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By G. H. Anderson

Cadmium-coppeR precipitate, a byproduct of the purification stage of the zinc plant, is composed mainly of zinc, cadmium and copper in varying amounts depending on the efficiency of precipitation and the cadmium and copper contents of the impure solution treated. The composition range is approximately: cadmium 10-12 pct, copper 6-8 pct, and zinc 3&35 pct. As received at the cadmium plant, the precipitate is a dark gray to black press cake produced by filter pressing the flocculent cadmium-copper precipitate formed when impure zinc solution is agitated with zinc dust. The details of the cadmium-copper precipitation are to be found in the paper "Electrolytic Zinc at Risdon, Tasmania," by W. C. Snow. Transactions AIME, 121, 501. The treatment, in addition to recovering cadmium as metal, yields a copper product as a residue for realization and recovers in solution for return to the zinc plant, the excess of zinc dust used during purification. The recovery of cadmium metal is approximately 200 tons† per year, but, with the amount of copper residue recovered, varies with the zinc production tonnage, composition of the original concentrates, degree of roasting and other operational factors. The operations involved in the recovery of the foregoing materials are carried out in a building adjacent to the purification section of the zinc plant and can best be described by dividing the process into the following stages: 1. Oxidation and Grinding of the cadmium-copper precipitate. 2. Leaching and Filtering. 3. Precipitation of cadmium. 4. Oxidation and Grinding of the precipitated cadmium. 5. Leaching oxidized cadmium precipitate and Purification of leach solution. 6. Electrolysis. 7. Melting, Casting and Packing. A feature of the process is the use of two distinct solution circuits. Spent electrolyte from the zinc plant circuit containing about 10 pct sulphuric acid is used as the primary solvent of the cadmium and zinc present in the oxidized precipitate. When the cadmium is later separated by precipitation and filtering, the filtrate, originally zinc plant spent electrolyte and now fortified in zinc, is returned to the zinc plant. The precipitated cadmium, after oxidation, is redissolved in spent electrolyte from cadmium electrolysis which solution is in closed circuit within the cadmium plant except for discards to the zinc plant circuit as mentioned in Section 5 hereinafter. (Fig 1.) Details of the various operations follow. I. Oxidation and Grinding of the Cadminm-copper Precipitate The precipitate having had a preliminary drying by compressed air in Dehne filter presses in the zinc plant is trucked to a storage platform of sufiicient length to store 3 to 4 days' production separately. After 48 hr exposure to atmosphere, oxidation is sufficient to enable all cadmium and zinc but only a portion of the copper to be dissolved in dilute sulphuric acid. The oxidized precipitate is shovelled into trucks and, after weighing and sampling, is broken in size by means of

Jan 1, 1950

By R. G. Clay, C. F. Seaman

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By Raymond Mancha

A TEST of any fan requires the determination of such data as fan pressure, air volume handled by the fan, and power input to the fan shaft. When testing operating mine fans of the centrifugal type, test methods have been employed, for the sake of convenience, that permit great latitude in detail of procedure resulting in wide variance in results by different investigators. Some controversial aspects of field fan testing are the result of past failure to define dearly the various forms of fan pressure and to a lack of understanding of the accurate procedure for the measurement thereof. Also, the anemometer for air-volume measurement, although convenient, if carelessly used and maintained is subject to a wide error range with variation in method of application and in instrument accuracy. The appearance of the axial-flow type of mine fan offers the possibility of refinements in field testing that were impractical with the centrifugal fan. First of all, the fan inlet and outlet are more sharply defined, giving loser agreement as to just where the fan begins and ends; thus there is less disagreement as to the proper location for measurements of fan pressure. The area of passage into the inlet of an axial-flow mine fan usually constricts, causing high air velocities with accompanying steadiness of flow. Thus the fan inlet offers an ideal location for accurate air-volume measurements with the Pitot tube, the proper use of which is well established by code. Thereby the controversial aspects of such measurements are eliminated. Use of the Pitot tube also facilitates a more accurate measure of fan pressure, since the individual pressures measured can be properly weighted, as will be explained later. Determination of power input to the axial-flow mine fan operating in the field is subject to the same inaccuracies as with the centrifugal fan if the axial-flow fan is also belt driven. However, because of the higher rates of rotation usual with axial-flow fans, direct-driven units are more numerous and it is unnecessary to assume a drive efficiency, thus eliminating one possible source of error. It is the purpose of this paper to define and explain the various types of pressure measurements involved in a field test of an axial-flow mine fan, and to further explain the proper use of the Pitot tube when measuring pressures and volumes. DEFINITION OF FAN PRESSURES The two fan pressures most commonly discussed are fan total and fan static pressure, each of which has the same meaning regardless of whether the fan is blowing, exhausting, or blowing and exhausting, as in the case of a fan with ducts connected to both outlet and inlet. Fan total pressure is the amount by which the total pressure of the air delivered by the fan exceeds the total pressure of the air received by the fan. This pressure is a measure of the maximum amount of the work done on the air by the fan that is available for external use.

