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By W. I. Garms, J. L. Stevens

Because of the many conflicting opinions regarding the functioning of the ball load and the characteristics of ball wear in ball-mill grinding, the following résumé of reports covering recent tests conducted at the Hayden plant of the Kennecott Copper Corporation is submitted with the view of pointing out certain salient features that should throw additional light on the subject. Mills Used Cylindrical drum mills with trunnion feed and discharge were used. The discharge trunnion was 13½ in. in diameter. Dimensions inside the lining were 6 ft. 6 in. in diameter and 9 ft. 6 in. long. The mills were Britannia-lined and nine lifter bars projected 2 in. from the lining. The mill speed was 162/3 r.p.m. Feed to the mill was 9 per cent on 10-mesh, 85 per cent on 100-mesh, 77 per cent solids. New balls measured 2 in. in diameter and the approximate weight of the ball load was 40,000 pounds. Initial Test During a test that was being conducted to determine the relative economic value of standard cast balls vs. forged steel balls in the local 'grinding operation, certain observations were made in regard to the functioning of the various ball sizes and the characteristics of ball wear, which warranted a supplementary test for the purpose of confirming such observations. For the initial test, the necessity for developing a graduated load of the new forged balls before the test could be started was realized, but the time required to obtain a graduated load and the functioning of the various ball sizes was not clear. No advantage could be seen for starting out with a full load of forged balls, as this initial load, at the least, would have to be completely ground out and replaced by daily additions of balls before the graduation could be complete and a normal static condition of ball sizes reached. For a considerable period prior to exhaustion of this initial charge there would be an unusual quantity of worn balls of a certain size, which would not be typical of a normal charge and could influence the characteristics and functioning of the entire load. Therefore, it was decided to start by daily addition of the forged balls to a mill that contained a normal operating load of cast balls. After 88 days of operation the mill was shut down for inspection. By that time the load consisted of a mass of forged balls functioning around a large core of the residual material of the former cast-ball load. This core of cast slugs, locally referred to as "the kidney," was well defined from the forged-ball load.

Jan 1, 1947

By W. I. Garms, J. L. Stevens

Because of the many conflicting opinions regarding the functioning of the ball load and the characteristics of ball wear in ball-mill grinding, the following résumé of reports covering recent tests conducted at the Hayden plant of the Kennecott Copper Corporation is submitted with the view of pointing out certain salient features that should throw additional light on the subject. Mills Used Cylindrical drum mills with trunnion feed and discharge were used. The discharge trunnion was 13½ in. in diameter. Dimensions inside the lining were 6 ft. 6 in. in diameter and 9 ft. 6 in. long. The mills were Britannia-lined and nine lifter bars projected 2 in. from the lining. The mill speed was 162/3 r.p.m. Feed to the mill was 9 per cent on 10-mesh, 85 per cent on 100-mesh, 77 per cent solids. New balls measured 2 in. in diameter and the approximate weight of the ball load was 40,000 pounds. Initial Test During a test that was being conducted to determine the relative economic value of standard cast balls vs. forged steel balls in the local 'grinding operation, certain observations were made in regard to the functioning of the various ball sizes and the characteristics of ball wear, which warranted a supplementary test for the purpose of confirming such observations. For the initial test, the necessity for developing a graduated load of the new forged balls before the test could be started was realized, but the time required to obtain a graduated load and the functioning of the various ball sizes was not clear. No advantage could be seen for starting out with a full load of forged balls, as this initial load, at the least, would have to be completely ground out and replaced by daily additions of balls before the graduation could be complete and a normal static condition of ball sizes reached. For a considerable period prior to exhaustion of this initial charge there would be an unusual quantity of worn balls of a certain size, which would not be typical of a normal charge and could influence the characteristics and functioning of the entire load. Therefore, it was decided to start by daily addition of the forged balls to a mill that contained a normal operating load of cast balls. After 88 days of operation the mill was shut down for inspection. By that time the load consisted of a mass of forged balls functioning around a large core of the residual material of the former cast-ball load. This core of cast slugs, locally referred to as "the kidney," was well defined from the forged-ball load.

Jan 1, 1947

By H. R. Stahl

The disseminated lead district of Southeast Missouri lies 70 miles south of St. Louis. The only metal of economic importance in the ore is lead, but minor amounts occur of iron, zinc, copper, cobalt, nickel, silver and cadmium. These minor metals, as might be expected, tend to concentrate in both the flotation and table middlings. In the past, some zinc has been recovered by differential flotation. The silver and cadmium tend to associate with the zinc, and some recovery is made of these two metals at the smelter. The lead occurs as galena, or lead sulphide. The minor metals are also present as sulphides. The presence of oxidized lead is suspected, but has not been definitely proved. The gangue is chiefly dolomitic limestone, with small amounts of sandstone and shaly glauconitic and chloritic minerals. The ore contains about 3.25 per cent lead, occurring as galena disseminated through the gangue. Some of the galena is very finely disseminated, and the mineral in the gravity tailings, or chat; is seldom or never in the free state. The method of treatment of the ore has been evolved over a period of 75 years; hence probably is not far from correct. The milling processes have tended toward simplification during the past 15 years, but recoveries have never been sacrificed for the sake of simplicity. In many cases, simplification of flowsheets has led to better recoveries and operating economies. A simplified flowsheet is shown in Fig. I. The St. Joseph Lead Co. operates four mills in the Flat River district, with a normal total capacity of about 21,000 tons per day, divided as follows: Federal mill, 11,000 tons; Leadwood mill, 4500 tons; Desloge mill, 3400 tons and Bonne Terre mill 2400 tons. The methods of concentration used at these plants are similar in principle; they differ mainly in details and in the type and size of the equipment used. The process of concentration may be divided into five main steps; i.e., dry crushing, wet grinding, classification with its attendant desliming, tabling and flotation. A brief description of these processes will be given, noting important differences in the various plants, and also noting the more important changes or developments of the past few years. Dry Crushing The run-of-mine ore is hoisted in skips, the capacity of these skips ranging from 23/4 tons to 81/4 tons. At the Desloge plant, the ore is hoisted in 2-ton mine cars on self-dumping cages. All ore is hoisted at shafts adjacent to the various crushing plants, no ore being transported on the surface. Primary breaking at all the mills is accomplished with 9E Telsmith breakers. Preliminary screening is provided ahead of most of these crushers, either by station-

