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By L. D. Jaffe, D. C. Buffum, F. L. Carr, M. Goldman

IN an earlier investigation by Jaffe and Buffum,' a time-temperature transformation diagram for isothermal temper embrittlement of an SAE 3140 steel showed some suggestion of two overlapping "noses." These authors mentioned, as a possible explanation, that their specimens embrittled at 525°C were slightly softer than expected, perhaps because of some slight unknown variation in tempering treatment. Because a double nose would be of considerable importance in elucidating the mechanism of the temper brittleness reaction,' it appeared worthwhile to redetermine the portion of the diagram near 525°C. The same heat of SAE 3140 steel was used as in the previous work, and experimental procedures were the same.' Blanks were austenitized 1 hr at 900°C, water-quenched, tempered 1 hr at 675°C, and water-quenched. Groups were then embrittled isothermally at the temperatures and for the times shown in Table I, and water-quenched. For comparison, one series was tempered various times at 675°C, water-quenched, and given no further treatment. V-notched Charpy impact specimens machined from the blanks were tested over a range of temperatures. Energy absorption and percentage of fibrous fracture were plotted vs test temperature for each group, utilizing both the new data and those for the same treatment obtained in the earlier work,' except that the earlier data at 525°C were discarded.* are given in Table I. After a small correction (-0.28°C/BHN) for the variation of hardness" from Rockwell C24, the transition temperatures were plotted against embrittling time for each embrittling temperature, using the data of Table I for the range of treatments covered there, and using previous dataL for other treatments. By cross-plotting, Fig. 1 was finally obtained. It appears from this figure that if there are two noses to the time-temperature-transformation diagram for isothermal temper embrittlement of SAE 3140 steel, the upper nose must lie above 575°C. As interest had been primarily in the region near 525°C, a thorough study of embrittlement above 575°C was not made. Fig. 1 shows that such a study is needed, especially since preliminary measurements at the tempering temperature of 675°C reveal much more rapid embrittlement than would be expected from the data at lower temperatures. The transition temperature increased, approximately, as the logarithm of the tempering time (about 81/2°C, after correction for hardness, for each doubling of the time).& Since three tensile specimens tempered 240 hr at 675°C and water-quenched had an average yield-tensile ratio of 0.844, with 70.7 pct reduction of area, it is assumed1 that the tempering temperature is not above Ae1. The omission of a quench-and-reheat operation below Ae1, for the groups tempered at 675°C, as compared to groups embrittled at lower temperatures, should not affect results." More detailed investigation of embrittlement above 575°C is now under way.

Jan 1, 1954

By D. S. Sanders, E. W. Witcomb

FLOTATION tests, from which the flowsheet for the concentrator was determined, were made in a pilot mill at Chuquicamata and later comparative testing of full size flotation cells was done at the Andes Copper Mining Co. concentrator in Potrerillos. The principal sulphide copper mineral is chalcocite with varying amounts of covellite, enargite, and chalcopyrite. The oxidized portion is principally antlerite, a green basic copper sulphate. The gangue minerals include quartz, sericite, pyrite, feldspars and kaolin. Chuquicamata pit ores have proved to be hard grinding. The results of the flotation and metallurgical work prove that good extraction and concentrate grade can be obtained under a rather wide range of conditions. Metallurgy In general the metallurgy of the ore requires a fairly close control of the grind to the primary rougher flotation and of the regrind of the primary rougher concentrate before cleaning. Regrinding is necessary to reduce the insoluble content of the final concentrate since a large part of the middling combinations are between the copper and silica minerals. The alkalinity of the pulps can vary between wide limits and control of the percent solids gives the desired grind. Alkalinity Control Arroyo Salado water is supplied to the concentrator head tanks. It contains about 5000 ppm of salts but test work shows no deleterious effect. Actually this water permits a decrease of as much as 1 lb of lime per ton of ore as compared to San Pedro water, which is used for leaching oxide ore. About one quarter of the water flow through the concentrator will be new water from the reservoir. The remainder will be reclaimed from the thickeners. For primary grinding and rougher flotation, new and reclaimed water are combined in the head tanks. To avoid erratic results, only new water is used in the regrind system. Milk of lime is added to the rod mill feed chute to obtain the maximum benefit of mixing with the ore. This protects the equipment from corrosion due to entrained acids and soluble copper salts. Lime consumption is 3 to 5 lb per ton of ore. Grinding and Flotation The six bin gates over each belt feeder are opened in controlled combination and sequence to keep a live bed of ore in the bin to prevent arching and to obtain mixing. The milk of lime added ahead of the rod mill is regulated to suit rougher flotation requirements. The primary promoter reagent, Aerofloat 238, is added to the 10-mesh discharge of the rod mill as it is laundered to the ball mill scoop box. The classifier overflow is regulated to produce a grind of about 12 to 15 pct +65 mesh and 60 pct -200 mesh. A secondary promoter Z-4 (aml zanthate) and the pine oil frother are added in the overflow launder where a pH of 10 to 11 is maintained. Comparatively heavy beds of froth on the