Jan 1, 1942

By C. E. Lesher

Two processes for agglomerating fine sized ores with low temperature coke are described. One process (Orcarb) agglomerates ores with limited amounts of carbon; the other (ore-carbon pellets) pelletizes fine sized ores, using low temperature cokes as the binder. Data are presented on the products obtained when taconite, magnetite, and hematite concentrates and several titanium oxide ores were used. REDUCTION of finely divided oxide ores has long been a major metallurgical problem. Various methods of agglomerating, notably briquetting, sintering, nodulizing, and pelletizing, have been developed and are in industrial use. Carbon is the usual reducing agent for oxide ores and attempts have been made to agglomerate fine ores with coking coal in order to get the metallurgical advantages of intimate contact and increased particle sizes suitable for subsequent smelting or reduction.1-3 For the past three years, research has been in progress on two processes for combining fine sized ores with reactive carbon in the form of low temperature coke. By one process, designated Orcarb, the ore is coated with coke carbon in amounts limited to that required for reduction and the result is an appreciable but limited increase in product particle size over that of the original ore. The second process, called ore-carbon pellets, produces pellets of fine ore bound by low temperature coke. The first objective was to develop a process for combining an oxide ore with only enough carbon for its reduction, the amount being that derived from the simple reduction equations. The carbon (that amount theoretically required for the reduction of the oxides in an agglomerate with ore), calculated to CO, ranges from 10 pct for zinc calcines Or phosphate Ores to 21 pct for rutile and 26 pct for silica. The carbon required for reducing iron oxides falls between the extremes—10 and 26 pet' When the carbon in the agglomerate is low temperature coke, reduction by hydrogen may be expected to take place, thus modifying the calculated carbon require- ment. On the other hand, in practice, an excess of reducing agent over that required in theory is normally used. Screen sizes of three iron ores on which the greater part of the experimental work was performed are given in Table I. Most of the testing was done with Freeport (Renton Mine) coal, but Pittsburgh and Elkhorn bed coals were also used. The first objective was to agglomerate the ore with sufficient low temperature coke to provide carbon for reduction, i.e., in the range from 10 to 25 pct. This was accomplished by preheating the ore, mixing with pulverized coking coal, and completing the coal carbonization in a plain steel rotating retort. The pilot retort, as it is now constituted, is shown by diagram in Fig. 1 and photograph, Fig. 2. Ore is fed into the upper kiln by a screw; it is heated to 900° to 1100°F by direct flame and it is then discharged into an insulating hopper from which it goes

Jan 1, 1956

By S. F. Ravitz

In his excellent book on the Principles of Flotation, Warkl makes the following significant statement concerning the theory of flotation: Two questions of first-rate importance must be considered . . . First, it must be decided whether the sulfide minerals require a collector in order that they may float, or whether as has been claimed, they possess an inherent float-ability and need no collector. Second, it must be decided whether the xanthate type of collector can react with or be adsorbed by the unchanged sulfide minerals themselves, or whether, as many claim, reaction with the collector is dependent upon initial oxidation of the mineral surface. In the theory of flotation there are no two questions of greater importance than these. Ravitz and Porter2 presented evidence that galena can be floated without a collector, and even without any reagent, provided oxidation products are thoroughly removed from its surface. They found also that thoroughly cleaned galena floated faster when xanthate was used than when a frother alone or no reagent was used, and that, without a collector, the recovery decreased gradually upon exposure of the mineral to air. Their work, however, has met with considerable criticism.l(127).3-6 This paper is presented to answer the objections that have been raised, and to present additional data that confirin their results. The apparatus used to provide oxygen-free conditions for flotation has been described in detail by Ravitz and Porter. An atmosphere of purified nitrogen was maintained within the apparatus; all water and solutions used were freed of dissolved air, and could be transferred where desired by the use of nitrogen Pressure; nitrogen was used as the gaseous medium for flotation. The flotation cell was essentially a glass cylinder 15 cm. long and 4.6 cm. in diameter. Near the bottom was sealed a fritted glass filter, which broke up the gas into small bubbles, and 6 cm. above this filter was a side arm through which the froth overflowed. The pulp was agitated by a glass impeller, which entered the cell through a mercury seal. For each test 17 grams of galena and 75 ml. of water were placed in the cell, after which the speed of the impeller and the rate of gas flow were adjusted. Water was added gradually during the test to keep the froth level at the proper height. The mineral was prepared and cleaned in the following manner. Exceptionally Pure galena was crushed by hand on a clean bucking board, screened to —I50+ 200 mesh, and deslimed on a 200-mesh sieve with a fine spray of water. It was then boiled on a hot plate for a number of hours with several changes of the desired cleaning solution, after which both the solution and the mineral were placed in the "mineral flask" of the apparatus. The cleaning solution was drained off, and the mineral was boiled with oxygen-free water

