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By L. A. Neumeier

The Bureau of Mines evaluated the properties and anisotropy in properties of rolled Zn-Ti and Zn-1.0 wt pct Cu-Ti alloys containing up to 0.36 wt pct Ti. Finish rolling was at varied temperatures to total reductions as high as 92 pct. Alloy properties are very sensitive to various combinations of composition, rolling temperature, and reduction. Properly rolled Zn-1.0 Cu-Ti alloys are superior in strength to Zn-Ti alloys. Conditions are described that lead to best tensile properties. Creep resistance for Zn-Ti alloys decreases with increased rolling temperature when finish rolled to 92-pct total reduction: but not when finish rolled to 33- or 67-pct reduction. Creep resistance for 7n-1.0 Cu-Ti alloys improves with increased rolling temperature to about 480° F. Heavily cold-rolled Zn-1.0 Cu-Ti alloys have fine grain sizes and poor hardness, tensile, and creep properties, but high ductility. Thermal expansivity is markedly affected by the copper addition. Data show the anisotropy of tensile, creep, and thermal expansivity properties. Large idiomorphic crystallites of TiZn15 appear in both Zn-Ti and Zn-Cu-Ti alloys when more than about 0.2 pct Ti is present. A new Zn-Cu-Ti intermediate phase was also identified. Empirical equations are presented relating creep rates with major process variables.

Jan 1, 1971

Executive JOHN M. LOVEJOY, Chairman ERLE V. DAVELER G. B. WATERHOUSE WISER JUDSON WILLIAM WRAITH J. V. W. REYNDERS, Consultant Finance HENRY KRUM, Chairman PAUL D. MERICA R. M. ROOSEVELT H. G. MOULTON, Consultant Admissions. RALPH M. ROOSEVELT, Chairman GEORGE D. BARRON, Vice-chairman JOHN L. CHRISTI PHILIP KRAFT W. A. SINSHEIMER

Jan 1, 1936

By P. J. van Staden

Several differences exist between the columns used for laboratory simulations of heap leaching and commercial heaps. Column walls support part of the weight of the ore and restrict the particle packing arrangements. Columns are typically of a diameter that is smaller than the dripper spacing to be used on the commercial heaps. The stacking practices that cause segregation and stratification on heaps are absent from columns. Intuitively, these physical differences would be expected to be the reasons behind the observed differences between column and heap leaching performance. But how can they be quantified, and what is the magnitude of the differences? Could some of these differences be built into column leaching experiments in order to better simulate commercial-scale heap leaching performance to ease the scale-up calculations from column leaching data? Alternatively, could heap construction practices be improved so that the ideal performance obtained in columns could be emulated on heaps? Data from a few heap leaching cases studies that permit direct comparisons between column and heap leaching performance is analysed, to reveal the extent of the differences in resistance to mass transfer. Discussion is provided on the implications for column and heap leaching practice, and for scale-up design procedures. INTRODUCTION Differences have been known to exist between leaching performance observed in laboratory columns and that of commercial-scale heaps. Both the heap leaching kinetics and finally achievable extent of extraction are generally found to be inferior in heaps compared to that achieved in columns. Heap leaching design engineers extrapolate the anticipated heap leaching performance by multiplying the time-axis of a column-leaching curve by a factor of about 2 and the extraction-axis by a factor of about 0.8, as explained by Jansen and Taylor (2002), Scheffel (2002), John (2011) and Scheffel, Guzmán, and Dreier (2016). The procedure is illustrated in Figure 1 by means of a hypothetical example. The solid line represents the leaching performance observed in a laboratory column, and the dashed line represents the anticipated leaching performance of a heap, extrapolated from the column leaching data using these empirical scale-up factors.

