switching coefficient increases and the squareness ratio of the hysteresis loop decreases proportionally with increasing cube pole concentration. PREVIOUS studies 1,2 on the 1/8-mil ultra-thin molybdenum permalloy tape established cold rolled and annealed textures and their magnetic performance for switch-core and coincident-current memory applications, The major cold-rolled texture was found to be of {110}<335> which is quite similar to the cold-rolled texture of silver and 70-30 cartridge brass. The major annealed texture components were found to be of {120)-- <001> and {113)<785>. Although the direction for easy magnetization for this type of alloy by the single-crystal technique is not avaliable, knowledge from related studies indicated [Ill] as the easy magnetization direction, similar to that of nickel.&apos; Therefore, a cube orientation, if present in the tape core, would be most detrimental to its magnetic performance.3 In its cold-rolled state the tape had < 111 > directions in the rolling plane 14.5 deg from the rolling direction.&apos; But the coercive force of the tape in such condition will be excessive due to large amounts of internal stresses, and therefore makes its use impossible for memory applications due to long switching time. In lower nickel alloys, suchas 50-50 nickel-iron and 65 pct Ni-Fe sheets, previous investigators4"7 found the presence of cube texture on annealing at 900° to 1000°C. Although no cube texture was found in annealed 1/8-mil tape,"&apos; its presence in some annealed 1,/4-mil tape cores was firmly established a year ago as verified by significant differences in magnetic properties. The present investigation is devoted to the orientation study and related magnetic measurements on three thicknesses of the ultra-thin tape, namely, 1/8 mil, 1/4 mil, and 1/2 mil, cold rolled directly from annealed 4-mil stock in the cold-rolled state as well as after they were annealed at 760°C (1400° F) and 927°C (1700°F). The results obtained might serve to elucidate the development of annealed texture as well as to establish clearly the importance of preferred orientation on the magnetic performance of such memory tape core despite its well-known low magnetic anisotropy exchange energy. EXPERIMENTAL PROCEDURE Samples were taken from one production heat (0.010 wt pct C, 0.004 P. 0.016 S, 0.80 Mn. 0.29 Si. 79.69 Ni. 4.42 Mo. 0.054 Co, 0.002 Ti, 0.12 Al, 0.04 Mg, 14.56 Fe). Annealed 4-mil sheets were cold re-

Jan 1, 1962

SELWYN G. BLAYLOCK, who has been awarded the Douglas medal for 1927 for his achievements in the production of electrolytic zinc and the treating of complex lead, zinc and silver ore at the Sullivan mine, British Columbia, is a Canadian and one of the notable men who have come from McGill Uni-versity, where he graduated in 1899. As the medal was founded in 1922 by a group of James Douglas&apos; friends to honor the memory of that distinguished Canadian metallurgist, who was an honorary LL.D. of McGill, it is extremely appropriate that it should be awarded to another Canadian metallurgist and McGill graduate, who has been notably successful in practice. Immediately after gradua- tion Mr. Blaylock obtained a position as chemist at the Trail smelter and two years later was made chief chemist and metallurgist. In 1907 he became superintendent of the Hall mines smelter and after a year became superintendent of the St. Eugene mine at Moyie. While there he was responsible for the purchase of the Sullivan mine and the use of methods of mining, especially sorting, which so improved the relative per- - centages of zinc and lead in the ore as to make it better material for blast-furnace smelting. The losses in zinc, both from sorting out and in the lead blast-furnace slag made a deep impression on him, and various methods of water concentration and the use of air-jigs and flotation to increase the recovery were tried under Mr. Blaylock&apos;s supervision as early as 1910. These were followed by experiments with direct fusion, volatilization in kilns, sulfite processes, and one similar to the Ashcroft proc-ess. In 1914 experiments were made with roasting, leaching, and electrolyzing the resultant zinc sulfate solution; the following year a plant was built at Tadanac which has been in successful operation since 1916, its capacity having been increased by successive steps to 275 tons of zinc per day. Experimental work continued on the feed for the zinc plant and a 150-ton mill to use the Horwood process was constructed, but as the results were not encouraging it was altered to wet magnetic concentration, which was successful and the plant was later increased in capacity to 600 tons per day. Meanwhile experiments with differential flota-tion showed promise and magnetic concentration was continued while flotation was developed to where it could be put into operation parallel to the magnetic plant. This flotation plant, started in 1920, was so successful that eventually the magnetic separation plant was abandoned. Lead smelting conditions at Trail were difficult on account of the necessity of mak-ing slags containing as much as 20 per cent zinc and because of the high percentage of fine flotation con-centrate. The improvements in sintering and smelt-ing made under Mr. Blaylock&apos;s direction have been notable. The Betts electrolytic refining process for lead was first used at Trail, and the plant has been brought to its present highly developed and efficient state under Mr. Blaylock&apos;s direction. The first com-mercial Cottrell plant in connection with lead smelt-ing was also built there. The capacity of the Tada-nac plant is now approximately 750 tons per day of refined metal, including lead, zinc, copper, antimony, cad- mium, gold and silver. All the work necessary for this nota- ble development has neces- sarily been the work of many hands, directed by Mr. Blay- lock. In 1924 he was awarded the McCharles Medal by the University of Toronto for his achievements in con- nection with the production of electrolytic zinc and the solving of the difficulty of treating the complex lead- zinc-silver ore of the Sullivan mine, and developing it from a small producer to one of the most important mines of the world, now producing 150,000 tons of lead and 90,000 tons of zinc annually. Having spent all his professional life in one remote place, Mr. Blaylock is relatively little known per-sonally to others of his pro-fession, and the award indi-cates the truth in Emerson&apos;s well known epigram about the mouse-trap.

Jan 2, 1928

By C. W. Andrews

The various methods of by-product coke-manufacture have been quite thoroughly discussed in a number of papers recently published, and I think it would possibly be of interest and value to discuss the growth and probable future of the byproduct coke industry, rather than details of manufacture. The first by-product coke-plant in the United States, consisting of 12 ovens, was put in operation at Syracuse in January, 1893. This installation was followed by one at Johnstown, Pa., in 1895; one at Dunbar, in 1896, and one at Glassport, in 1897. The plant at Syracuse was primarily intended to supply the owners with ammonia for their alkali-works. At Glassport, for the first time, gas was sold for illuminating purposes, although it was mixed with water-gas to make it marketable. The New England Coke & Gas Co. started the plant in Boston in 1899, which was destined to change the whole trend of the establishment of by-product plants, since it manufactured for the first time illuminating-gas suitable for sale without enrichment. This was accomplished by dividing the gas on the battery into two portions. One portion, consisting of the richer part of the gas, was saved for sale, while the remaining, or leaner, portion was used in the battery. This idea was originated by Dr. Schniewind, of the United Coke & Gas Co. The coke was used for domestic, industrial, and locomotive fuel, as its sulphur-content was too high to make it suitable for use in the foundry. The successful operation of this plant led to the building of additional plants at points where the production of illuminating-gas could be sold to advantage and where there was a fair market for the coke-output. The following list of illuminating-gas plants in order of construction will give a good idea of the rapid growth of this method of combined gas- and coke-manufacture: Hamilton, Ohio; Detroit, Mich.; Camden, N. J.; Baltimore, Md.; Chester, Pa.; Milwaukee, Wis.; Duluth, Minn.; Geneva, N. Y.; Chicago, Ill.; Indianapolis, Ind. In addition to these there are now under construction plants at Muncie, Ind.; Joliet, Ill., and Waukegan, Ill., all of which will produce illuminating-gas and metallurgical coke.

Jan 1, 1913