Jan 1, 1952

By C. E. Birchenall, M. H. Davis

The rate of oxidation of titanium in the temperature range 650° to 950°C has been measured. 'The linear rate law obtained is explained by interface reaction control of the process. Tracer experiments indicate the oxygen diffuses in at least one phase of the scale. THE kinetics of oxidation of titanium, principally at low temperatures and under varying oxygen pressures, have been reported.¹,² In the present investigation the rates of oxidation of titanium have been determined by the gravimetric method in the temperature range 650" to 1000°C under 1 atm of oxygen. Some of the scales formed were powdered and an endeavor was made to identify the phases present by X-ray diffraction. Alexander and Pigeon1 ascribe a logarithmic rate law to the growth of thick films formed at temperatures up to 565"C, and Gulbransen and Andrew' suggest that the parabolic rate law holds up to 600°C. In both cases large initial deviations from those relationships are apparent and no explanation is offered. In the present experiments it is shown that a linear rate law holds in the temperature range 650" to 1000°C, and it is suggested that the initial deviations observed are due in part to a spurious surface area effect and also to the high solubility of oxygen in titanium metal. Positive information on the mechanism of oxidation is not available, but oxygen ion diffusion through the lattice is inferred from the re- sults of trace experiments. X-ray analysis of the scale confirms Gulbransen and Hickman's' observation on the presence of rutile TiO². Experimental Details The titanium used in these experiments was prepared by the reduction of titanium tetrachloride with magnesium and was used in the form of rolled sheet. Samples approximately 0.040x0.5x0.75 in. were cut and polished to a fineness of 4/0 on metal-lographic polishing papers. Individual samples, suspended on a platinum wire, were lowered into the hot zone of a furnace, in vacuo, by means of a winch, and oxygen was admitted after the desired temperature was attained. The increase of weight of single samples oxidized for fixed times at certain temperatures was determined by weighing the sample before and after oxidation on a chemical balance. In these instances the sample was used for only one time at one temperature. In some cases the course of oxidation was followed by attaching the sample to a sensitive spring balance and noting periodically the extension of the spring with time of oxidation. Kirkendall type experiments were carried out by marking the surface of the titanium prior to oxidation with a solution containing radioactive silver. After drying, the specimen was heated in vacuo to decompose the silver nitrate used, leaving a small amount of metallic radioactive silver on the surface. Following oxidation, these samples were mounted, ground, and then placed in intimate contact with

Jan 1, 1952

By E. R. Williams

Continuous casting of nonferrous metals has become a commercial reality. After years of slow and arduous experimentation by a number of independently working inventors, starting with Lord Henry Bessemer in 1865, success has finally been achieved. Nonferrous metals—aluminum, copper and brass—extremely vital to our production of war equipment, are being continuously cast every day in large quantities. In fact, some nonferrous metals —for instance, high-strength aluminum alloys—could not have been produced in the huge quantities required for planes and other war equipment if it were not for continuous casting. These casting processes for nonferrous metals have not been entirely perfected but continuous casting is here to stay. Continuous casting of carbon and alloy steel has not yet gone into commercial production but it is well on its way. A great deal of hard work has been expended and -small fortunes in money have been and are being spent in marching down the long road to success. Considerable tonnage of plain carbon, stainless and other alloy steels of high quality have been produced by pilot-plant continuous-casting units and at speeds that give indisputable evidence of commercial success. What has made the difference between failure and success? What is there left to do to revolutionize completely the method of casting metals, which has existed for centuries? These questions are being rapidly answered. Fundamental Methods Present continuous casting processes have many common problems but differ considerably in their methods of solution. The two fundamental methods are: I. Forming the solidifying metal against mold walls that move with the metal, so that there is no relative movement between the walls and the metal. This is a development of Lord Bessemer's idea of pouring metal between two rotating rolls. 2. Moving the solidifying metal along a stationary mold wall so that there is relative movement between the wall and the metal. This was proposed by Trotz in Germany as far back as 1889. Neither Bessemer nor Trotz achieved commercial success. The chief reason seems quite apparent today, inasmuch as they did not solve the fundamental problem of rapid heat removal from the molten metal. The Williams Continuous Casting Process is based upon this fundamental idea of fast heat removal and dissipation. The bessemer method of forming the ingot against a moving mold wall—either rotating rolls or endless moving plates—was carefully considered by the author and promptly abandoned as being commercially impractical. Primarily, only a very small tonnage output was possible except at terrific lineal speeds, because the very nature of this method required casting thin sheets rather than ingots. Secondly, the casting of a sheet between lolls that are pressed together meant the rolling of two layers of solidified skin separated by a molten core. Internal bursts and segregation of the metal usually resulted. Moving Solidifying Ingot Along Mold Wall It was decided therefore to devote our research activities chiefly to- the idea of moving the solidifying ingot along. a stationary mold wall. It was soon dis-

Jan 1, 1945

By Fred W. Stiefel

THIS paper covers the operation and maintenance of Diesel locomotives underground on a portion of the Delaware River Aqueduct.[t] This part of the tunnel is r5 miles long, with shafts 14 ft. in diameter spaced about 5 miles apart. These shafts (2, 2a and 3) are 825 ft., 1551 and 840 ft. deep, respectively. The grade of the tunnel is almost level, except for the 12,100 ft. north of shaft 2, which has a grade of 2.31 per cent. The tunnel is circular, 19 to 19 ft. in diameter, It lies at a depth of 400 ft. below sea level, with a maximum overburden of 2400 feet. The tunnel driving was carried on in six headings, using 15 storage-battery locomotive each of 13 gross tons and having a [inum] speed of 8 miles per hour. The longest haul was 3 miles, which proved to be the maximum distance for which storage-battery locomotives could be employed economically for the amount of material to be moved. One locomotive was used for switching cars at each heading, one for switching cars at the bottom of each shaft and one on the main-line haul from the heading to the shaft. The maximum amount of material hoisted in one shaft from two headings when 104 ft. of tunnel was driven in 24 hr. was 1000 cu. yd., or about 2000 tons of rock in place. The total quantity excavated was 950,000 cu. yd. (1,500,000 tons of rock). The tunnel-lining operations using con¬crete presented a problem of hauling a maximum distance of 5 miles in the tunnel. The concrete-placing equipment had a capacity of about 2000 cu. yd., or 4000 tons in 24 hours. Experience had shown that storage-battery equipment could not handle this quantity of material within the time required at the distance of 5 miles, Trolley-locomotive haulage was considered dangerous and expensive, and would cause interference because transmission equipment would be required. Therefore the performance of Diesel locomotives of 160 hp., with a speed of 18 miles per hour, in comparison with that of other types of locomotives, was studied. CHOICE OF LOCOMOTIVES In deciding upon the type of locomotive to be used for the 5-mile haul, four alternatives were considered, as follows: I. To purchase additional storage-battery locomotives with the necessary batteries over the 6 locomotives and 12 batteries available. 2. To convert the existing battery locomotives to the trolley type. 3. To purchase new trolley locomotives. 4. To purchase Diesel locomotives. PROPOSITION No. I The go-hp. storage-battery locomotives available (each weighing about 12 tons), when hauling a load of 64 tons, with a drawbar pull of about 1900 lb. on a level track, developed a speed of 8 miles per hour and could make five round trips in 8 hr. with a