Jan 1, 2019

By Booth S. A, Milton D. W

Stratabound/stratiform and remobilised mineralisation have been recognised as the two dominant styles of mineralisation which constitute the Broken Hill ore deposit. Recognition of prospective sites for accumulation of ore depends on understanding the heterogeneity of the ore environment and the distribution of the two styles of mineralisation. The internal stratigraphy of the Broken Hill ore deposit and its enveloping rocks, is described, and a generalised stratigraphic setting of the stratiform mineralisation is outlined. Fhiphasis is placed on the existence of closely-spaced sub-basins as potential sites for ore accumulation close to the Broken Hill ore deposit. Further from the immediate ore environment, recognition of the same diagnostic stratigraphic sequence

Jan 1, 1983

By E. Contestabile, B von Roser

Schedule IV containers in Canada and IME- containers in the United States have made possible the transport of mixed loads. That is, using these containers for detonators, one can simultaneously transport detonators and explosives on the same vehicle. As a result of a recent accident, the Canadian Chief Inspector of Explosives asked for a review of the appropriateness of the Schedule IV container in its present use. The Canadian Authority asked for a re-assessment of the adequacy of the Schedule IV container in preventing propagation of a reaction from the detonators in the container to the remainder of the explosives on the truck, and its ability to contain fragments generated by the detonators. A series of tests was performed to determine the initiation threshold temperature for non-electric detonators and the failure threshold temperature for the shock tube to establish some guidelines for preventing communication from one case of detonators to another within the transport container at elevated temperature. Tests fell into the following categories: inter-case sympathetic detonation, detonator thermal auto-initiation, shock tube thermal sensitivity, and intra-case sympathetic detonation. The results indicate that temperature is a critical factor, which determines to what extent, if any, communication occurs.

Jan 1, 2002

By Strathmore R. B. Cooke, Iwao Iwasaki, Hyung Sup Choi

The effect of temperature as a parameter in ore flotation has not been systematically studied, although for some ores it has been known for many years that selectivity and grade of concentrates can be improved by conditioning or flotation at moderately or substantially elevated temperatures. In 1934 Coghill and Clemmer' stated that "elevated temperature seems to make hard water more tolerable" in soap flotation, and a cursory examination of the literature indicates that some earlier investigators were aware of beneficial results obtained by floating above room or mill-water temperatures. Tartaron,' Mitchell et al.,3 and particularly Hamilton et al." have shown that high pulp temperatures give high recovery and selectivity in soap flotation of fluorite from a variety of gangue minerals. As a result of the work of these and of other investigators, high-temperature flotation has been successfully applied to industrial flotation of fluorite. Falconer" stated: "The use of high temperature conditioning (above 35°C and preferably above 60°C) in connection with the soap flotation of nonsulphide ores, as covered in B. Kalinowski's French Patent No. 847,- 215, Dec. 7, 1938, is claimed to provide more intense activation, better separation of values from gangue, and reduction in quantity of reagents." The various investigations quoted above clearly demonstrate the improvements to be realized by increasing the temperature at which fluorite is floated, and it is surprising that little or no information has been published on systematic investigation of the same method as applied to other ores.

