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By R. F. Mehl, M. Gensamer

Although a large number of X-ray photograms of cold-rolled steel have been published, two circumstances have led to the experimental work reported in this paper. The first is that no complete study, making use of the pole-figure method of Weverl, has been made of cold-rolled low-carbon steel. Wever has published pole figures of mild steel subjected to plane parallel deformation1, 2 and undoubtedly he has applied the method to cold-rolled mild steel, but the authors have not found these data published. Kurdjumow and Sachs3 have applied the method to the study of rolling and recrystallization textures in electrolytic iron; their results will be referred to later. The reason for this dearth of papers using the pole-figure method is obvious; a greater number of photograms and more analytical work (although this work is quite simple and straightforward) are required to construct a pole figure than most investigators have felt worth while, despite the fact that no other method of which the authors are aware will satisfactorily depict the orientations of the crystallites in a polycrystalline aggregate. The second circumstance that led to the work reported in this paper was the appearance of a paper by Wayne A. Sisson4 on the same subject. In his valuable study of the factors influencing the development of preferred orientations in cold-rolled mild steel, Sisson made one observation that required confirmation; that is, that at some stage of the rolling process the type of preferred orientation changed from the type that gives a photogram characterized as a "six-point" X-ray pattern to that known as a "four-point" pattern. A similar observation had been made by Tam-mann and Heinzel5, who studied the orientations developed in the rolling of steel by the specular reflection of light from the etched surface of rolled samples. These workers also found that the type of preferred orientation changed during the rolling process. The authors had some difficulty in seeing how this behavior could be consistent with modern views on the genesis of preferred orientations, particularly with those of Wever2

Jan 1, 1936

By J. W. Halley, T. S. Washburn

Rimming steels derive their name from their action during solidification in the molds. As a result of incomplete deoxidation, gas is evolved during freezing, and the metal has a characteristic rolling action. The metal rises along the mold walls and descends in the center of the ingot. This action keeps the top of the ingot open and the advancing rim of solid metal can be seen. The rim grows until the rolling action is insufficient to prevent the top of the ingot from freezing or until the action is stopped by capping the ingot with a heavy plate. The violence of the reaction and its characteristics, such as growth or drop in the molds, are determined by the composition, the furnace practice, and the deoxidation in the ladle and in the molds. Though rimming steels represent a large proportion of the amount of plain carbon steel made, they have received comparatively little attention in the technical literature. The manufacture of rimming steels has been described by Hibbard,1,2 Fleming,3 Pierce4 and Jackson.5 The reactions occurring during rimming have been considered from the theoretical point of view by Chipman and Samarin.? The structure of rimmed-steel ingots has been discussed by Nead and one of the present authors6 Meyer: in studying segregation, investigated the distribution of carbon, sulphur and phosphorus in rimmed-steel ingots by analyzing the rim and core of blooms. In their studies of segregation, the Joint Committee of the Iron and Steel Institute and the British Iron and Steel Federation9 studied distribution in one billet and one ingot of rimmed steel. A detailed study of the distribution throughout the ingot has not been published. The present investigation was undertaken to determine accurately the distribution of carbon, manganese and sulphur in normal rimmed ingots and the general distribution of other common elements present. The study included some work on the effect of pouring conditions and furnace practice on the distribution, though there is still much to be done in this direction. The structure of a rimmed ingot that has been split by dynamiting and a schematic diagram of the zones in the ingot are illustrated in Fig. 1.

Jan 1, 1938

By J. W. Halley, T. S. Washburn

Rimming steels derive their name from their action during solidification in the molds. As a result of incomplete deoxidation, gas is evolved during freezing, and the metal has a characteristic rolling action. The metal rises along the mold walls and descends in the center of the ingot. This action keeps the top of the ingot open and the advancing rim of solid metal can be seen. The rim grows until the rolling action is insufficient to prevent the top of the ingot from freezing or until the action is stopped by capping the ingot with a heavy plate. The violence of the reaction and its characteristics, such as growth or drop in the molds, are determined by the composition, the furnace practice, and the deoxidation in the ladle and in the molds. Though rimming steels represent a large proportion of the amount of plain carbon steel made, they have received comparatively little attention in the technical literature. The manufacture of rimming steels has been described by Hibbard,1,2 Fleming,3 Pierce4 and Jackson.5 The reactions occurring during rimming have been considered from the theoretical point of view by Chipman and Samarin.? The structure of rimmed-steel ingots has been discussed by Nead and one of the present authors6 Meyer: in studying segregation, investigated the distribution of carbon, sulphur and phosphorus in rimmed-steel ingots by analyzing the rim and core of blooms. In their studies of segregation, the Joint Committee of the Iron and Steel Institute and the British Iron and Steel Federation9 studied distribution in one billet and one ingot of rimmed steel. A detailed study of the distribution throughout the ingot has not been published. The present investigation was undertaken to determine accurately the distribution of carbon, manganese and sulphur in normal rimmed ingots and the general distribution of other common elements present. The study included some work on the effect of pouring conditions and furnace practice on the distribution, though there is still much to be done in this direction. The structure of a rimmed ingot that has been split by dynamiting and a schematic diagram of the zones in the ingot are illustrated in Fig. 1.