Jan 1, 1943

By C. B. Smith, A. G. Guy

The power and convenience of tensor analysis are employed in deriving the equations that describe three-dimensional diffusion in an n-component system plus vacancies. A "Kirkendall coordinate system, '' defined in principle by Kirkendall markers, is used in defining the intrinsic diffusion coefficients that are of primary interest. The defining equations are then transformed to local cartesian coordinates and to fixed cartesian coordinates, the latter being convenient for experimental measurements. Finally, it is shown how the intrinsic diffusion coefficients can be calculated from the experimentally determined quantities. It is generally recognized that any treatment of the essential diffusion phenomena in metallic systems—the jumping of atoms from one lattice position to another—must be based on movement of the atoms past Kirkendall interfaces.' A logical extension of this idea is the proposal that, fundamentally, diffusion must be described mathematically in terms of "Kirkendall coordinates." The basic equations thus obtained may then be transformed into other coordinate systems, when such transformation leads to simplification of either the experimental or mathematical treatment of diffusion data. Tensor analysis is clearly the appropriate mathematical tool for this purpose, since it provides methods both for determining general equations that are valid for all coordinate systems and for conveniently transforming the equations from one coordinate system to another system related to the first in some known fashion. For the present purpose three coordinate systems are used, and it is convenient to begin with a description of the relations among these three systems. FICK LAW FOR KIRKENDALL COORDINATES The reference coordinate system, and the one that is useful for most experimental measurements, is a cartesian coordinate system, xi (i = 1, 2, 3), fixed with respect to some point in the specimen, such as an end of the alloy bar undergoing diffusion. A second set of coordinates, the Kirkendall coordinate system, is one that in principle can be described by the location of inert markers (Kirkendall interfaces). A convenient mathematical relation between Kirkendall coordinates, x'i, and the reference cartesian coordinates, ii, is obtained by specifying that the Kirkendall coordinate system coincide with the cartesian system at time, t = 0. After diffusion has occurred for some time, t = t, the Kirkendall coordinate system will differ from the reference system in the manner shown schematically for the one-dimensional case in Fig. 1. The third coordinate system is Darken's1 local cartesian-coordinate system, xi, which is identical with ii in orientation and units, but which moves with the point under consideration in the Kirkendall coordinate system. Even though most diffusion problems are essentially one-dimensional, there is an advantage in first carrying out the analysis of the three-dimensional problem to gain a clearer insight into the assumptions that are made in the usual one-dimen-

Jan 1, 1962

By V 7. 0 / 300 dpi

LAST fall, having a number of ore samples from mine-development work carried on in spite of the "Revolution," I was forced to do my own assaying again, after a lapse of many years. This gave me an opportunity to build a furnace for crucible assaying, using wood for fuel, along lines which had been developing in my mind since the time 20 years ago, when as general factotum at a little mining camp in western Chihuahua, Mexico, I had converted a crucible charcoal furnace into one using the reverberatory principle. Wood-burning furnaces have been described in technical literature, but I believe the present application will prove helpful to those of the profession who are working in out-of-the-way places. The furnace is shown in detail in Figs. 1, 2 and 3. Remembering the trouble caused by stoking my first charcoal furnace too heavily, I fired this one with but a few sticks of wood at a time. By this method I could run out a charge of 12 crucibles in 30 to 35 min., using 50 sticks of ordinary stove wood. Later I fill up the firebox, holding about 40 sticks, get everything ready, put the crucibles into the melting chamber and "fire up." Just as soon as the fire is well started the draft door is closed and the furnace allowed to heat up slowly for 5 min. Then the draft is opened (the door to the ash-pit) and the fire allowed to roar, which it does. In 20 min. from the time the fire is lighted the fire door is opened, the fire stirred along the grates, and 5 or 6 more sticks of wood thrown in. In 25 min. from the start, the draft is closed again, the cover of the melting chamber pushed back and the charges poured. The melts will be found to be perfect, and the crucibles to pour absolutely clean, if everything has gone right, as will usually be the case. A new charge may be put in, and fired with S to 10 sticks of wood at a tine, and be ready to pour within 20 min.; but it is just as well to leave the crucibles in the melting chamber a full 2.5 min. By using 10 sticks of wood to start the fire and adding 5 sticks as often as needed, which will be as soon as the last sticks fed in are in perfect combustion, the smelting can be completed in from 30 to 35 min., using from 35 to 40 sticks of wood. I consider this to be the best method of stoking.