Jan 5, 1960

By L. R. Plitt, E. O. Lilge

Tests carried out with both kaolin and silica slurries show that flocculants of the polyacrylamide type can be used to improve the thickening performance of hydrocyclones. This thickening improvement demonstrates that contrary to previously held theories, flocs can be formed which are capable of resisting the shear forces in a hydrocyclone. For the 1.25-in.-diam cyclone used, optimum thickening occurs when the flocculant solution is injected into the slurry stream at or near the feed inlet. Hydrocyclones offer many distinct advantages over gravity thickeners. These advantages include simplicity, low initial and operating costs, small space requirements, and flexibility of operation. In spite of these advantages hydrocyclones have not been widely used as thickeners. The main reason for their lack of application is that hydrocyclones are unable to efficiently remove semicolloidal particles (less than 5 microns) from the suspending solution. In gravity thickeners, the fine particles can be induced to form particle aggregates, or flocs, which have the settling characteristics of large particles. In the hydrocyclone, however, it was always assumed that the existence of shear would prevent the formation of flocs.1-3 Thus, if no floc structure is retained, the thickening hydrocyclone must be designed so that its separation size is below the size of the smallest particle to be recovered. To obtain a very small separation size requires the formation of very high centrifugal forces, which necessitates the use of small-diameter cyclones. Small hydrocyclones have, in turn, very low throughput capacities which render them impractical for many industrial thickening applications. Thus, if the effects of flocculation cannot be utilized, the usefulness of hydrocyclones as thickeners remains limited to pulps which contain no very fine particles. It is well established that most substances acquire a surface electric charge when brought into contact with an aqueous medium. The repulsive interactions between similarly charged particles act to prevent flocculation. One method of flocculating a dispersion is to neutralize the repulsive surface charges by the addition of an electrolyte. The reduction of the electrostatic surface charge (Zeta potential) then permits the universal van der Waals attractive forces to operate between the atoms of the various particles and form particle aggregates. The lime additions used by thickener operators promotes flocculation by this mechanism. A second method of flocculating a dispersion is by the addition of long-chain macromolecules. In this case flocculation is brought about by a bridging mechanism in which the molecules are adsorbed with part of their length on two or more particles, thus forming a molecular bridge between the particles.4,5 The molecules also form bridges between themselves when a single particle is bonded to several polymer molecules. The type of bonding between the flocculant molecule and the particle may be hydrogen bonding, chemical bonding, or electrolytic attraction.6 The flour and glue which are used as thickener additives are believed to promote flocculation in this manner. In the past decade synthetic long-chain polymers with extremely high flocculating capabilities have been developed. The flocs formed by these new polymeric flocculants are larger and more shear-resistant than those formed by the presence of electrolytes.7 The advent of these flocculants raises the question: Are the flocs formed by these synthetic polymers stable enough to resist the liquid shearing forces in a hydrocyclone? The investigations reported in this paper were carried out in order to provide some answers to this question. EXPERIMENTAL WORK For the bulk of the test work a 5% slurry of commercially available kaolin with an average particle size of 5 microns was used. An initial evaluation of the available flocculants was carried out to determine which flocculant formed the most shear-resistant flocs. A combination of the techniques used by Healy7 and Booth8 was used for this purpose. This initial study revealed that the neutral high-molecular weight polyacrylamides produced the most shear-resistant flocs. One of these flocculants, Separan MGL*, was selected and then used for all the hydrocyclone tests. The flocculant was dissolved in water and added to the hydrocyclone as a dilute solution.

Jan 1, 1968

By Howard G. L

The gold treatment plant of Great Western Consolidated N.L. is at Bullfinch, Western Australia, 22 miles north of Southern Cross, which is 100 miles west of Kalgoorlie. A railway to Southern Cross connects the mine with the main Perth-Kalgoorlie line.The plant, which commenced operations on the 31st October, 1952, was, designed by Messrs. Alex: Stewart & Co. in consultation with the technical staff of Western Mining Corporation Ltd. to treat 1,000 long tons a day, made up of approximately equal parts of quarry oxidized ore and underground sulphide ore, by moderately coarse grinding (2-3 per cent plus 65 mesh Tyler) and straight cyanidation without straking. The design had to allow for future treatment of all sulphide ore at the rate of 40,000 tons per 28-day periodand to make provision for the possible later inclusion of preaeration treatment.The test work on .which design was based was done on representative development samples, the mine being divided into zones for this purpose, to check possible local variations from average treatment behaviour. Proposed mill feed mixtures were made up from time to time for complete ore dressing tests, which were conducted in the Western Mining Corporation laboratory at Kalgoorlie.The chief feature of the design was that the plant had to handle the extremes of very soft, poor settling and filtering oxidized ore and very hard, good settling and filtering sulphide ore. On one hand, crushers and mills had to be large enough to handle the sulphide ore and on the other, thickeners and filters had to be large enough to handle the oxidized ores.The design is marked by simplicity, the use of large units of equipment, and flexibility to cope with a variety of ores.