Jan 1, 1943

By Norman Weiss

The need for more efficient utilization of labor in the metal-mining industry has been the subject of several recent editorials in the mining press, and one attractive possibility for such improvement lies in the designing of future milling plants particularly with economy of manpower in mind. We are accustomed to think of mill design as the selection and coordinated arrangement of process equipment, but if labor is to be efficiently utilized it must be given early consideration in planning the new mill. There are three good reasons for economizing in labor: (I) to reduce costs, not only in the form of wages but also in housing, hospital, welfare, and other obligations that a company accepts when a new man is added to the payroll; (2) to prepare for recurrent periods of manpower shortage, during which a well-designed plant can manage to get along with skeleton crews; (3) because a milling plant that must depend upon an inordinately large labor force for its operation and maintenance is certain to be inefficient in other respects, for many of the shortcomings in design that make necessary the excessive use of labor are generally not corrected merely by having too many men. A simple example of this is a poorly designed ore-feeding system : if this is inadequate to the extent that one man must be constantly unclogging feed chutes, it may help the situation to put two men or three at this task instead of one, but the feed rate will still be inconstant. First of the factors that affect labor saving is the location of the mill. This is often determined by transportation costs alone, and as a result the mill may be placed close to the mine regardless of the nature of the terrain. The designer must then adapt his layout to the conditions, favorable or not. This happens often at small properties, where capital investment may seem more important than operating expense. In most instances, however, a mill site carefully selected for the advantages it gives the designer in laying out the plant is the most economical in the long run, even if it involves a slight increase in cost of ore haulage. Given some freedom of choice, the design engineer may be able to find just the piece of ground that provides the proper gradient for the needs of the flowsheet, permits installation of essential yards and trackage, and perhaps supplies an excellent site for a tailing dam. Further justification for having placed the mill not too close to the mine sometimes emerges only years later as an opportunity to treat ores from other properties. When it is known from the beginning that the new mill will serve a number of scattered mines, a free choice of mill site is naturally stimulated. The most important single factor in limiting the labor requirements of a mill is its materials-handling system for bring-

Jan 1, 1947

By Gerhard Derge

The extension of metallurgical control of steelmaking processes has always made it desirable to have some quick method for determining the oxygen content of molten steel. To meet the practical demands of the steel industry, this method must also be simple enough so that the sampling can be done by the regular furnace crew. This requirement was kept in mind continually in the work to be described. An oxygen value is required for proper control, whether it be of a rimming steel or of a fully killed steel. However, two factors have discouraged the development of such direct methods: i. The reaction C + FeO Fe + CO is rapid enough to make the correct sampling Of the bath a difficult problem. The most generally accepted means of circumventing this difficulty have been the aluminum-killed bomb test of McCutcheon and Rautiol and the aluminum-killed spoon test of Herty.2 The theory of these tests is that all of the oxygen in the iron is converted to Al2O3, which can then be determined by chemical methods. Even when simplified so that the analysis is made by a measurement of turbidity, this method is too slow to be used in routine control. Furthermore, the results obtained have never been entirely satisfactory, and some of the reasons for this will be discussed later in this paper. z. The ordinary chemical methods of analysis for oxygen in steel are not reliable. The only analysis that seems to possess sufficient dependability to be of value is vacuum fusion,3 and the apparatus and technique developed for this have been too complex and slow for routine work. Indirect Methods Indirect methods of estimating the oxygen have been resorted to in mill practice. Generally these depend upon the chemical equilibria involved. The oxygen in the bath is related to other components in the slag or metal that can be analyzed the readily. These methods have many difficulties in common. Steelmaking processes involve a complicated set of interrelated reactions, all of which tend to approach equilibrium at varying rates. The degree to which any particular equilibrium is approached is not well established. Moreover, a proper use of theoretical equilibrium constants requires a more extensive knowledge of the activities of the con- than we now have. This is well illustrated by the irregularities observed when trying to determine the oxygen in the metal by its ratio to the iron oxide in the slag. Equiliblium relations may be used mole readily if they are evaluated by plant experience, but the extent of approach to equilibrium remains an undetermined variable and lack of an entirely satisfactory mean, of determining oxygen for the calibration is felt keenly, For example, the statistical analysis of the relation between carbon and oxygen4.6 illustrates that even iia a single plant wide variations exist under llormal conditions of operation. Never. theless, this is one of the best methods now

Jan 1, 1943

By R. W. Gurry

When a steel in the austenitic state, with all its carbon in solution, is maintained, at constant temperature, in contact with a gas that removes the carbon from the surface, yet without otherwise altering the composition of the steel, the change in weight of the specimen, after any period of exposure, is a direct measure of the amount of carbon it has then lost. The purpose of the present paper is to show how, by application of established mathematical reasoning based on the fundamental diffusion law and on certain postulates, this loss of weight yields a complete picture of the distribution of carbon throughout a specimen of regular shape after any period of exposure, with all the accuracy required in any practical occurrence of decarburization. Precisely the same is true for the gain in weight during the reverse process of carburization, or for the gradual gain or loss of any element, so long as the steel comprises only a single phase at temperature, provided always that the experimental conditions conform to the postulates. The change in carbon content from surface to center of the specimen is always continuous, when everything is in solution, even though under the microscope the decarburized zone (or carburized rim) may appear to be sharply defined. The experimental procedure is described and some typical results are presented, which are then interpreted by application of the diffusion law. Experimental Procedure Initially an endeavor was made to follow the gain of weight during the process of carburization in a hydrogen-toluene atmosphere, but it was soon found that the specimen became coated with graphite, which could not be removed without some loss of metal; therefore it was decided to continue carburization until the steel was saturated throughout, to remove all graphite from the surface of the specimen, and then to follow the loss of weight during decarburization in a moist hydrogen atmosphere at the same temperature as that used for carburizing. The carburizing procedure has been described in an earlier paper;' it suffices to state here that the carburization lasted about I40 hr. at 760°C. (1760°F.) and 44 hr. at iiio°C. (2030°F.), these periods being long enough-—on the basis of the calculations presented later—to ensure that the specimen was about 99.9 per cent saturated at each temperature. The specimen was a cylinder about 3.8 cm. long and 0.475 cm. in diameter; the materials were carbonyl iron, a commercial low-metalloid iron, and a steel with 3.6 per cent nickel and the usual proportion of other elements. The fully carburized specimen, after cooling at the top of the furnace tube, was placed in a lathe and sandpapered to remove all graphite; then the diameter was measured with a micrometer caliper, and the initial weighing (corresponding to complete saturation) was made. By means of a