Jan 10, 1914

By J. L. Lummus, Frank O. Jones

A field procedure for determining graphite in drilling mud is presented which is sensitive to 0.25 lbs/bbl and accurate to 20 per cent. The method, utilizing oil flotation principles, is suitable for use in all oil-free water base muds except those containing ligno-sulfonates, lignites, or surface active agents. INTRODUCTION Graphite was proposed as a drilling lubricant as early as 1940, but has not been used extensively until recently.' Graphite has been substituted for oil in certain instances to avoid oil contamination of core samples since it has been found that graphite does not interfere with routine core analysis. A major difficulty in using graphite is the lack of a satisfactory field test for determining the concentration of graphite in mud. DISCUSSION OF METHOD Several possible methods of analysis were investigated, including chemical separation, density fractionation, radioactive tracers, and oil flotation.', ' The investigation indicated that a satisfactory method for field use could be devised using an oil flotation method. It was found that toluene shaken with a highly diluted mud sample would recover sufficient graphite to be apparent in the oil phase. Refinement and standardization of technique made it possible to detect as little as 0.25 lb/bbl graphite and to estimate graphite content to within 20 per cent by using a visual comparison method. The test is limited to oil-free water base muds. Quebracho, phosphates, or barytes in the mud do not interfere with the test but surface active agents, lignosulfonates' and lignites evidently render !he graphite water-wet so that none may be recovered by oil-flotation. PROCEDURE Instructions for determining graphite in mud: 1. Collect representative one gallon sample.- from mud stream. Transfer immediately 350 ml portions of this Sample into four mixing cans labeled A, B, C, and D. Measure out graphite (same as is used to treat mud) for addition to cans as indicated below, using either a specially designed scoop or a balance. Can: A B C D Grams Graphite to Add: none 2 4 6 Stir samples B, C, and D for 15 minutes with propeller type mixer. 2. Clean and rinse well with water four 250 ml volumetric flasks labeled I, 11, 111, and IV. Using a 5 ml hypodermic syringe, transfer the indicated volume of samples from cans A, B, C, and D into designated flasks as follows: From Can: A B C D To Flask: I II III IV Add This Volume of Mud: 4 ml 2 ml 2 ml 2 ml 3. Measure 100 ml fresh or distilled water to each flask and swirl gently to disperse mud. Add exactly 5 ml toluene to each flask. Stopper and shake each

Jan 1, 1953

By L. D. Ricketts

Discussion of the paper of L. D. RICKETTS, presented at the San Francisco meeting, September, 1915, and printed in Bulletin No. 100, April, 1915, pp. 711 to 737. FREDERICK LAIST, Anaconda, Mont.-I am going to tell you of some of the results we have been getting in Anaconda during the past two or three months in a large plant which was finished in May. We commenced work in Anaconda a year and a half ago on the construction of a leaching plant which was to have a capacity of 2,000 tons per day. The plant was to treat tailings from the concentrator, containing about 12 1b. of copper to the ton. During the past year we changed our plans and decided to use the flotation process on the tailings from the concentrator, which contained no oxidized copper, and apply the leaching process only to tailings which had accumulated in a large dump below the works. This dump contains about 15,000,000 tons of tailings, and since some of the copper in it is oxidized, we felt that the application of the flotation process would result in the loss of too much of the oxidized copper, particularly as time went on. In starting up we had a number of difficulties to overcome, as is usually the case. They were not of a serious character. The results obtained, so far, have been decidedly encouraging. During July the plant treated a total of 50,000 tons of tailings, or an average of 1,610 tons per clay. The extraction was about 83 per cent., which, on material of the grade treated, meant a. tailing assaying under 0.1 per cent. copper. The acid consumption was about 65 lb. per ton, and the fuel percentage in. roasting was about 4 ¼ . We expect to cut the fuel percentage in the next month or two to about 3 per cent., the high percentage during July being clue to starting up and shutting clown furnaces. We find that iron builds up in the solution to a marked degree. This is not a serious platter with us, because we have cheap sulphuric acid and can reject solution whenever it gets too foul. The extraction of silver has been remarkably good in the large plant. It has averaged better than 75 per cent. It seems that so long as we maintain salt in the solution to a percentage of 10 per cent., the extraction is good. We allowed the salt to fall off in order to find how low a percentage we could use. At about 7 per cent. the silver extraction fell to 65 per cent. This was to be expected, because silver chloride is soluble in strong brine, but not in weak brine.

Jan 12, 1915

By S. Weinig, E. S. Machlin

THE mechanism of the martensitic transformation has been the subject of a remarkable number of papers in recent years.' Because the task of evaluating all the available theories is a formidable one, the authors will confine themselves merely to the presentation of results of an experimental investigation of the habit orientation, lattice relations, magnitude and direction of average transformation shear strain, and orientation of subband markings for a single variant of the possible habit systems in an 11 pct Mo-Ti alloy. It is hoped that these data will assist in the establishment of a rigorous theory for the mechanism of the martensitic transformation. Experimental Procedure Specimen Preparation: A sample, approximately 1/2 x 1/4 x 1/4 in., was cut from a double arc-melted ingot composed of commercial-process A titanium and high purity molybdenum (nominally, 11 pct Mo). It was capsuled in evacuated Vycor at 10.-3 mm Hg subsequent to a double argon wash and homogenized for 48 hr at 1100°C. After a direct quench into H2O at room temperature, a large-grained specimen resulted, which, by X-ray diffraction and metallographic means, was subsequently shown to be in the body-centered-cubic phase, ß. Two adjacent, approximately perpendicular faces were polished and the ends of the specimen were ground parallel in a V-block. Fiducial lines were inscribed at 45" to the common edge of the polished surfaces in order that they would be perpendicular to the habit traces of the subsequently formed a plates.2 The specimen was then compressed between the ground ends in a Buehler press to stress induce the transformation ß to a'. Experimental Measurements: The habit-plane orientation, the orientation of subband markings, the orientation relation between the parent ß and the product a', and the transformation strain for the same plate were determined. To further establish the nature of the subbands, one surface was repol-ished and etched. The bands in the a' were still visible and were found to be optically anisotropic under polarized light. Back-reflection Laue patterns were obtained from the ß matrix and from a region comprising both a' plates and adjacent ß matrix (a Cu target was used). By superposition of the two X-ray diffraction patterns, it was possible to obtain a tentative orientation relationship between the two phases. This result was confirmed by a pattern taken after the specimen was rotated to make the basal-plane pole of the hexagonal a' phase strike the film. This necessitated cutting the specimen to expose the a' plate in the new position relative to the X-ray beam. Analysis of Experimental Measurements On the basis of the aforementioned X-ray patterns, the orientation between parent and product