Jan 1, 1956

By D Burgess

Since the days of mandatory high cost pilot plant testing in 6 ft + 2 ft SAG mills autogenous griding/semi-autogenous grinding (AG/SAG), mill specific energy is now calculated by various empirical formulae and pilot testing carried out in small laboratory mills. Barratt (1989) described a method using Bond Indices than adding a single factor of 1.25 to arrive at the AG/SAG specific energy. In 1991 the Canadians developed the SAGPower Index (SPI), test using 2 kg of material in a 300 mm diameter mill. This test has been further developed by Starkey, Hindstrom and Orser (2006). This test now includes a Bond Ball mill work index test.The Macpherson test is still being offered by some testing facilities. This dry test requires at least 180 kg of material.Morrell (2004) formulated an AG/SAG specific energy generator based on a drop weight index Value (DWi). This DWi value is related to the Ab parameter and the material specific gravity. Morrell reconfigured the BondÆs Third Theory formula to cause extra power requirements to be inputed. This technique incorporates the material specific gravity (SG). The AG/SAG specific energy values produced using this method are susceptible to higher values of crushing and rod mill work indices.This paper offered describes a more simple method to generate AG/SAG specific energies. This programme incorporates Bond work indices, the JKMRC parameter Ab measuring impact breakage and the parameter ta measuring the abrasion rate, then employs a series of efficiency values nominated for each transition size (AG/SAG T80). These efficiency values have been compiled over the years from actual operating plants and documented reports.CITATION:Burgess, D, 2012. A method of calculating autogenous/semi-autogenous grinding mill specific energies using a combination of bond work indices and Julius Kruttschnitt parameters, then applying efficiency factors, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 37-44 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

By H. R. Peiffer

The increase in electrical resistivity, ?pT,at 78°K was measured as a function of elongation, E, at 78°K for a 2 pct (approximately) by weight finely divided alumina in silver material. The amount of increase in electrical resistivity (?pT) is a function of the amount of alumina present and can be represented by: ?pT = cen where c and n are constants. The constant c increases and the exponent n decreases with increasing alumina content. Approximately 33 pct of the resistivity increase is recovered upon annealing at room temperature. RECENTLY, dispersed phase alloys' have shown promise for use at elevated temperatures. These materials have a high creep resistance, a high-temperature strength considerably above that of the pure matrix, and a high electrical conductivity. These materials also show other properties such as retardation of re crystallization and resistance to flow even above the melting point of the matrix.' It is believed that high-temperature creep is retarded in these materials because the dispersed phase blocks the motion of dislocations. According to schoeck3 and also Weertman,4 the rate controlling process is the climb of dislocations around the dispersed phase particles. Any mechanical property of these materials, or any other material for that matter, can be explained on the basis of defect structures and their interrelations. It would certainly be interesting to learn something about the effect of a dispersed phase on dislocation densities and also their effect on point defect concentrations following cold work. This is the purpose of the research discussed herein. The method used was the measurement of the electrical resistivity as a function of elongation at low temperatures. This scheme has already been applied to the studies of pure metals.5 Electrical resistivity measurements conducted on pure metals such as copper, silver, gold, and so forth, have indicated that defects such as vacancies and dislocations are formed during deformation.5 It has been shown by van Bueren6 that the resistivity change, ?p, should increase with elongation, E, according to the following relation: Ap=Al/2 + B£3/2 where A and B are constants of the system; ?p is in µ? cm. The first term refers to the amount of resistivity change associated with line defects and the second term with the change associated with point defects. A later calculation yielded exponents of 3/4 and 5/4, respectively.6 Manintveld7 has shown that the resistivity of copper increases with strain according to a 3/2 power law, whereas van Bueren has shown that the resistivity of silver increases according to a 5/4 power law. Assuming then that van Bueren's calculations are correct, it appears that line defects contribute very little to the electrical resistivity. There is some doubt, however, as to the validity of van Bueren's expressions for all cubic metals.' In fact, the data on copper and silver are not conclusive. In this in-