Jan 1, 1942

By R. W. Gurry

When a steel in the austenitic state, with all its carbon in solution, is maintained, at constant temperature, in contact with a gas that removes the carbon from the surface, yet without otherwise altering the composition of the steel, the change in weight of the specimen, after any period of exposure, is a direct measure of the amount of carbon it has then lost. The purpose of the present paper is to show how, by application of established mathematical reasoning based on the fundamental diffusion law and on certain postulates, this loss of weight yields a complete picture of the distribution of carbon throughout a specimen of regular shape after any period of exposure, with all the accuracy required in any practical occurrence of decarburization. Precisely the same is true for the gain in weight during the reverse process of carburization, or for the gradual gain or loss of any element, so long as the steel comprises only a single phase at temperature, provided always that the experimental conditions conform to the postulates. The change in carbon content from surface to center of the specimen is always continuous, when everything is in solution, even though under the microscope the decarburized zone (or carburized rim) may appear to be sharply defined. The experimental procedure is described and some typical results are presented, which are then interpreted by application of the diffusion law. Experimental Procedure Initially an endeavor was made to follow the gain of weight during the process of carburization in a hydrogen-toluene atmosphere, but it was soon found that the specimen became coated with graphite, which could not be removed without some loss of metal; therefore it was decided to continue carburization until the steel was saturated throughout, to remove all graphite from the surface of the specimen, and then to follow the loss of weight during decarburization in a moist hydrogen atmosphere at the same temperature as that used for carburizing. The carburizing procedure has been described in an earlier paper;' it suffices to state here that the carburization lasted about I40 hr. at 760°C. (1760°F.) and 44 hr. at iiio°C. (2030°F.), these periods being long enough-—on the basis of the calculations presented later—to ensure that the specimen was about 99.9 per cent saturated at each temperature. The specimen was a cylinder about 3.8 cm. long and 0.475 cm. in diameter; the materials were carbonyl iron, a commercial low-metalloid iron, and a steel with 3.6 per cent nickel and the usual proportion of other elements. The fully carburized specimen, after cooling at the top of the furnace tube, was placed in a lathe and sandpapered to remove all graphite; then the diameter was measured with a micrometer caliper, and the initial weighing (corresponding to complete saturation) was made. By means of a

Jan 1, 1942

By Norman Weiss

THE need for more efficient utilization of labor in the metal-mining industry has been the subject of several recent editorials in the mining press, and one attractive possibility for such improvement lies in the designing of future milling plants particularly with economy of man-power in mind. We are accustomed to think of mill design as the selection and coordinated arrangement of process equipment, but if labor is to be efficiently utilized it must be given early consideration in planning the new mill. There are three good reasons for economizing in labor: (I) to reduce costs, not only in the form of wages but also in housing, hospital, welfare, and other obligations that a company accepts when a new man is added to the payroll; (2) to prepare for recurrent periods of manpower shortage, during which a well-designed plant can manage to get along with skeleton crews; (3) because a milling plant that must depend upon an inordinately large labor force for its operation and maintenance is certain to be inefficient in other respects, for many of the shortcomings in design that make necessary the excessive use of labor are generally not corrected merely by having too many men. A simple example of this is a poorly designed ore-feeding system: if this is inadequate to the extent that one man must be constantly unclogging feed chutes, it may help the situation to put two men or three at this task instead of one, but the feed rate will still be inconstant. First of the factors that affect labor saving is the location of the mill. This is often determined by transportation costs alone, and as a result the mill may be placed close to the mine regardless of the nature of the terrain. The designer must then adapt his layout to the conditions, favorable or not. This happens often at small properties, where capital investment may seem more important than operating expense. In most instances, however, a mill site carefully selected for the advantages it gives the designer in laying out the plant is the most economical in the long run, even if it involves a slight increase in cost of ore haulage. Given some freedom of choice, the design engineer may be able to find just the piece of ground that provides the proper gradient for the needs of the flowsheet, permits installation of essential yards and trackage, and perhaps supplies an excellent site for a tailing dam. Further justification for having placed the mill not too close to the mine sometimes emerges only years later as an opportunity to treat ores from other properties. When it is known from the beginning that the new mill will serve a number of scattered mines, a free choice of mill site is naturally stimulated. The most important single factor in limiting the labor requirements of a mill is its materials-handling system for bring-