Jan 1, 1942

By P. W. Barton, A. A. Johnson, E. J. Hughes

THE only systematic work on the effect of carbon on the recovery and recrystallization kinetics of iron so far reported in the literature is that of Venturello et al.1 These workers found that doping high-purity iron so that it contained up to 86 ppm C caused a marked retardation of recovery but had little or no effect on the recrystallization kinetics. Such evidence for an effect as was obtained was inconclusive. In the course of a study of electron-beam zone-refined iron the authors have recently examined the effect of added carbon on the recrystallization kinetics of this material. The amount of carbon added was about 50 pct more than was used by Venturello et 01. and definite evidence was obtained that carbon at this level retards recrystallization. The starting material was taken from a batch of 0.2-in.-diameter Johnson-Matthey spectroscopically pure iron rods which had been further purified by being given three passes at 4 in. hr-1 in a standard MRC electron-beam zone-refining unit. The purity of this material. which was used in two other investigations of the properties of high-purity iron,2,3 was established by an extensive series of analyses. Four laboratories, the National Research Corp., Newton, Mass., the Naval Research Laboratory, Washington, D.C.. the Battelle Memorial Institute, Columbus, Ohio, and the National Physical Laboratory. Teddington. Middlesex, England. carried out analyses for carbon. oxygen. and nitrogen. and in addition a complete mass-spectrograph analysis was carried out at Battelle. In all, twelve analyses were carried out for carbon, seventeen for oxygen and twelve for nitrogen. The amounts of these elements found to be present were 10 * 7, 8 i 5, and 0.9 i 0.1 ppm, respectively. The limits given are rms deviations. Two results, one of 46 ppm of oxygen and another of 31 ppm of nitrogen, were rejected as anomalous. The mass-spectrograph analysis revealed the major impurity to be 25 ppm of tungsten. All of the other elements detected were present in amounts less than 0.5 ppm. Some of the rods were doped with carbon by being immersed in a suspension of carbon in oil, heated in a vacuum at 400°C, and then zone-leveled to make the distribution of carbon uniform. Carbon analyses carried out on three of the doped rods gave a mean value of 130 ± 45 ppm. Some of the doped and undoped rods were deformed 58 pct by swaging and cut into specimens 1/2 in. in length. Isothermal recrystallization experiments were then carried out on them at 575", 600°, 615", 630°, and 650°C, the volume fraction of each specimen recrystallized being measured by a point-counting method described by Hilliard and Cahn.4 The results for the two types of specimens are shown separately in Figs. 1 and 2. In Fig. 3 the pairs of recrystallization curves for each temper-

Jan 1, 1965

By C. H. Bushell, M. Malnarich

REAGENT control in flotation is more an art than a science. Operators vary the amount of reagents used according to the metallurgy obtained. The amount of collector may be increased, for example, if tailings losses are high and decreased if concentrate grades are low. Often control decisions are based on only one assay, such as the tailings analysis by microscopic or quick assaying techniques. This method of control has given reasonably satisfactory results but is clearly far from perfect. No adjustment in reagent addition can be made until information is received that metallurgy has deteriorated, that is, until the optimum time for adjustment has already passed. The first adjustment probably will be an overcorrection or an undercor-rection, and time will be required to establish the best new addition rates. The seriousness of the problem depends on variability of the ore. In ore of continually shifting composition and complexity, mill results will shift back and forth from excessively high concentrate grades and high tailings losses to excessively low concentrate grades and low tailings losses. Economic incentive for improved control also increases with increasing size of mill. Even a small improvement in percentage recovery in a large mill pays for expensive control equipment. The Sullivan concentrator of The Consolidated Mining & Smelting Co. of Canada Ltd. at Kimberley, B. C., treats 11,000 tpd of complex lead-zinc ore, a scale of operations which justifies an extensive research program on automatic control. Work has followed two separate lines: 1) development of faster analytical methods for pulps, and 2) study of metallurgical control by measurement and control of reagent concentrations in flotation solutions. Although work is incomplete in both phases, a progress report can be given on the development of reagent control. The primary requirement in flotation is selective attachment of particles of one mineral to air bubbles while the other minerals remain detached. Hydrophobicity of mineral surfaces is established by the collector, which coats the minerals by adsorption. The requirement of selective affinity for air bubbles is therefore equivalent to requiring selective adsorption of a collector. Adsorption theory, supported by abundant experimental evidence, shows that the amount of a reagent adsorbed at a surface must be a function of its residual concentration in solution. As the concentration in solution rises, the amount adsorbed increases, usually following some type of logarithmic curve. Selective adsorption must depend on a natural difference in the minerals, but ordinarily the differences in affinity are not great enough to obtain complete coverage of one mineral while the others are completely uncoated. An optimum concentration will exist, below which coverage will not be great enough to float enough of the valuable mineral, and above which too much unwanted mineral will float. All other reagents, except some frothers, affect the amount of collector adsorbing at constant residual collector concentration in solution, either by adsorbing competitively with the collector, or by a