Jan 1, 1962

By J. Craig Green

Companies mining for coal, oil shale, and minerals in the western US must often complete extensive geotechnical studies to satisfy federal and state regulations. The modern mining process is further complicated by the legal problems and liabilities associated with mineral rights and land leasing. Yet the study of water rights-unique to the West is often incomplete or ignored until the latter stages of mine planning. Water rights in the West are not adequately understood by many mining companies and their geotechnical consultants. Some firms, headquartered outside the western US, have not had to contend with the water supply problems of a semi-arid region. Water problems encountered by eastern mining companies, for example, are primarily water quality, drainage, and flood control. Even when water quantity is a problem in these areas, ownership and/or use of it is governed by entirely different legal and administrative concepts than in the West. Both legal and engineering issues have to be considered when studying western water rights. Legal issues, such as existence and title, must be addressed to insure that a particular water right is properly established and maintained. The exercise of a water right, however, often requires engineering analyses that focus on preventing injury to the properly established water rights of others.

Jan 8, 1981

By R. C. Freas

Limestone is a basic raw material in the production of cement, lime, a wide variety of industrial products, and a significant percentage of the crushed stone produced in the United States annually. It is used in fluxes, glass making, mineral fillers, environmental applications such as flue gas desulphurization and water treatment, soil conditioners, animal feeds, and as a construction aggregate. This presentation provides an overview of limestone distribution in the United States, it’s production, the variety of products produced, and a brief review of the primary uses for limestone.

Jan 1, 2006

By Tobias Hens, Andrew Frierdich, Barbara Etschmann, Joël Brugger, Barrie Bolton

"Advancing our knowledge about trace metal distribution and abundance provides insight into formation, alteration, and trace element enrichment processes of ferromanganese nodules and crusts. A fundamental understanding of these mechanisms, in turn, has implications for the selective recovery of critical metals and the development of sustainable hydrometallurgical techniques. Synchrotron-based X-ray Fluorescence (SXRF) is a powerful, state-of-the-art analytical tool that can examine geochemical features of Fe-Mn nodules and crusts, such as abundance and distribution of trace elements, in an integrated and efficient manner. By correlating trace element composition and Mn oxidation states in these highly complex samples, it can be determined whether these parameters are related to the Mn oxide mineralogy (vacancy content of crystal lattice) and chemistry.In this study, eleven Fe-Mn nodule and crust samples from three circum-Australian locations – the South Tasman Rise, Lord Howe Rise, and Coral Sea – were analyzed at the Australian Synchrotron, Victoria, Australia. Micrometer-scale spatial resolution of geochemical features, such as concentric growth structures and element distribution, was achieved with high sensitivity (Fig. 1). For instance, Ni could be detected at 10's of ppm to 1 wt% levels. Information on other redox-sensitive trace elements, which are present at 10-1000 ppm levels (e.g., Cu, Co, Ti), were gained by means of fully quantitative elemental composition maps."

Jan 1, 2017

[ ] S. H. Ash (Mining Industry Continues with Lower Fatal-Injury Rates, P. 257) was born in Roslyn, Wash. and attended Lehigh University and University of Washington. He received a B.S. and E.M. degree in mining engineering. Mr. Ash was supervising engineer in charge of work of Safety Div., Bureau of Mines, eastern Pennsylvania and the states of N. Y., Del., N. J. and the New England states. He was appointed chief engineer of Metal and Nonmetal Mining Section of the Mineral Production Security Div., Bureau of Mines, July 1942. At present he is chief engineer, Safety Branch, U. S. Bureau of Mines. Besides holding membership in AIME, he is a member of National Honorary Engineering Society, Society of American Military Engineers, Tau Beta Pi and many other societies.