Jan 1, 1946

By Norman Weiss

The need for more efficient utilization of labor in the metal-mining industry has been the subject of several recent editorials in the mining press, and one attractive possibility for such improvement lies in the designing of future milling plants particularly with economy of manpower in mind. We are accustomed to think of mill design as the selection and coordinated arrangement of process equipment, but if labor is to be efficiently utilized it must be given early consideration in planning the new mill. There are three good reasons for economizing in labor: (I) to reduce costs, not only in the form of wages but also in housing, hospital, welfare, and other obligations that a company accepts when a new man is added to the payroll; (2) to prepare for recurrent periods of manpower shortage, during which a well-designed plant can manage to get along with skeleton crews; (3) because a milling plant that must depend upon an inordinately large labor force for its operation and maintenance is certain to be inefficient in other respects, for many of the shortcomings in design that make necessary the excessive use of labor are generally not corrected merely by having too many men. A simple example of this is a poorly designed ore-feeding system : if this is inadequate to the extent that one man must be constantly unclogging feed chutes, it may help the situation to put two men or three at this task instead of one, but the feed rate will still be inconstant. First of the factors that affect labor saving is the location of the mill. This is often determined by transportation costs alone, and as a result the mill may be placed close to the mine regardless of the nature of the terrain. The designer must then adapt his layout to the conditions, favorable or not. This happens often at small properties, where capital investment may seem more important than operating expense. In most instances, however, a mill site carefully selected for the advantages it gives the designer in laying out the plant is the most economical in the long run, even if it involves a slight increase in cost of ore haulage. Given some freedom of choice, the design engineer may be able to find just the piece of ground that provides the proper gradient for the needs of the flowsheet, permits installation of essential yards and trackage, and perhaps supplies an excellent site for a tailing dam. Further justification for having placed the mill not too close to the mine sometimes emerges only years later as an opportunity to treat ores from other properties. When it is known from the beginning that the new mill will serve a number of scattered mines, a free choice of mill site is naturally stimulated. The most important single factor in limiting the labor requirements of a mill is its materials-handling system for bring-

Jan 1, 1947

By William B. Senseman

For reasons well known to mining engineers, wet grinding is quite universal in plants having to do with the extraction of metallic values from crude ores. In the processing of the nonmetallic and industrial minerals, dry grinding is the more common method. The crude materials usually contain less moisture than is required for wet grinding and the products usually are marketed in dry form; most of them as powders. To grind wet, therefore, would entail the addition of water and an expense for drying the product, which is not justified. Furthermore, the material would be a dried cake that would still have to be pulverized or disintegrated before use. For producing fine products, air separators have largely replaced the earlier practice of sizing with screens and bolting reels. Mills in which the air-separating feature is an integral part of the design are now commonplace. Mill designs are many and air systems have developed along various lines, but for the purpose of this discussion, only the fundamental idea need be emphasized— the idea of using air for sizing and removing pulverized material from a grinding mill. About 20 years ago, when the use of mills equipped with air separation had become common practice, heated air or gas was introduced to such a system for the removal of moisture. The purpose of this discussion is to outline the developments and trends that have followed this experiment. It is an important development, for mills equipped with air separation are now universally used for pulverizing not only nonmetallic and industrial minerals, but a great variety of other materials. All such mills are essentially the same. Dried material is pulverized, removed by an air stream, trapped in a cyclone collector, and relatively clean air from the cyclone is returned through a fan to the mill. There is no theoretical need for a vent from the system. The practical need is to vent the air that leaks into the system, and, since the cyclone is not 100 per cent efficient as a dust collector, this vented air carries some dust with it. The initial attempt at drying was made on a system like that described, by the addition of a furnace. The minus pressure within the mill draws hot gas from the furnace. Infiltration air, along with the drying gases and evaporated moisture, is vented as before. The first application was made to a mill used for pulverizing coal. The coal contained very little moisture, so very little drying effect was needed for the success of this experiment. Actually, some calculated efficiencies of over 100 per cent were obtained before due allowance was made for the conversion of mill power into useful heat. When moisture is low, the heat obtained from the conversion of mill power is an appreciable part of that required. That drying can be effected in the short time interval available in a grinding mill of this type is better understood by visual-

Jan 1, 1947

By W. C. Muehlhof, L. D. Lamont

THROUGHOUT the anthracite industry's history, the problem of handling and dis¬posing of refuse material has been one of major importance. In the early days of the industry's activities, only the large sizes of anthracite were satisfactory for market requirements and the smaller sizes were deposited as refuse. As time went by, more efficient coal-burning equipment was developed and a market demand was created for the finer sizes of coal. This demand made it possible to recover market¬able coal from banks deposited as refuse in the earlier days of mining. The same condition holds good today and we can see a definite trend toward the objective of utilizing the No. 4 and No. 5 buckwheat sizes. Under present conditions, we know that a limited market exists for No. 4 and 5 buckwheat; we know that the finer sizes of anthracite can be burned as pulverized fuel; we know that these fine sizes can be processed and sold as bri¬quettes, and we also know that the heat content of these fine sizes of coal is compar¬able with that of the sizes now merchan¬dized. Therefore, it should be the objective of the industry to conserve these fine sizes of coal in such manner that if and when processing facilities are developed, and market demands exist, the material can be salvaged and made available to meet these demands. From the standpoint of economical future reclamation, the ideal method of disposing of refuse containing the finer sizes of coal is to deposit it in banks separate and apart from other breaker refuse. By this procedure, the minimum amount of material will be processed in future salvage operations to recover the No. 4 and No. 5 buckwheat sizes. Many collieries in the anthracite region are operated under leases that have a specific term. These lease agreements provide that, upon expiration of the lease by term, exhaustion or otherwise, the ownership of the refuse material reverts to the lessor. Therefore, a question was raised in the minds of these operators: Why should I assume the expense of such pro¬cedure when I have no assurance that I will benefit by the future processing and sale of these fine sizes? Under present conditions, the answer to this question is not one of choice but of necessity, as restrictions imposed by legislation make it mandatory on operators of preparation plants to take necessary steps to properly treat silt-laden waters before permitting them to enter the streams. By reason of these restrictions, the disposal of fine sizes becomes a secondary_ rather than a primary operating con sideration, and the procedure for disposing of this refuse becomes a matter of choice governed principally by topographical and geographical conditions. METHODS OF HANDLING REFUSE There are three principal methods of handling and disposing of fine-size refuse; namely :