Jan 1, 1955

By W. E. Hall

The Darwin mining district of California, 160 miles north of Los Angeles, has yielded an estimated $45 million in lead, silver, zinc, and copper since 1875. The deposits are in silicated limestone of Penn-sylvanian and Permian age close to a stock of horn-blende-biotite-quartz monzonite. Steep left-lateral strike-slip faults that strike N 50o 70o E served as feeder channels for the ore solutions, and minor folds, favorable beds, and faults close to the feeder channels localized the orebodies. The orebodies are irregular, steep, pipe-like replacements that cut across bedding, bedding replacements, and fissure fillings in a calc-silicate country rock consisting mainly of wollastonite, grossularite-andradite garnet, idocrase, and diop-side. Near surface the ore consists of a soft, crumbly mass of cerussite, hemimorphite, jarosite, limonite, and some secondary copper minerals. The primary ore consists of galena, sphalerite, pyrite, pyrrhotite, and chalcopyrite with smaller amounts of andorite, matildite, scheelite, tetrahedrite, and a high silver-lead-bismuth-selenium-bearing galena. Preliminary X-ray fluorescence study of the primary ore has indicated unusually high concentrations of bismuth, selenium, and silver in galena from certain parts of the mine. One phase of this study concerns the distribution of these elements. Twenty-two samples of galena and thirty of sphal- erite have been purified and analyzed or are being analyzed. The selenium in 21 samples of galena ranges from 0.0045 to 2.11 pct; four contain more than 1 pct. These four samples were collected from the Essex orebody, where most of the selenium is concentrated. Three are from high-grade, near-surface ore, and one is from the 700 level. Four of the seven analyzed galena samples from the Thompson workings immediately north of the Essex orebody contain 0.29 to 0.74 pct Se; the others show less than 0.05 pct. Galena from the Defiance workings about 2000 ft south of the Essex has less than 0.03 pct Se. The selenium-bearing galena also contains concentrations of bismuth and silver. Preliminary spectographic analyses indicate as much as 3 pct Bi and more than 1 pct Ag in galena with high selenium content. Antimony is the only other minor element present in galena in amounts greater than 0.02 pct; most galena contains between 0.1 and 0.2 pct and the maximum is 0.72 pct. A second phase of the study is an attempt to determine the temperature range of the formation of the ore. The iron content of sphalerite deposited in equilibrium with pyrrhotite is being used as a geologic thermometer as it is a function of the temperature of deposition of the ore.' Preliminary results indicate temperature range from 430°C in replacement orebodies in the Defiance workings to 180°C in shallow fissure fillings.

Jan 1, 1960

By Walter R. Hibbard

ALTHOUGH extensive pole figure studies have been reported by Brick1 and others2,3 showing preferred orientations in rolled and annealed 70-30 alpha brass, and by Hermann and Sachs4 in 70-30 alpha brass cups, comparatively little information has been published on textures in drawn 70-30 alpha brass tubes. Crampton5 reported a (110), [III] texture in drawn brass tubes based on selected individual X ray photograms following the method of Norton and Hiller,6 but the extent and ease of developing this texture and the annealing texture have not been reported and pole figures of these structures were not found in the literature. The present investigation was further stimulated by the possible influence of textures on the mercury stress-corrosion cracking of brass tubes. According to Snoek,7 iron-nickel alloy sheet with a strong cubic texture was more resistant to corrosion than a randomly oriented sheet. Crampon5 reported that single crystals of 70-30 alpha brass drawn into tubes did not crack after 24 hr s in a standard mercurous nitrate test. Edmunds 12 found that single crystals of 70-30 alpha brass did not crack in a standard mercurous nitrate test but did crack when subjected to ammoniacal vapor. Wassermann reported a similar failure of a single crystal of brass in ammonia. A tube with a strong texture might approach single crystal behavior and tend to resist mercury cracking if the differences in orientation between adjacent grains were sufficiently slight. PREPARATION OF SPECIMENS Crampton's5 tubes drawn 28 pct reduction in area developed a slight degree of preferred orientation while in tubes drawn 48 pct it was more marked. Based on this information, in order to obtain specimens with markedly different degrees of preferred orientation but with the same final draw, two sets of 70-30 alpha brass tubes were made following the mill schedules shown in Table I. The final draw involved a large reduction in outside diameter, and a ratio of reduction in area to reduction in outside diameter similar to that reported by Crampton,5,10 to cause mercury cracking of brass tubes. The brass contained 71.31 pct copper, 0.01 pct lead, 0.01 pct iron and zinc by difference. DETERMINATION OF POLE FIGURES A 6-in. length from each tube was rotated rapidly by an electric stirrer in 25 pct nitric acid until the gauge was reduced to 0.003-0.005 in. The tubes were carefully cut longitudinally and the walls gently flattened. They were further etched in 25 pct nitric acid to 0.0015 in. thickness and (III) pole figures were determined using copper K-alpha radiation in the manner described by Brick.1 The three general types of pole figures