Jan 1, 1952

By Thomas L. Watson

The occurrence of zircon in pegmatites of acidic composition is recorded by many observers both in this country and abroad, and they form one of the most important geologic modes of occurrence of the minera1. As is well known, commercially the most important American locality for zircon is near Zirconia in Henderson County, North Carolina, where many tons of the mineral have been obtained from a kaolinized pegmatite dike 100 ft. wide and traced for 1 1/2 miles along the direction of strike N. 50" E. Zircon may be developed in pegmatites as inclusions in the principal rock-forming minerals, chiefly quartz and feldspar, and as a separate megascopic constituent in the form of grains and crystals not exceeding, as a rule, 2 in. in size, and usually smaller. In the Henderson County, North Carolina, pegmatite, zircon occurs in prismatic crystals with pyramidal terminations measuring up to 30 mm. in diameter, associated chiefly with the feldspar. In some of the apatite veins of Canada, which are closely allied to pogmatite dikes, zircon is reported in crystals upward of 6 in. in length and 2 in. and more in thickness. Zircon in large masses occurs in Virginia in the well-known pegmatites near Amelia in Amelia County, and near Gouldin in Hanover County. Both localities are in the middle eastern portion of the Piedmont Plateau province, and are separated by a distance of about 40 miles in a northeast-southwest direction. The rocks of both areas, including the pegmatite bodies, are in an advanced stage of decay, and the hard and moderately fresh rocks are concealed beneath a cover of variable thickness of rock decay derived by the normal processes of weathering. Exposures, there-

Jan 1, 1917

By H. R. Reed

THE cost of hoisting with electricity in deep mines is a matter about which little information has been published. This paper makes a detailed study of such costs and presents the information in a series of curves. In particular the curves show the cost of hoisting ore from depths varying from 500 to 7000 ft. Vertical with the shaft at 70° incline for the double cylindroconical and double cylindrical drums, the hoists operating balanced, using Ward Leonard and Ilgner-Ward Leonard control, and designed for most economical operation from each depth. Operation with the double cylindrical drum is expanded to cover hoisting from 2000 to 5000 ft. using equipment designed for 5000 ft. and from 4000 to 7000 ft. using equipment designed for, 7000 ft. Additional curves are included to show the ratings of the necessary hoist motors, energy input per hoisting cycle and size of flywheel in tons for complete equalization when the Ilgner-Ward Leonard system is used, also hoisting efficiency. The duty imposed on a mine-hoist motor fluctuates between wide limits during a cycle of operation. The demand on the motor may vary from a large horsepower value' near the beginning of the cycle to a small or a negative demand near the end of the cycle. Such factors as tonnage to be raised per hour, depth of shaft, and balanced or unbalanced opera-tion have such a marked effect on the operating cycle and vary to such an extent for different mines that no set of rules can he formulated for quickly determining the size of the motor required. The duty cycle from which the motor rating is (derived must he determined for each case according to the conditions of the proposed hoisting operations. The motor torque required at the cardinal points of the duty cycle is determined from the moment or torque in pound-feet necessary to raise and lower the load and ropes, accelerate and retard the load, ropes, drum, etc. The torque diagram depends largely upon the type of drum. Two different types of drums, if used on exactly the same duty cycles, may have widely different peak demands and capacities of hoist motors. The essential feature in the design of hoisting drums and in determining