Jan 1, 1957

By H. R. Hendricks

THE ore of the Iron King Mine, being very hard and having a very fine crystalline structure, presents many problems in milling that are not present in ordinary lead-zinc ores. This very fine crystallization causes much interlocking of minerals and is most prevalent between the gold and pyrite and the sphalerite and pyrite. The recovery of gold is extremely important in the economic value of the ore and therefore constitutes a large part of the metallurgical problem to be solved. All of the gold is free but it has not been observed in any of the mill products except under very high magnification. Some of the gold is tarnished, and this condition also makes recovery difficult. The gold values have been determined to be highest in association with chalchopyrite and arsenopyrite. The silver occurs mostly in association with galena and tennantite. A small amount of native silver has also been observed occasionally. The galena, being softer than the other minerals, frees much better in grinding. Some oxidation of galena is present in the ore, as the nonsulphide lead content amounts to approximately 10 pct of the total lead. Some oxidation of sphalerite is also present in the ore. Microscopic examination of the zinc concentrate shows many grains in the minus 325-mesh to be half pyrite, together with the fact that some of the zinc is marmatite, makes production of a high-grade zinc concentrate very difficult unless recovery is sacrificed. In the grinding, it has been determined that the flotation feed must be maintained at 85 pct minus 200-mesh to free the minerals sufficiently to make the best economic recoveries. Table I illustrates the effect of fine grinding. It shows that the bulk of the recovered minerals is in the minus 325-mesh size and is the part of the minerals that is more readily freed. The pyrite concentrate is not given in the table because it contains practically no minus 325-mesh. DEVELOPMENT OF PRESENT FLOWSHEET AND MILLING PRACTICE The first milling operation started in 1938, with a mill capacity of 140 tons daily, producing a bulk gold-silver-lead concentrate. After several stages of expansion and changes of the mill flowsheet, the mill is currently treating 380 tons of ore daily and three separate concentrates are produced. In order of their importance and value, these are: (1) lead concentrate, (2) zinc concentrate and (3) pyrite concentrate. The present average grade of mill heads is: gold 0.11 oz, silver 4.00 oz, lead 2.50 pct, zinc 7.60 pct, copper 0.22 pct, iron 22 pet and 30 pct insolubles. Specific gravity of the ore is 3.8. During the period when bulk flotation was practiced, the flowsheet was very simple. The ore was ground fine enough to liberate as much gold, silver and lead as possible. Necessary collectors and frothers were added in sufficient quantity

Jan 1, 1947

The Pinto Valley, Arizona, plant of Cities Service Company is one of the newest copper producers in the United States, having started production in 1975. It is located a few miles west of Miami, Arizona, on the site of the Castle Dome Mine, a World War II operation which was shut down and its mill later moved and renamed the Copper Cities plant. The Pinto Valley concentrator uses six 18 ft (5. 5 m) diameter by 21 ft (6.4 m) long, 4000 hp, single-stage ball mills, taking crushed ore from four 7 ft (2. 13 m) standard and eight 7 ft shorthead crushers. It has a design capacity of 40,000 stpd but has higher tonnage capability for an expanded concentrator. The copper porphyry-ore is mined by two 15 yd and one 20 yd shovels (11.5 and 15.3 cu m) and is transported to the primary crusher by 150-ton trucks. The dump pocket is relatively shallow (14 ft or 4. 27 m deep); and because of the width necessary for the 150-ton trucks, it was designed to use the ore itself instead of manganese steel as an abrasion- resistant lining, although a heavy steel cap protects the dump pocket top edge. The pocket has a capacity of 300 tons of ore. Water sprays under the dumping positions direct a fog toward the truck being dumped for excellent dust control. Breaking of oversize is done by a hydraulic impact breaker mounted on a hydraulically operated articulated arm situated at the side of the dump pocket. Again because of the shallow pocket, no splitter is used to protect the spider of the 60 in. by 89 in. Traylor gyratory crusher, which is equipped with a hydroset mechanism. Crushed ore at 8 in. (203 mm) drops into the 360 st surge chamber lined with AR plate and then is fed directly to the 60 in. (1.52 m) stock- pile conveyor by an 84 in. by 20 ft (2. 13 m by 6.1 m) four-speed apron feeder at a rate of up to 4000 stph. Early in 1977, Cities Service will replace the four -speed electric drive with a hydraulic pump and motor with infinitely variable speed. They are also studying the replacement of the apron feeder with an 8 ft wide (2.44 m) belt feeder similar to those at Similkameen and Palabora. Peak crushing capacity would then be increased to 5000 stph. The AR plate in the crushed ore pocket was installed in large, rectangular sheets which required much less embedded metal to hold it in place than would the more conventional, smaller arid square pieces of cast manganese steel. For duty in the crushed ore pocket, where wear is not severe, the AR plate is a satisfactory material. The new Palabora Figure-8 crusher also uses this design.

Jan 1, 1978

By J. C. M. Li

A dislocation model is constructed for a cylindrical tilt boundary of which the axis of tilt is parallel to the axis of the cylinder. The strain energy per unit area calculated from this model agrees with thut of an infinite vertical wall of edge dislocations. Based on this model, a dislocation mechanism for the shrinking of such a cylindrical boundary is proposed. The power relation between the grain boundary velocity and the driving force is accounted fm by the stress dependence of dislocation velocity. The activation area for grain boundary motion is found to correlate with that .for Creep in aluminum. In the preceding paper' the velocity of a curved grain boundary is found to be a power function of the driving force, [1] where M is the velocity at unit driving force, y is the grain boundary energy, and p is the radius of curvature. This finding adds another nonlinear rate equation to metallurgical phenomena. While a macroscopic theory for the steady state as just presented2 can be used to understand the shape of the moving boundary, such theory usually does not provide any information about the atomistics of the process. It is attempted in this paper to discuss the nonlinear rate equation in terms of a dislocation mechanism. A DISLOCATION MODEL FOR A CYLINDRICAL TILT BOUNDARY Before a dislocation mechanism can be proposed for the motion of a cylindrical tilt boundary, it is necessary to construct a dislocation model for the boundary. For simplicity, a tilt boundary in a simple cubic lattice will be examined. Although this does not represent the situation in aluminum, it is expected that the general features will be similar. The details of the model in the fcc lattice can be constructed when a refinement of the simple model is called for. A cylindrical low angle tilt boundary in a simple cubic lattice is shown in Fig. 1. Two sets of edge dislocations are needed as shown. A schematic representation is shown in Fig. 2(a). In terms of a continuous distribution of dislocations, the density of dislocations n, of Burgers vector b, in the small segment ds of the grain boundary is Similarly the density of dislocations n, of Burgers vector bz in the same segment ds is where is the tilt angle of the grain boundary. Based on this continuous distribution of dislocations in the grain boundary, it can be shown that the interaction between a small segment of the boundary and the rest of it is zero. Furthermore, the stress produced at any point inside or outside the cylinder is also zero. This shows that the model is acceptable in the sense that the boundary has no long-range stresses. Since the model of continuous distribution of dislocations in the grain boundary does not reveal the interactions among individual dislocations, a discrete dislocation model has to be used for such purpose. For simplicity, an equivalent model is shown in Fig. 3. This arises because the total Burgers vector is, from