Jan 1, 1947

By H. E. Hostetter

GROSSMANN'S method1 for calculating the hardenability of steel from the composition and grain size has gained wide acceptance, and when properly used, has been well proved in practical application, both for estimating the hardenability of known compositions and for selecting compositions to meet known hardenability requirements. The Grossmann method uses as the criterion of hardenability the "ideal critical diameter," DI, which is the size round that will just harden to a structure of 50 per cent martensite at the center when given the theoretically fastest quench, or "ideal quench." The DI can be related to actual quenching practice if the severity of quench is known; or the DI can be correlated with hardenability as determined by the Jominy end-quench method.2 The effect on hardenability of each element present in a steel can be represented by a multiplying factor, the value of which depends on the amount of the element present and the degree of effectiveness of the particular element. (The value of the factor for a given carbon content varies with the grain size.) When the factors for all elements present in a steel have been determined from charts, they are multiplied, and the resultant product is the D, of the steel. The validity of Grossmann's scheme of multiplying factors has been confirmed in a number of other investigations.3 -6 It can be appreciated that a certain desired level of hardenability can be obtained by adding chromium, molybdenum, nickel and vanadium to a base composition containing carbon, manganese, silicon and the usual impurities. Many of the ranges of hardenability used commercially can be obtained by any of the foregoing alloying elements used singly in the base composition. The same hardenability levels also can be obtained through use of any of a great number of multiple element combinations of these alloying elements. In this report a method will be presented for determining what one combination is the most efficient (based on hardenability and cost of added alloy) for a given level of hardenability. It should not be assumed from the emphasis given to hardenability, that hardenability is the only criterion for the selection of an alloy composition. Such properties as the hot-working characteristics, machinability, carburizing characteristics, amount of distortion during quenching and the retention of austenite may be considerations of importance in many applications. However, hardenability is one of the most important properties of an alloy steel and one that must be considered in almost all cases involving heat-treatment by quenching. Therefore, a method for determining the lowest-cost alloy combination should be of considerable value. DETERMINATION OF MOST EFFICIENT COMBINATIONS For comparing the alloying elements on standards of cost and hardenability, it

Jan 1, 1945

Duval Corporation, the copper -producing subsidiary of Pennzoil United, Incorporated, operates two plants south of Tucson, Arizona: Esperanza and Sierrita. The newest, the Sierrita mine and mill, came into full production in 1971 at a rate of 72,000 tpd (now 87,000 tpd) which at that time made it the largest designed copper plant in the world. In 1968, the largest milling units then available were selected; these included fourteen 16$ ft by 19 ft long ball mills (5 m by 5. 9 m), each equipped with a 3,000 hp motor. The primary crushing plant contains two 60 in. by 89 in. Allis- Chalmers gyratories; each is constructed for two-side dumping by 120- and 150-ton trucks. Access to the maintenance areas to service the drive motors, eccentrics, apron feeders, and speedup conveyor is through hatches between the crushers. This area also contains a good- sized elevator and a stairway. The rather shallow dump pocket, the 300 -ton crushed ore pocket, and the overall design make for a fine, compact arrangement for two 60 in. crushers installed in a total height of 89 ft (27. 13 m) from dump level to top-of-concrete at the lowest level beneath the collecting or speedup belt. This efficient layout is reflected in the short changeout time for mantle and concaves (8 and 30 hours respectively) and the low crushing cost of $0.0416 per ton (first half of 1975). The crushers are installed on 64 ft (19.5 m) centers (Bougainville and Palabora, both with 54 in. crushers, are 55 and 43 ft or 16.76 and 13. 1 m respectively). The switchgear room is located between the truck dumping positions on one side of the installation and the hydroset units on the other side, with dust collecting equipment mounted over the latter. Protection of these units from rocks falling from the trucks is provided by concrete walls. Truck dumping is controlled from a central 25 ft high (7.6 m) control room, which permits fine observation of operations. Removal of oversize boulders from the dump pockets is done with a hydraulic grapple or a rock hook or sling attached to the auxiliary crane hook. Spare mantle stands and a maintenance room with a zinc melting furnace are located at one end of the surface facilities. The maintenance room is equipped with an access door for the crane, which is equipped with 80- and 20-ton hoists. The two 84 in. (2. 135 m) apron feeders receiving the 6 in. (152 mm) crushed rock discharge onto a 72 in. (1. 83 m) collecting and wear belt operated at a speed compatible with the 60 in. (1.524 m) takeaway belt running at a speed of 630 fpm (192 mpm). This belt feeds a 54 in.

Jan 1, 1978

By R. D. Fisher, A. G. Metcalfe, F. A. Crossley

AN investigation of the application of magnetic stirring to the consumable arc melting of aluminum and nickel demonstrated that grain refinement could be obtained when there was sufficient stirring force. Also, it was found that grain refinement could be obtained in aluminum by alternately reversing the polarity of the current through the magnetic stirring coil. The effect of the latter was to periodically reverse the direction of rotation of the molten metal.' Because of their promise, these findings were applied to conventionally cast aluminum. Shortly after this work was begun, Kondic2 reported experiments in which aluminum cast with 40° C (72° F) of superheat exhibited a grain size of 10 to 20 grains per cm2. When interrupted rotation was applied during solidification by mechanical means, a grain size of 800 to 1000 grains per cm2 was obtained. Magnetic Stirring Applied to Conventional Casting— The applicability of magnetic stirring to producing grain refinement in conventionally cast metal was evaluated in the following way. Three pounds of aluminum were cast into a clay-bonded graphite mold. The mold was surrounded by a magnetic stirring coil of 1000 turns. A direct current of 200 amp was passed through the casting by means of a graphite electrode which was immersed from the top. A current of 5 amp was passed through the stirring coil. The aluminum (1100 or 25) was superheated to 1600° F (871° C) and fluxed by passing chlorine gas through the melt for 2 min prior to casting. Two additional castings were made for comparison: 1) a conventional casting of metal superheated to 1600° F and 2) a similar casting through which 200-amp current was passed during solidification, but without current in the magnetic coil. Macrosections of these three ingots are shown in Fig. 1. Very remarkable grain refinement resulted in the magnetically stirred casting. It may be noted that the casting through which 200-amp current was passed has a coarser structure than the conventional casting. It would appear that this is due to the additional heat input generated by the current to the former casting. Mechanical Rotation Applied to Conventional Casting—A casting temperature of 1600° F was used throughout. Some 1 1/2 -lb castings rotated at 150 rpm were predominantly columnar with small areas of coarse, equiaxed grains close to the lower mold wall. Following these, a series of 3-lb castings was made. Macrosections of these ingots are shown in Fig. 2. The imposed rotational speed was 250 rpm. Turbulence or acceleration in the solidifying liquid was the significant factor. Fine grain size resulted in the material solidified during the period of acceleration of the liquid to the imposed rotational speed. When uniform rotational motion was achieved, the grain structure reverted to columnar. The final casting shown in the lower right-