Jan 1, 1946

By Junius D. Edwards

WHEN Charles Martin Hall invented the electrolytic process for the production of aluminum, one basic requirement was a supply of pure alumina. Now, more than 50 years later, the same requirement still holds, and there are no technical straws in the wind to indicate that this situation is likely to change. Long ago, Aluminum Company of America took a number of steps to see that this country should not face shortages of alumina and aluminum in time of emergency. As one step to this end, the Company developed and used a supply of bauxite in South America in order to conserve dwindling Arkansas reserves. Another, and more a program than a step, was to investigate every type of process that might bear the possibility of utilizing the low-grade ores of aluminum that are more plentiful than bauxite. It is a long story, but the investigation included electric-furnace processes, acid extraction processes and alkaline extraction processes, with or without preliminary furnace and sinter treatments. Some of the processes developed were operated in commercial production. This work has been continuous since the first World War. The turn of the wheel came sooner than was expected when, in 1942, bauxite shipments through the Caribbean became uncertain and hazardous through the operation of submarines. The tremendous requirements of aluminum for the war effort were consuming high-grade bauxite at an alarming rate and a national search for substitutes began. High-grade bauxites, for technical and economic reasons, have been defined as those containing about 7 per cent or less of silica. While bauxite of this grade was limited in amount in the United States, there was in sight a substantially greater volume of low-grade, high-silica bauxite not amenable to economic treatment by the Bayer process, but richer in alumina than the clays or other available low-grade ores. It was at this point, and after several years of intensive research on this particular phase of the problem, that Aluminum Research Laboratories brought to completion the invention of one of its research staff, which made practical the utilization of large reserves of low-grade bauxite in America. BAYER PROCESS The Bayer process has been almost universally used for the 'production of alumina from bauxite. This process, in outline, comprises the digestion of finely ground bauxite in a solution of caustic soda (NaOH) under pressure and at an elevated temperature. Alumina in the bauxite goes into solution as sodium aluminate and the impurities-oxides of iron, titanium and silicon-remain undissolved. The undissolved residue, known as "red mud," from its color, is separated and run to waste. Aluminum hydrate (A1202.3H20) is precipitated from the solution of sodium aluminate, and the liquor containing the caustic soda and some alumina is returned to the process for digesting more bauxite. The aluminum hydrate, after washing, is calcined at a high temperature to produce

Jan 1, 1945

By R. G. Clay, C. F. Seaman

THE mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that time, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered, the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. ORIGINAL EQUIPMENT AND INSTALLATION A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots and the up position would connect it to the line from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1942

By A. W. Gauger

IT is with some trepidation that I approach my subject, for I know that I shall at once incur the suspicion of the mechanical engineer, with his concern for boiler tests and efficiencies; of the mining engineer, with his mechanization problems; and of the chemical engineer with his unit processes. Nevertheless, it must be recognized that the technology of fuels, the systematic body of knowledge relating to the production, processing and utilization of mineral fuels, has grown to such proportions and involves so many widely diverse applications of fundamental science and engineering that it behooves us to stop and consider just what its status is in relation to other technologies, as well as in relation to the present and future needs of the social order. The time has come when technologists, applied scientists and engineers must consider the possible effects that their achievements may have 0n Society. The development of a machine to replace several hundred men is a fine achievement when considered solely from the point of view of creative endeavor. But what of the men who are displaced by the machine! Does not the responsibility for the development of this machine embrace as well the consequences of obsolescence and technological unemployment resulting from such development? I believe that it does and that we must recognize this fact in our practice of engineering and technology as well as in the training of future engineers and technologists. MAGNITUDE OF THE FUELS INDUSTRY Fuel technology is the body of knowledge that relates to the mineral fuels. The fuels industry is by far the most important of all our mineral industries. The chart in Fig. I, reprinted from U. S. Bureau of Mines Information Circular No. 6643, indicates the comparative values of metals, fuels and other nonmetallic minerals for the year 1929. Of the grand total of 13 1/4 billions of dollars in value of minerals at the mines, the mineral fuels contributed 48 per cent. Coal, with 34.4 per cent, exceeded the value of any other single mineral. In fact, the value of the production of the Pennsylvania anthracite mines alone annually exceeds the value of the gold production of the entire world. Petroleum and natural gas in 1020 contributed 14.3 per cent and 3.3 per cent, respectively, of the total value of world mineral production. So much for the world picture; let us now look at the situation in the United States. Over the 60-year period 1880 to 1939, $46,409,727,000 of metallic products, S80,788,852,000 of fuels (coal, petroleum, natural gas, natural gasoline) and $27,871,000,000 of nonmetallic minerals other than fuels were produced. The mineral fuels have contributed 52.7 per cent of the entire mineral wealth produced over this period that has seen such tremendous industrial development in our country. As a matter of fact, our present industrial civilization has been made possible through the use of the