Jan 1, 1970

By H. C. Chao, Y. E. Smith, L. H. Van Vlack

ThE phase relationships for the MnO-MnS system have been investigated only in the eutectic region. wentrupl reported a eutectic at 1280°C (2345°F) with approximately 50 wt pct of each component, as based on the earlier work of Andrew, Maddocks, and Fowler.2 More recently silverman3 corrected these figures to 2275° ±10°F (1246°C) with 63 wt pct MnS and 37 pct MnO. Neither Wentrup nor Silverman investigated the solid-solution limits in this system. However, Wentrup did suggest with his phase diagram sketch that the maximum solubility of MnS in MnO is 10 wt pct, and the solubility of MnO in MnS is nearer 20 wt pct, Fig. 1. These figures have been utilized in subsequent publications without further verification.4 There had been evidence found in this laboratory that the above solid-solution figures were high. An investigation was therefore conducted using the following procedure. Samples containing MnO and MnS were contained in platinum foil, heated just above the eutectic temperature in a purified nitrogen at- mosphere, and quenched into distilled water. The MnO had been prepared by calcining reagent grade MnCO3 in nitrogen at 1400°C. Only MnO was detected in subsequent X-ray diffraction analyses. The MnS had been prepared by sulfur deoxidation of reagent grade MnSO4 according to the reaction: MnSO4 + S2 ? MnS + 2SO2 The details of the sulfide preparation procedure, which utilized atmospheric sulfur pressures at 950°C, are discussed elsewhere.5 The sulfur content of the sulfide was determined to be 36.8 * 0.2 pct, which matches the stoichiometric value of 36.9 pct. No evidence of either MnO or MnSO4 was detected microscopically or by X-ray diffraction. The single-phase combinations of MnO and MnS at the eutectic temperature indicate the limit of solid solution as shown in Fig. 2(a) and (c), and the revised phase diagram is shown in Fig. 2(d). The melting temperature for MnO is that provided by Glasser.6 The eutectic temperature of 1232° ± 5°C and eutectic composition of 64 wt pct MnS are in close agreement with Silverman's correction of Wentrup's work. The maximum solubility of MnO in MnS was determined to be 1.7 ± 0.2 wt pct. The maximum solubility of MnS in MnO was determined to be 1.8 ± 0.2 wt pct. These figures represent the lower limit for the presence of liquid at 1260°C and are significantly lower than the earlier estimated solid-solution figures. The authors wish to acknowledge the support of U.S. Steel's Edgar C. Bain Laboratory for Fundamental Research. The comments and critical suggestions of Drs. A. S. Keh, R. Rickett, and R. B. Snow are appreciated.

Jan 1, 1963

By P. H. Cardwell

The effect of various plastics upon the permeability of a sand and the mechanism of their ensuring permeability within the consolidated sand body is discussed. Data are presented on various properties of the plastic materials used in sand consolidation and the effect of temperature and usable temperature ranges are considered. The recovery of petroleum from uncon-solidated sands presents serious production problems, owing to sloughing of the formation. The solution of some of these problems usually has been by mechanical methods.' The ideal way of controlling the sloughing tendencies of these un consolidated formations, however, is to prevent the movement of the individual sand grains by the use of a bonding medium. A step in the desired direction was taken by Irons and Stcesser2 when they taught the use of liquid plastics to produce a consolidating deposit that prevents sloughing. More recently, Wrightsman and Buckley3 proposed a modification of this process. In this use of the liquid plastic, the sand grains become coated and the ensuing hardening of the plastic by polymerization or condensation cements together the sand Particles. Liquid plastics that have been suggested for this Purpose are of the following types: phenolformaldehyde, alkyd resins, vinyl type resins, polystyrene and the acrylic esters. When the flow of fluids through a sand must not be materially reduced, it is essential only to cement together the sand grains with liquid plastic and leave a portion of the interstitial space for fluid flow. TO be assured of obtaining fluid flow through the sand interstices after injection of plastic, it has been necessary in the past to remove the excess plastic from the interstices between the particles before it sets. Wrightsman and Buckley suggest two methods for the removal of the excess plastic from the interstices of the sand: (I) to flush the excess plastic back into the formation, away from the well bore, by use of oil or water; (2) to produce the well before the plastic has set, thereby moving the excess plastic into the well bore. Under certain well conditions it is very difficult to remove the excess plastic from all parts of the formation by either of these methods and leave the formation consolidated with no sections completely sealed. The method* described herein of retaining permeability within a sand body is to use a liquid plastic that upon setting ensures permeability within the consolidated sand body. This cements the sand grains but obviates the need for removing plastic from the interstices of the sand. One of the ways of accomplishing this is to use a liquid plastic that occupies a considerably greater volume than the solidified resin it yields on further polymerization or condensation in situ. In order to be satisfactory, the volume of the plastic while liquid should be at least one and one half times the volume of the solid formed. Preferably,