Jan 1, 1962

By E. I. Larsen

Powder metallurgy has been described as being "as old as the pyramids and yet as new as the latest bomber." While this may be true literally, it has been only in the last Io or 1 5 years that widespread use has been made of this method of fabricating various articles One of the first items fabricated by powder metallurgical methods on a commercial scale was electrical contacts, which were fabricated from metal powders as early as 1920. The reason why these contacts have been made for so many years by powder metallurgy is because the compositions of most of these materials arc such that they could not be made by the conventional methods of melting, casting, rolling, drawing, etc. Classes of Contacts In general, electrical contacts fabricated from metal powders fall into two classes, both from the point of view of the method of fabrication and the components that make up the chemical composition. One class is composed of silver or copper and a refractory metal such as tungsten, molybdenum, or compounds thereof such as carbides. The second class contains a predominant amount of silver and a semi-refractory material such as cadmium oxide, nickel, cobalt, or graphite. Electrical contacts of the first class have been developed primarily as facing materials for heavy-duty circuit-interrupting devices. This class of material unites the high melting point, high boiling point, great resistance to electrical erosion, and the refractory characteristics of molybdenum, tungsten, or their carbides with the high thermal and electrical conductivities and low contact resistance properties of silver or copper. The second class of material is used for lighter duty or specialized applications such as heavy-duty, direct-current aircraft relays. Varieties Electrical contacts fabricated from metal powders are made in a wide variety of compositions, shapes, and sizes. In general, the sizes of electrical contacts fabricated from powders vary from contacts weighing approximately 0.001 oz. up to those weighing 3 to 4 Ib., the contacting face being as small as 0.001 sq. in. up to approximately 6 sq. in. in area. The shapes consist of: sound or rectangular buttons, some of which have radii or bevels on the contacting faces; rings, cups, rectangular bars, the contact face of which may have a radius across the width and/or at one end; and form-pressed pieces of a wide variety such as angles, both acute and obtuse, irregular shapes comprising a radius blended into a flat, which in turn may have steps. The size or shape of the contact is dictated principally by the mechanical design of the manufacturer, who in turn may have problems to be met such as: (I) size, weight, or space limitations; (2) electrical conditions such as blowout coils arc chutes, baffies, gas passages.

Jan 1, 1945

By A. G. Guy

IN the course of a research on steady-state diffusion it became necessary to make diffusion calculations for a finite solid. This problem was found to be sufficiently different from the corresponding problem for the semi-infinite solid for an entirely new approach to be necessary in obtaining a useful solution. The assumption of an exponential variation of the diffusion coefficient with concentration' was no longer a practical necessity, so that the new solution offered the possibility of handling diffusion in semi-infinite solids for arbitrary variations of the diffusion coefficient with concentration. In discussing the principle on which this method of calculation is based, it is convenient to consider the relatively simple problem of one dimensional diffusion into a matrix that initially contains none of the diffusing substance. The solution of the set- ond Fick Law,?C/?t=D0?2C/?x2, for diffusion into a semi-infinite solid is3 C/C=1-F(x/2vD0t [1] where C is the concentration of the diffusing substance at a depth x cm below the surface after dif- A. G. GUY, Member AIME, and M. GOLOMB are Professor of Metollurgicol Engineering and Professor of Mathematics, respectively, Purdue University, Lafayette, Ind., A. S. YUE, formerly with Purdue University, is now with the Research Laboratory, Dow Chemical Co., Midland, Mich. TP 4505E. Manuscript, July 18,1956. fusion has occurred for t seconds; C, is the concentration of the diffusing substance at the surface of the matrix; Do is the diffusion constant, expressed in sq cm per sec; and + represents the Gauss error function. The corresponding solution for diffusion into a finite solid of thickness, 1, is conveniently written for the change in concentration, C+, at a point midway through the matrix and for times, t, great enough for the following equation to be used4 c1/2/C1=(1/2-2/?e-?2D0t/l2) [2] In Eqs. 1 and 2 the diffusion coefficient is represented by the symbol Do (rather than by D) to emphasize that these equations are valid only for constant values of D. To understand how these equations can be modified to take account of actual variations of D, such as that shown in Fig. 1, first consider the fact that for a given time of diffusion theconcentration at a given point is determined, not by the D value alone, but by the ratio x/vD0 in Eq. 1 and by a similar ratio, l2/D0 in Eq. 2. Even though in fact D varies, it is possible to keep this ratio constant by using a special coordinate system, x,, that varies with D in the appropriate manner. The necessary character of this special coordinate system can be easily visualized by imagining that