Jan 1, 1942

By M. C. Flemings, R. Mehrabian, G. E. Nereo

Analytical expressions derived previously are used to describe quantitatively effects on macrosegregation of some solidification and mold design variables. Al-4.5 pct Cu alloy is used as example. It is shown that abrupt changes in heat transfer rate result in solute-rich or solute-poor "bands". Severe solute-rich bands accowzpany "reheating" that can result, for example, from abrupt formation of an air gap at a mold-metal interface. With planar isotherms, solute-poor regions also result from reduction in cross section through which fluid flow to feed shrinkage takes place. It is suggested that the negative "cone" of segregation often found at the base of large ingots results from this reduction in cross section for feeding. Consideration is given to the positive and negative segregation to be expected in laboratory ingots in association with changes in ingot cross section. Under-riser positive segregation is considered and it is concluded this segregation results when vertical temperature gradient is low in the region of the riser; the segregation is accentuated by the temperature gradient reversal which can occur in inadequately insulated risers ("'hot tops"). Finally it is shown for Al-4.5 pct Cu alloy that the assumption of constant k and p in analytic solutions gives calculated results close to those obtained for the case of ziariable k and 0. A basic "local solute redistribution equation" was derived earlier' for calculation of macrosegregation resulting from flow of solute-rich liquid to feed solidification shrinkage, thermal contraction, and/or other contractions occurring during solidification. This expression relates fraction liquid, g~, to liquid composition, CL, at a given position x, y, z, in a solidifying Eq. [I] is for the general case of three-dimensional heat and fluid flow assuming constant solid (but not necessarily liquid) density during solidification, negligible net solute change from diffusion, and no pore formation. In this paper we first demonstrate application of this equation for calculation of macrosegregation in several examples involving one-dimensional heat and fluid flow. Specifically considered are examples in which segregation results from sudden changes in heat flow conditions. Next, Eq. [I] is used to describe segregation in examples where isotherms remain planar but where fluid flow is no longer one-dimensional. Application of this expression illustrates the type of segregation to be expected in unidirectionally solidified ingots of gradually or abruptly varying cross section. For segregation calculations, A1-4.5 pct Cu alloy is employed with constant k and p Vz = 0.172, 4 = 0.055), eutectic composition (CE = 33 pct Cu). In all examples, fraction eutectic is neglected in velocity calculations; this was previously shown to introduce little error.' A linear distribution of fraction liquid in the liquid-solid zone is assumed. In a concluding section it is shown that the assumption employed of constant k and /3 gives results close to those obtained when the actual (variable) values of k and 0 are employed.

Jan 1, 1969

By F. Musuka, D. Jenkins, A. Keyes

BHP Olympic Dam operates a Direct-to-Blister copper flash smelting Furnace (DBF) and an Electric Slag Cleaning Furnace (ESCF). Tapping blister copper from these furnaces via tapholes is a routine task carried out by trained personnel. Blister tapping is a high-risk activity with personnel operating in close proximity to molten metal. BHP’s risk management framework has identified taphole integrity as key to managing the potential for loss of molten metal containment and molten metal – water steam explosion, both of which have the potential to cause multiple fatalities. As part of the focus on managing this risk, since 2017, a dedicated Taphole Development Team has been embedded in the operation to run an accelerated improvement program focused on safely and sustainably embedding an improved taphole design with greater integrity. The program of works includes selecting and testing of new materials as well as improving operating discipline by collaborating with tapping personnel to convert their knowledge into codified condition monitoring and operating strategies. Computational fluid dynamics (CFD) is being used to derive critical risk controls, and to develop live-alerts to infer excessive refractory insert wear. To date, improving DBF blister taphole mudgun clay has led to more consistent drilling depths and subsequently reduced oxygen lancing requirements, resulting in improvements in taphole insert service life. On the ESCF, insert performance testing and development of a new taphole condition monitoring strategy, is being used to increase taphole insert tonnages to more sustainable levels. Operations personnel have been closely involved in different phases of the change management, this has led to standardization of tapping practices for both furnaces, and improved ownership of the proposed changes. The paper discusses the maturity in taphole integrity risk management as a result of recent and ongoing developments. It then provides an overview of future taphole developments proposed at Olympic Dam. INTRODUCTION Tapping operations at Olympic Dam and learning from significant events The Direct-to-Blister furnace (DBF) at Olympic Dam was built with a conventional design blister taphole, which was designed for manual oxygen lancing and clay stopper bar plugging tapping operations (Campbell, 2005). Manual tapping is a high-risk activity, where personnel work in close proximity to molten metal, and molten metal personal protective clothing and equipment must be used. Opening tapholes by oxygen lancing is a skilled activity and unintentional damage to the taphole inserts can occur. Since 2008, a Mudgun and Drill unit has been developed and trialled and is now in use on the DBF blister tapholes, to remove personnel from the line of fire when opening and closing tapholes. This has the additional benefit of reducing lancing damage to taphole inserts. A new taphole design labelled the Modular Blister Taphole (MBTH) was developed at Olympic Dam for the Mudgun and Drill operation. The intent of the design was for mudgun clay to fill the entire tapping channel when plugging after a tap. The drill would then drill through this clay plug to open the taphole.