Jan 1, 1947

By C. T. Cox

The title of this short paper necessarily implies that friction articles comprise a distinct and unique field and are different from the other metal-powder articles, such as bearings, electrical contacts, and magnets. However, this distinction is rather hard to define in that it is really one of degree and not Of kind. It would appear that the main distinction is that friction are are articles of rather large pressing area. In other words, they are articles in which the area Of the mold in which the metal powder is pressed is relatively large. Other articles made from meta1 powders are generally of rather small area, usually a few square inches, while in friction articles the area ranges up to several hundred square inches. The satisfactory production of meta1-powder friction articles involves many production problems and a large number Of these are accentuated by the large area' Since it is impossible in the short space of this Paper to describe adequately all the problems involved, I have thought it best to merely mention some Of them and to place particular emphasis upon a problem that is common to all metal-powder articles, although it is of greater importance in the making of friction articles. This problem has to do with the effect On the properties of a friction article of small variations in the composition. However, before discussing the problems, it is probably appropriate that some mention be made of the limitations on the use of friction articles. Most engineers are familiar with the general field that can be characterized as constituted by clutch facings and brake linings, and undoubtedly have heard many stories of how metal-powder friction articles have had a service life 10 times as long as nonmetallic friction materials, and of instances where metal-powder friction articles have operated satisfactorily where no other material could be used. These instances should not be regarded as criteria for the assumption that metal-powder friction articles are capable of universal use in the friction field and will always outperform other materials. On the contrary, metal-powder friction articles are unsuitable for numerous applications, either on a performance or on a cost basis. Metal-powder friction articles will not replace wooden block brakes for wagon wheels nor cast-iron brake shoes for railway car wheels, to mention just two obvious examples. Metallic friction articles cannot be used where they must act as heat insulatorsl for they have a relatively high thermal conductivity as compared with asbestos. This is a detriment in some uses, as in certain airplane brakes having hydraulic actuators provided with rubber seals. The metallic friction material readily transfers the heat developed at the braking surface back through the material to the rubber seal and destroys it. Within its limited field, metal-powder friction material has proved satisfactory, but many problems are involved in the Selection of the proper material to use in an application and in the commercial pro-

Jan 1, 1945

The San Manuel district of Arizona was first prospected prior to the Civil War, but there was little or no production until 1881. Small operators tried to develop ore bodies from time to time after that without success. Then, in 1917, a drilling program was carried out; but it did not indicate a copper content of interest. Finally, in the 1940s, the U. S. Geological Survey drilled the property and obtained indications of a mineralized area. At that time, Magma Copper and its principal owner, Newrnont Mining Corporation, drilled many deep holes through the gravels and conglomerates and in 1948 proved an ore reserve of 367 million tons of 0.785% copper sulfide ore. Underground exploration and development were started in early 1948, and in 1953 design and construction of the surface facilities commenced. By the end of 1955 plant construction was completed. The ore deposit of disseminated pyrite, chalcopyrite, chalcocite, molybdenite, silver, and gold is in quartz monzonite porphyry gangue. Mine production, which is entirely underground, is by full-gravity block-caving. Initially, ore mined at a rate of 36,000 tpd was hoisted through two identical shafts. From the 5,000 ton coarse ore bins located at the collar of each shaft, the ore was hauled by a shuttle service rail- road in 100-ton bottom-dump rail cars a distance of seven miles to the rail bin. Fragmentation from block-caving supplemented by underground grizzlies at the loading stations controlled the oversize material to give a feed size acceptable to the two 7 ft standard Syrnons crushers. During operation, however, it became evident that a large amount of secondary breaking in the grizzlies was necessary to produce a suitable feed for the standard crushers. The decision was then made to install primary gyratory crushers on the surface adjacent to the shaft site ore bins. Under each 5,000-ton coarse ore bin, an 8ft by 20 ft (2.44 m by 6. 1 m) apron feeder transfers the ore to a 13 ft 6 in. by 10 ft (4. 1 m by 3. 05 m) grizzly set at 6 in. (152 mm) and sloped at 45 degrees. About 65% of the ore passes the grizzly and thus bypasses the 42 in. by 65 in. (1.07 m by 1. 65 m) Allis -Chalmers gyratory equipped with a Hydroset mechanism and having an open side setting of 5 ½ in. (140 mm). Grizzly undersize and crushed ore drop onto a 6 ft by 30 ft (1. 83 m by 9. 14 m) variable- speed apron feeder for transfer to a 54 in. by 90 ft 5 in. (1.37 m by 27.6 m) inclined conveyor equipped with a tramp iron detector. Indication of tramp iron automatically stops the conveyor and both apron feeders. When the iron is removed from the belt by the operator,

Jan 1, 1978

By W. C. Ellis, E. S. Greiner

WHEN metallic materials are deformed plastically, the process may be considered as one in which hardening and recovery occur simultaneously. The net hardening is that produced by deformation in the absence of recovery lessened by recovery during and immediately following deformation. Strain hardening takes place through movements of large groups of atoms acting to a degree as an entity. A complex microstress system is developed. Recovery is a diffusive-type process involving single atoms or, at most, only a few atoms in concerted action. This is accompanied by elastic strain relaxation. Recovery is strongly temperature-dependent. The driving energy for the process resides in thermal fluctuations of the atoms augmented by the energy of the stress field producing the deformation, and by that produced by the deformation. At very low temperatures recovery is essentially absent, or very small. At high temperatures in the hot working range, recovery is so rapid that deformation results in no appreciable net hardening. The interplay of strain-hardening and recovery results in differing end structures depending on the temperature and the strain rate of the process. For example, there is a prominent difference in the spacing of the slip markings and in the detail of X-ray diffraction patterns. High temperatures and slow strain rates favor a wide spacing of slip markings and sharp reflections.',' This suggests that change in rate of the recovery concurrent with deformation modifies to a degree the manner of the deformation. Different slip systems may operate. For very slow rates, the mechanism may, as has been suggested, differ from that of "classical slip." In the experimental work that follows the findings will be interpreted from the point of view developed in this section. Recovery and Electrical Conductivity Electrical conductivity is a sensitive indicator of the state of strain in a material3' * and was used in this investigation. The materials studied in wire form were: OFHC copper, a high purity (99.99 pct) and 2s grade aluminum, and grade A nickel. Wire annealed and initially 0.050 in. in diam was drawn through diamond dies in a simple holder maintained at 25", —75°, or — 195°C. Precautions were taken that for each material all conditions of drawing were the same except the temperature. Electrical resistivity was measured for each condition at 20°C. This was done as soon as practical after drawing to minimize possible room temperature recovery. The data for the four materials are plotted in Fig. 1 in terms of the change in resistivity for drawing at the several temperatures. The effect of temperature of drawing on the increase in resistivity is most pronounced for copper. Drawing at 25°C to 80 pct reduction in area increases the resistivity approximately 2 pct; drawing at —195°C to the same reduction increases the resistivity about 6 pct. The increase in resistivity" & in this metal is due to extra scattering of conduction electrons originating in the lack of orderly arrangement of the atoms induced by the deformation process. It has been suggested