Jan 1, 1958

By W. R. Seyfried

The problem of extracting alumina from clay and low-grade bauxites has been the subject of considerable interest for some time. The basic reason, of course, lies in the fact that known reserves of high-grade bauxite are not inexhaustible and that the present commercial process cannot successfully cope with low-grade bauxites or clay. Ultimately it may be necessary to use these poorer raw materials for all commercial operations. An unprecedented demand for aluminum during the war emergency and the constantly threatened shortage of imported bauxite caused the United States Government to take an active role in this problem. Through the National Academy of Sciences, the War Production Board and the Defense Plant Corporation, it investigated processes and finally selected four to be used in the construction and operation of four experimental plants. Each plant was to be designed in such a way as to facilitate its incorporation into a larger commercial plant should the need arise. The Chemical Construction Corporation process, or "Chemico" ammonium sulphate process as it is better known, was one of the four selected. The extensive literature and patents on record before the war generally follow one of two process-types depending upon the extracting reagent used. One group uses an acid reagent, the other an alkali. The Chemico process uses an acid reagent. For those interested in other processes described in literature, there is a short bibliography at the end of this paper. The basic steps of all acid extraction processes are: (I) to make the aluminum silicates acid soluble, (2) to separate the released alumina from impurities and insoluble gangue, and (3) to recover the reagent for re-use. The Process as Planned In April 1941, a group of chemists and engineers of the Chemical Construction Corp. became interested in the problem of extracting alumina from clays and low-grade bauxites and, under the leadership of Mr. Blakeslee Barnes, Vice President and General Manager of the company, first drafted a process in July of that year—five months prior to Pearl Harbor. The work continued in our research laboratories and with a small pilot plant. By June 1943, the work had progressed to a point where it was possible to apply for a patent on the process which was U.S. patent 2,405,426, granted August 6, 1946. Immediately prior to the development of the Salem plant, the process was envisaged to be as follows: I. The first step was to make the aluminum silicates acid soluble. It was planned: (a) Clay or lour-grade bauxite would be roasted at 700 to 800°C (1292° to 1472°F). (b) The roasted ore would then be leached with hot aqueous ammonium bi-sulphate to form a slurry. In this way the

Jan 1, 1949

By Robert T. Pring

IN addition to the humanitarian aspects of a dust-control program, certain economic benefits are becoming more fully recognized and now furnish a greater incentive to the mill operator to eliminate the dust from his plant. It is well known that a favorable working environment promotes increased efficiency and improves the morale of employees, with a consequent decrease in labor turnover rates. One progressive mining company1 estimates its direct and indirect labor turnover costs at from $40 to $50 per man; for skilled labor, such as machine operators, the cost may approach $100 per man. While these figures may not be typical of every plant, they do indicate a potential savings in operating costs. Dust signifies waste; its elimination decreases maintenance costs of mechanical and electrical equipment; its recovery is often justified from the standpoint of values contained in the collected material, although in large-scale milling operations handling low-grade ore this may not be an important factor. Perhaps the most significant trend in industrial sanitation is that dust-control equipment and procedures are becoming recognized as integral parts of the mill flowsheet. Exhaust ventilating systems are now afforded equal care in design and maintenance to that given other machinery more closely related to the process. METHODS OF DUST CONTROL "Dust control" is a general term descriptive of a number of methods used singly or together to prevent particulate atmospheric contamination. These methods are, in order of their desirability, as follows: I. Change of process. 2. Isolation of process. 3. Wet methods. 4. Ventilation. a. Local exhaust ventilation. b. General ventilation. 5. Personal respiratory protection. In ore-concentration plants it is seldom practical to change a process to eliminate the dust it creates It is possible in some cases to isolate a particular dust-producing operation from the rest of the plant, provided that constant attention on the part of the operators is not required. Frequently water is used in fine crushing and fine grinding, primarily to facilitate the processing of the ore and incidentally to allay the dust produced. Wet methods, however, are not feasible in coarse and secondary crushing operations, because of the relatively great volume of water required in instantly wetting each freshly fractured rock surface. Such quantities of water would entail conveying and storage difficulties. Local exhaust ventilation is the most effective and most commonly applied control measure, wherein directional air currents through the zone of dust production draw the contaminants into a duct system. Unfortunately, only a limited amount of reliable information on exhaust ventilation is available.2-5 Of necessity these publications are rather general, being primarily concerned with engineering fundamentals. While many organizations have successfully

Jan 1, 1940

By S. Feuerstein, J. M. Galligan

SURFACE effects in the deformation of metal single crystals have been noted by a variety of workers.&apos; A large majority of these experiments have used surface roughening or a second chemical constituent at the surface. We performed the following experiment to see how the presence of a fine polycrystalline layer influenced the stress-strain behavior of single crystals of zinc. Spherical single crystals oriented for single slip were acid-machined into cylindrical dumbbell-shape shear specimens having a gage length of 3.2 mm and a cross section of 20 sq mm. The crystals were then deformed varying amounts at the surface of the gage section with emery paper. Following this treatment, the crystals were heated for approximately 8 hr at 200°C. This resulted in a polycrystalline surface layer, with an observed grain size of approximately 1/10 mm in diameter. Deformation of these crystals was undertaken at 78°K in pure shear, the shear direction <1120> lying in the basal plane. The deformation apparatus has been described elsewhere.2 The presence of a polycrystalline layer markedly raises the first observed plastic deformation, as can be seen from Fig. 1. This yield stress is then followed by a region of low work hardening comparable to that observed in shear tests on similar untreated zinc single crystals. However, the extent of easy glide is quite seriously curtailed, and is followed by a more rapid increase in work hardening, reminiscent of Stage II observed in the deformation of fcc crystals. The higher yield can probably be attributed to the difficulty of penetrating the grain boundaries by dis locations from sources interior to the boundaries. If we associate this yield stress with this breakthrough, then we might expect that more dislocations would be trapped in the crystal than is normally the case. If there is an increased density of trapped dislocations, this could lead to an increase in the amount of trapping with increasing strain, giving a larger strain hardening in the crystal. Although other explanations can be given, it is quite tempting to characterize the region of increased work hardening as a manifestation of Stage II-type deformation. This seems reasonable if the polycrystalline layer is quite efficient in holding up a large number of dislocations. In zinc crystals, since it is well-known that the slip distance can be as large as the specimen dimensions, one would only expect to entrap a large number of dislocation with very large crystals or by applying surface layers. If this interpretation of increased strain hardening is correct, it might present a method of studying Stage II in single crystals of zinc. This research was supported, in part, by the Office of Naval Research (1959). The encouragement and stimulation given by Professor E. R. Parker is gratefully acknowledged.