Jan 1, 1937

By H. R. Reed

THE cost of hoisting with electricity in deep mines is a matter about which little information has been published. This paper makes a detailed study of such costs and presents the information in a series of curves. In particular the curves show the cost of hoisting ore from depths varying from 500 to 7000 ft. vertical with the shaft at 70º incline for the double cylindroconical and double cylindrical drums, the hoists operating balanced, using Ward Leonard and Ilgner-Ward Leonard control, and designed for most economical operation from each depth. Operation with the double cylindrical drum is expanded to cover hoisting from 2000 to 5000 ft. using equipment designed for 5000 ft. and from 4000 to 7000 ft. using equipment designed for 7000 ft. Additional curves are included to show the ratings of the necessary hoist motors, energy input per hoisting cycle and size of flywheel in tons for complete equalization when the Ilgner-Ward Leonard system is used, also hoisting efficiency. The duty imposed on a mine-hoist motor fluctuates between wide limits during a cycle of operation. The demand on the motor may vary from a large horsepower value near the beginning of the cycle to a small or a negative demand near the end of the cycle. Such factors as tonnage to be raised per hour, depth of shaft, and balanced or unbalanced operation have such a marked effect on the operating cycle and vary to such an extent for different mines that no set of rules can be formulated for quickly determining the size of the motor required. The duty cycle from which the motor rating is derived must be determined for each case according to the conditions of the proposed hoisting operations. The motor torque required at the cardinal points of the duty cycle is determined from the moment or torque in pound-feet necessary to raise and lower the load and ropes, accelerate and retard the load, ropes, drum, etc. The torque diagram depends largely upon the type of drum. Two different types of drums, if used on exactly the same duty cycles, may have widely different peak demands and capacities of hoist motors. The essential feature in the design of hoisting drums and in determining

Jan 1, 1937

By A. Weiss

Operations Research developments within the mining industry have evolved over the last few years through stages typical of any new technique. By 1965, most companies had passed from mere recognition of the usefulness of operations research techniques to cautious acceptance of and, in some in- stances, to confident utilization of this relatively new tool. This is reflected in the growing interest and increasing participation of mining companies in sponsoring individual papers on the subject published in mining journals or presented in full sessions at annual meetings. Chief among the sources which served, in 1965, as a barometer for measuring operations research progress within mining is the Annual Symposium on the Applications of Statistics, Operations Research, and Computers in the Mining Industry, which is now sponsored by the University of Arizona, Stanford University, Colorado School of Mines, The Pennsylvania State University, and the Society of Mining Engineers of the AIME. Two significant trends were evident in the applications of operations research techniques to problems typical of the mining industry. On the one hand, unique and imaginative O.R. applications were being developed by several companies, government agencies, and at the university level. These techniques were usually developed to solve specific existing problems. On the other hand, there was unmistakable evidence of attempts to look for problems which could be solved with those "operations research" approaches which would involve the use of the digital computer. In the latter instance, solutions may become applications or demonstrations of how the computer can be used in the reduction and interpretation of data, without the benefit of a decision model. This is a common tendency as most companies have computers, at least in their accounting sections, and many operations research solutions require the use of a digital computer. From the applications observed in 1965 and in prior years, the successful techniques have evolved.

Jan 2, 1966

By A. Dorsey Lyon

INTRODUCTION A GREAT DEAL of attention has recently been given to the metallurgy of zinc, and much of that which has been said and written on this subject has been in the nature of a criticism. For instance, we are quite accustomed to hearing it said that the metallurgy of zinc has stood still, that it is antiquated, out of date, etc. We are, however, inclined to agree With George C. Stone1 When he says that, although there have been no such spectacular and radical changes as Were worked in the metallurgy of copper and iron by the introduction of converters, there has nevertheless been a decided and marked progress in the metallurgy of zinc, and especially during recent years. However, it must be admitted that, due to the fact that the chemistry of zinc is so different from that of the other metals, the retort process is the only process by which zinc is produced commercially; and, moreover, that losses which occur in the zinc industry-that is, in the mining, milling, and smelting of zinc ore-are very great, for it is stated' that probably less than 50 per cent. of the zinc mined, even

Jan 7, 1914

By T. E. Purcell

(a) Moisture. - Moisture, "as fired," is specifically significant in pulverized-coal firing because of its effect upon the continuity of raw-coal feed to the pulverizer, pulverizer capacity, flame stability, and speed of ignition. The extraneous moisture content is a rough index of potential difficulties in unloading and maintaining the flow of raw coal to the pulverizer and of pulverized coal to the furnace, with the troubles increasing geometrically as the percentage increases. Sticking or other stoppage of coal flow at any point between the bunker and the furnace is especially serious with pulverized-coal firing because there is danger of immediate loss of output. Where moisture is a problem, the designer must exercise great care in the design of coal-handling and preparation apparatus.

Jan 1, 1951