Jan 1, 2019

By A. Keil

TO measure the absolute conductivity of metallic specimens it is generally necessary to have exact geometrical shapes and good electrical contact with the test-piece. To overcome limitations of the standard methods, an electronic conductivity-meter* Trade name, Sigmatest; model Dr. F. Forster; Reutlingen, Ger-many. has been developed, which works on the principle of inducing high frequency eddy-currents in the specimens and measuring the reaction of these curients on a measuring coil.',' To attach the instrument to the test specimen, a flat area of about 10 mm diam is all that is necessary. The reading is taken off a scale calibrated in pet of International Annealed Copper Standard. The induced eddy-currents are propagated in a circular motion parallel to the surface. Thickness of the layer included in a measurement depends on the frequency used. With a frequency of 60 kc per sec, the penetration of the currents was 0.8 mm at the upper limit of the scale, and 2 mm at the lower limit. The method is limited to nonferrous metals and other applications have been described elsewhere." The object of the present investigation was to detect the formation and growth of surface layers of differing conductivity, and to determine the depth to which these changes could be observed. A well known phenomenon which illustrates the problem is the internal oxidation of certain silver alloys containing base metal solutes, such as copper or cadmium, when they are annealed in an oxidizing atmosphere.4, 5 Until now, the only methods available for following these phenomenon have been metallo-graphic examination and hardness tests.0, 7 Changes in mechanical properties of copper alloys by internal oxidation were recently discussed in a more general way by Martin and Smith." When measuring these conductivity changes it must be remembered that the only absolute values of conductivity are those of 1) the nontreated homogeneous specimen, and 2) the final values obtained with a surface layer of depth comparable to the penetration of the eddy-currents. All intermediate values refer to the mixed conductivities of the transformed surface layer and the untransformed alloy, and are thus readings of relative value only. Experimental Results Specimens were annealed for 1/2 hr at 800°C in hydrogen, and the internal oxidation was carried out by thermal treatment in air at 800°C. Fig. 1 shows the results of the first series of experiments. With internal oxidation of a solid solution of 8 pet Cd in silver, the conductivity rises continuously until the surface layer formed has become thick enough to be the deciding factor (about 400µ). Precipitates of CdO are mainly in the form of points, Fig. 2, and, as a result, the electric contact between the silver crystals is not distorted by them. Brine11 tests were run with a 1.25 mm ball and a 7.812 kg load. The specimen annealed in hydrogen and oxidized to a depth of 400µ had a hardness of 40 Bhn,

Jan 1, 1959

By Gerhard Derge

The extension of metallurgical control of steelmaking processes has always made it desirable to have some quick method for determining the oxygen content of molten steel. To meet the practical demands of the steel industry, this method must also be simple enough so that the sampling can be done by the regular furnace crew. This requirement was kept in mind continually in the work to be described. An oxygen value is required for proper control, whether it be of a rimming steel or of a fully killed steel. However, two factors have discouraged the development of such direct methods: i. The reaction C + FeO Fe + CO is rapid enough to make the correct sampling Of the bath a difficult problem. The most generally accepted means of circumventing this difficulty have been the aluminum-killed bomb test of McCutcheon and Rautiol and the aluminum-killed spoon test of Herty.2 The theory of these tests is that all of the oxygen in the iron is converted to Al2O3, which can then be determined by chemical methods. Even when simplified so that the analysis is made by a measurement of turbidity, this method is too slow to be used in routine control. Furthermore, the results obtained have never been entirely satisfactory, and some of the reasons for this will be discussed later in this paper. z. The ordinary chemical methods of analysis for oxygen in steel are not reliable. The only analysis that seems to possess sufficient dependability to be of value is vacuum fusion,3 and the apparatus and technique developed for this have been too complex and slow for routine work. Indirect Methods Indirect methods of estimating the oxygen have been resorted to in mill practice. Generally these depend upon the chemical equilibria involved. The oxygen in the bath is related to other components in the slag or metal that can be analyzed the readily. These methods have many difficulties in common. Steelmaking processes involve a complicated set of interrelated reactions, all of which tend to approach equilibrium at varying rates. The degree to which any particular equilibrium is approached is not well established. Moreover, a proper use of theoretical equilibrium constants requires a more extensive knowledge of the activities of the con- than we now have. This is well illustrated by the irregularities observed when trying to determine the oxygen in the metal by its ratio to the iron oxide in the slag. Equiliblium relations may be used mole readily if they are evaluated by plant experience, but the extent of approach to equilibrium remains an undetermined variable and lack of an entirely satisfactory mean, of determining oxygen for the calibration is felt keenly, For example, the statistical analysis of the relation between carbon and oxygen4.6 illustrates that even iia a single plant wide variations exist under llormal conditions of operation. Never. theless, this is one of the best methods now

Jan 1, 1943