Jan 1, 1953

By M. J. Stewart, F. Weinberg

PREVIOUS investigations into convective flow in molten metals have examined systems in which the length-to-height ratio of the enclosure is large, usually using long graphite boats.''2 In normal castings this ratio is generally small and it is not possible to extrapolate the results of previous work to casting configurations. The present investigation was undertaken to examine liquid flow in systems of low length to height ratio. The apparatus used for the experiments is shown in Fig. 1. The molten zone of tin is contained in an aluminum mold of inside dimensions 6.4 by 6.4 by 0.31 cm. A heating block containing a resistance coil is attached to one end of the mold and a cooling block is attached to the other end. These are used to produce a wide range of temperature differences across the mold. The mold is contained in a stainless steel tank which is heated from the outside by resistance heaters to give the required operating temperature during the experiment. Temperature measurements in the molten tin are made with an iron-constantan thermocouple PT equal to 5 atm. The phase diagram derived by us was the same as that found by Krivsky and Schuhmann, except for a similar discrepancy in one boundary line representing the gas composition in equilibrium with SO a. The vapor pressures of the pure components can be neglected at 1700°K, 1 atm and 2500°K, 5 atm. There-fore, in these cases PT is essentially the total pres-sure in the system. This is valid for all the diagrams presented by Krivsky and Schuhmann except for the one at 2500°K and 0.01 atm, where the vapor pressure of silicon is comparable to PT.6 This work was supported in part by the American Iron and Steel Institute. 'C. E. Wicks and F. E. Block: Thermodynamic Properties of 65 Elements, U S. Bur Mines, Bull. No. 605, 1963. R. P. Elliott: Constitution of Binary Alloys, 1st supplementary, McGraw-Hill Book Co., New York, 1965. 3W. A. Krivsky and R. Schuhmann, Jr.: Trans. TMS-AIME, 1961, vol. 221, pp. 898-904. 4M. F. Ancey-Moret, M. Olette, and F. D. Richardson: Mem. Sci. Rev. Met., 1964, vol. 61, pp. 169-76. 5J . F. Elliott and M. Gleiser: Thermochemistry for Steelmaking, vol. 1. Addison-Wesley, Reading, Mass. 1960. 6 J.Drowart, G. De Maria, and M. G. Ingraham: J. Chem. Phys., 1958, vol. 29, p. 1015. probe of 0.4 mm diam inserted into the open top of the mold. Radioactive sn113, in the form of a small sphere, was gently lowered into the molten tin at point A in Fig. 1 THERMOCOUPLE> WATER ARGON END HEATER QUENCH ARGON Fig. 1—Apparatus for detecting liquid flow. Point A indicates position the tracer was introduced.

Jan 1, 1970

By G. A. Koehler

The Hiawatha mine, a producer of iron ore, is near the City of Iron River, in Iron County, Michigan. This mine, operated by The M. A. Hanna Co., has been chosen by the author as a typical iron mine in this district that is using the cut-and-fill mining system. General Geology The Hiawatha ore body lies along the steeply dipping north limb of a syncline that closes around on the east end to form a bathtub-shaped structure. Throughout this major structure there are very many smaller wrinkles or folds, which vary the direction as well as the thickness of the ore body. The total length of the ore formation with minor interruptions so far developed is about 3600 ft. The ore body throughout the entire present depth of 2000 ft. has averaged approximately 40 ft. in thickness; it is somewhat narrower at the upper levels and shows a marked increase in thickness on the bottom levels. This increase in thickness probably indicates that the bottom of the trough is near or that a major monoclinal fold has been encountered, which flattens the ore and gives apparent horizontal thickening. The footwall of this ore body is generally a black, graphitic slate, very rich in iron sulphides, quite combustible when exposed to air, and very treacherous to handle during mining operations. The hanging wall, in most places, is a cherty, fairly hard iron formation, and in general stands quite well ill open stopes. There are places, however, where black slate is found in both the hanging wall and the footwall, which makes mining difficult, as the black slate tends to slough off after being exposed to air. In these places it becomes necessary to leave a scale of ore 5 to 6 ft. thick on both hanging wall and footwall, which is not recovered. Fig. I shows a typical vertical cross section through the Hiawatha ore body. Type of Stoping Because of the size and shape of the ore body and the hard, dense structure of the ore, sublevel stoping has been an efficient and economical method of extracting the ore. This method, with various minor changes, has been used since the mine was opened in 1893. There have never been any stope-filling operations, however, until recent years, when crushing of pillars and black-slate fires forced the issue and methods were arrived at to fill all stopes. Development of Deeper Levels Shaft sinking to deeper levels in development operations has been done successfully by two methods without interference or hindrance to production. These systems are: I. Sinking an auxiliary winze a short distance from the shaft, then drifting back to the shaft location and putting up a pilot raise through to the bottom of the shaft. This raise is always in the manway compartment of the shaft, so that no work is ever done under a cage or skip. A 20-ft.