Jan 1, 1965

By P. Campbell, E. M. Gardy, L. Hachkowski

Introduction The efficient process control of a uranium mill and its associated mining operation requires a nearly continuous knowledge of the uranium concentration in the ore. Generally, the approach is to use laboratory assays of grab samples from the mill feed belt. In some cases, elaborate and expensive systems have been used to ensure random sampling, but even with these systems, mass-balance discrepancies still exist. There is a requirement for an on-line instrument that can give a prompt, accurate analysis of a large portion of the feed stock. The authors have recently evaluated a laboratory system that achieves these goals using neutron activation and delayed neutron counting. The development of the on-line uranium ore analyzer is a consequence of previous work done at the Whiteshell Nuclear Research Establishment (WNRE) (Campbell et al., 1978 and 1981), and is based on the emission of delayed neutrons after the neutron irradiation of fissionable materials. The mechanism of delayed neutron emission has been described (Keepin, 1965), but briefly is as follows. The fission fragments resulting from irradiation are in an excited state. Certain of these delayed fission products, precursors, decay to a more stable state by the emission of a delayed neutron. The delayed neutrons can be divided into six groups with effective half-lives ranging from 0.2-55 sec. It is important to note that more than 50% of the delayed neutrons are emitted within the first 6 sec after irradiation; this has an influence on the design of the analysis instrument. Measurement Cycle The measurement cycle involves a neutron irradiation, a delay time, and a counting period. The laboratory demonstration model of the ore analyzer comprises a californium-252 neutron source, a source chamber, a pneumatic transfer facility, and a neutron irradiation and counting module [(Fig. 1)]. The ore sample is placed in the irradiation chamber. The neutron source is transferred to the center of the chamber for a 5-sec irradiation while the helium-3 thermal neutron detectors are switched off. As the source returns to the storage chamber, the detectors are switched on and the delayed neutrons counted. The total assay cycle is 16 sec. The current instrument design calls for the ore to be cut from the main belt every 16 sec by a mechanical arm and dropped into the analyzer directly, or via another belt. The cycle starts when there is a sufficient mass of ore in the irradiation chamber to depress a switch and stops when the delayed neutron count is finished. The ore is then dropped through the bottom of the analyzer chamber and returned to the main belt. The counts from each cycle are accumulated and the data reduced to give the desired information.

Jan 1, 1985

By C. W. Haworth, M. G. Hall

THE diffusion of copper in silver has been studied by Sawatzky and Jaumot1 and Cahoon and Youdelis.2 Sawatzky and Jaumot measured tracer diffusion in single crystal specimens but the 12.9 h half-life of CU64 limited the time of the diffusion anneal to about 50 hr, and in consequence the minimum temperature of the investigation was 716°C. Cahoon and Youdelis employed polycrystalline diffusion couples, and measured the interdiffusion coefficient over the range 0 to 2 at. pct Cu. These experiments were performed at temperatures down to 630°C and a discontinuity in the Arrhenius plot was observed at -700°C. The high D-values at lower temperatures were attributed to the effect of grain boundary diffusion. The present experiments were undertaken to investigate this anomaly and study the low temperature region of the diffusion of copper in silver. The method employed single crystal samples and a plated source of copper. Silver single crystals 2½ by 4 in. sq were grown from zone-refined silver bar (>99.99 pct Ag) by the Bridgeman technique, and $ in. thick slices were cut by spark machining. The specimens were electro-polished, and a 2 µm layer of copper was electroplated on to the planar surface at one end of each specimen using a potentiostatic voltage source. The following solution was used for plating: CuSO4 . 5H2O 125 gm H2SO4 (S.G. 1.8) 53 ml H2O 1000 ml The diffusion couples were sealed in capsules, in an argon atmosphere, and were annealed at either 701, 597, or 498°C. The temperature measurements were made to an accuracy of *l°C After the anneal, the penetration of copper into the silver was determined using an electron probe micro-analyzer. 100 sec scaler measurements of the intensity of Cu Ka! X-radiation were performed at 2 to 5 µm intervals along a line normal to the source, with a beam potential of 25 kv. Allowances were made for the background radiation, and the dead-time of the counting system in the normal way, and the intensity ratios were corrected to weight concentration by the method of Duncumb & de-Casa.3 For the analysis of the diffusion profiles, it was first observed that the copper layer did not completely dissolve in the host crystal for the specimens annealed at 498 and 591 whereas for the specimen annealed at 701°C, a single phase solid solution was formed. For the 498" and 597°C specimens, the concentration of copper in the silver at the interface was therefore constant throughout the anneal and the analysis for a

Jan 1, 1970