Jan 1, 1946

A general discussion on the subject of Crushing and Grinding was held on August 24, 1927, during the meeting of the Institute at Salt, Lake City. G. H. Clevenger, Chairman of the Milling Methods Committee of the Institute, made the following introductory remarks "Several years ago, a number of us felt that the time was ripe for a fundamental investigation of crushing and grinding. At the outset, it appeared that the first thing we needed was a yardstick whereby grinding efficiency could be measured. "The U. S. Bureau of Mines has been most generous in cooperating with the Milling Methods Committee in getting this work well undrr way. John Gross, who served the committee for a year as secretary, was placed in charge of the investigation and has been ably assisted by S. R. Zimmerley. "Messrs. Gross and Zimmerley will now present the Progrcss Report covering their work to date." Progress Report J. Gross, Salt Lake City, Utah.—-This has not been a simple problem, and when I tell you that me can give you a verbal report of what we have done during three years in five or ten minutes you will realize that we have necessarily done a great deal of work that dors riot appear in the final results. We have been in a virtually new field. Therefore, we have had to feel our way at each step so as to be fairly sure we were correet. The first thing that was realized in this problem ass that unless we had some method of measuring what happened when we crushed an ore, we might as well quit and not go further. we realized that if we tried to measure the results of crushing or grinding by the ordiuary sizing test we would not get anywhere. We could sieve down to say 200 mesh fairly accurately; we could even go one sieve size further with a little less accuracy, and then we were absolutely lost, unless we went into microscopic measurement. Even this latter method would not have been satisfactory. The question was: what were we going to measure? If we were going to measure size, we might measure one particular dimension of a particle and be off 200 per cent. or more. If we tried to measure surfaee, we would have something tangible; the surface of a particle, no matter how irregular, is something definite, far more definite than any linear dimension would be. Moreover, if we could measure surface, we would immediately have a method of checking the Rittinger law as to whether it was or was not correct. So the first thing that we sought. was a method whereby we could measure surface on crushed particles or, we will sag., on particles; it would not make any difference whether crushed or not. It then became a question of what kind of particles we were going to use in obtaining a method for surface measurement. We realizcd that it would be rather a corn-

Jan 1, 1928

By Frank G. Miller, Ralph A. Seban

Problems relating to thermal methods of oil recovery have been given increasing attention during the past year. The nature of the physical and chemical processes underlying thermal recovery are not yet well understood. The need for pertinent basic information is recognized generally. One of the processes involved is the transfer of heat by conduction in oil reservoir rocks containing moving reservoir fluids. This technical note deals with heat conduction in a laboratory sand column filled with a moving hydrocarbon and therefore should be contributory to the presently inadequate fund of knowledge on thc over-all subject. In an analysis of the flow of vaporiziny propane through a horizontal column of sand, one of the present author9 demonstrated correspondence between thc measured values of pressure along the length of the column and those predicted by a theory which postulated that the enthalpy of the propane was invariable along the length of the column. This postulate way implied as a deduction from the principle of conservation of energy but the details of the deduction were not given. The specific purpose of this technical note is the presentation of these details both as an adjunct to the earlier analysis and to supply some additional details on the processes involved in this type of flow. Results disclose that rates of heat transfer by conduction, in the direction of fluid motion, generally would be small for systems of the type studied. FUNDARIENrAI. ENERGY EiQUAI-[ON The principle of conservation of energy may be applied to a section of the porous material on the basis that heat transfer, like fluid flow, is possible in the x direction only. Transfer of heat in the y and z di'rections is prevented by suitable insulation or by taking the medium as infinite in directions perpendicular to the x axis. For steady state heat conduction and conditions of steady fluid flow the principle of conservation of energy specifies that any reduction in the enthalpy rate increases the flow of heat in the same proportion. Stated mathematically, dH KA 87 —- = - — - — U) * dx W dx........ Lvhere is the flow of heat in the positive x W dx direction. TEMPERATURE DISTRIBUTION IN DIRECTION OF FLOW The column of sand used in the previoi~sly described research is divisible into three different flow regimes. each of which may be analyzed independently. In the entrance region the liquid approaches the saturation state. A transition region is postulated following the transverse section at which the saturation state of the liquid is attained. In this transition region gas and liquid phases are both envisioned as present, with no vaporization occurring and with fluid movement restricted to the liquid. The upstream end of the exit region is at the transverse section of the column where the fraction of the pore space occupied by gas first becomes great enough for gas flow and where vaporization of the liquid begins. Fig. 1 indicates schematicaHy the thermodynamic processes undergone hv the fluid in the three flow regimes.

Jan 1, 1956

By John K. Hammes

INTRODUCTION From the point of view of the consumer, the cost of mineral commodities might be viewed as the total price industry pays for mine output. Similarly, the mining company engaged in the ongoing business of finding, developing and operating new mines is concerned very much today with the distribution of all the revenue paid for mine output and not just with its direct or controllable cash costs. In other words, what will the labor force, the supplier of energy, explosives, and reagents, and the supplier of government services, management, capital, and the ore reserves receive for their contribution to production? A significant amount might be categorized as what economic theory classifies as economic rent. Mining companies are finding it increasingly difficult to negotiate concession or work contracts to develop ore reserves in lesser developed countries, just as they find themselves accused of taking excess profits in their own countries. The principal difficulty in reaching mutually satisfactory agreements in many cases is the difficulty in establishing an acceptable distribution of income, in particular the distribution of economic rent. (See Chapter 7). Further, it is logical that the decisions and policies of the interested parties be they government or industry should be and will be influenced by their competitive position relative to determining the distribution of economic rent. The raising of finance for a mining project today may be inhibited by too biased an extraction of economic rent by entities other than the risk takers, the mining companies. ECONOMIC RENT A British economist (Ricardo D. McGraw-Hill, 1955) first proposed the concept of economic rent (See Drechsler, H.D. earlier in this Chapter.) Ricardo pointed out that no rent would be paid until land became scarce. Ricardo also stressed the distinction that the price of farm produce was not higher because of the rent paid on farm land but that the rent or price of more fertile farm land was higher because the price of produce is higher. This led to the concept that economic rent is not a cost and that it is therefore not price determining, but is itself price determined. Economic rent may be viewed as a cost if cost is equated with the concept of payment or distribution of income to any factor used in production. However, if it is a criterion that a payment must be made to the supplier of the factor of production, then economic rent is not truly a cost because it is not necessarily distributed or "paid" to the supplier. Consider the commodity A which occurs in three grades: Al, A2, and A3. Any grade may be used as a substitute for another grade in all of the industrial uses of A. However, in industries X and Y the different grades have different productivities. Table 1 shows the relative productivities of each grade in industries X and Y, and in all other industries [~]. [ ]

Jan 1, 1985