Jan 1, 1938

By L. E. Young

THE purpose of this paper is to start a discussion on: (1) methods of supporting the immediate roof, particularly of local areas of poor roof, during the mechanical loading of coal; (2) methods of increasing the rapidity of the movement of the mining faces so that, without the use of cribs, jacks, pack walls, and filling, there will be no danger to men, equipment,, and working places from the subsidence of the overlying strata. In the Pittsburgh district the application of mechanical loading devices in mines working in the Pittsburgh seam, has been retarded by the difficulties of holding and handling the draw slate. Normally the thickness of this slate averages about 8 in. but the hardness varies, and a cut of coal cannot always be loaded without placing posts or crossbars to support the slate. Locally, the slate may be thicker and special methods of support may be required. If crossbars are placed to hold the slate when the coal averages 5 ft. in thickness, the standard methods of transporting the coal cannot be followed, and the use of conveyors may be desirable. Very little information is available on the effect of increasing the rate of movement of advancing or retreating break lines or longwall faces. A few British publications have presented some data but all such data relate to longwall faces that have moved not more than 6 ft. a day. The most rapid rate of movement of longwall faces in Germany is 6 ft. per day. In the ordinary hand-loading mine in western Pennsylvania break lines on full retreat will move about 2 ft. per day. If loading machines are used under similar conditions the, rate of retreat may reach 4 ft. per day. In a mine where coal is loaded by hand on conveyors three shifts per day, the rate of retreat has ranged from 11 to 16 ft. per day, with an average of 13 ft. If mobile loading machines could be used to load on con-veyors on break lines, the rate of retreat should be increased substantially.

Jan 1, 1938

By A.A. Grubb, F.L. Wolf

The tests here described were rrlade to obtain information regarding costs, efficiency, etc. of baking cores in an oil-fired oven and two electric ovens, which were installed, early in 1920, in the core room of The Ohio Brass Co. and were used almost continuously during the rest of that year. Operating costs were rather high so a complete test was undertaken early in 1921. The ovens will be referred to as Nos. 1, 2, and 3. Nos. 1 and 2 are electric ovens of different makes while No. 3 is an oil-fired oven made by the firm that made No. 2. The three arc. equipped with similar ventilating devices. They have full-height doors at front and rear. The cores are placed on iron racks, which arc lifted by Cowan trucks and rolled into the ovens through the front doors. When baked, they are removed through the rear doors. Fig. 1 shows the general plan of the elecric ovens while Fig. 2 shows the plan of the oil-fired oven. Oven No. 1 is 91 in. (231 em.) wide, 83 in. (210 em.) high. and 79 in. (200 cm.) deep, inside measurements. The walls are constructed ill panels about 18 in. (46 cm.) wide and 43/16 in. (10.6 em.) thick, packed with blocks of insulating material and covered with sheet iron. Thc concrete floor oil which the owns set is 13 in. (33 cm.) thick. 4 recess about 5 in. (13 cm.) deep was cut in the floor to provide room for the heating units. Channels and plates supported on I beams serve as supports for the trucks and racks. The ventilating system consists of a Sirroco No. 11/2 fan, connected so as to draw air from the top of the oven or from the room, through damper C, and to deliver it either to the bottom of the oven or to the floor; the delivery is controlled by the vane A. The fan 2nd pipes are covered with 1/4-in. (6.3 mm.) asbestos-cement lagging. The heating units have a capacity of 91.4 kw. and arc located along

Jan 1, 1923

By F. L. Wolf, A. A. Grubb

The tests here described were rrlade to obtain information regarding costs, efficiency, etc. of baking cores in an oil-fired oven and two electric ovens, which were installed, early in 1920, in the core room of The Ohio Brass Co. and were used almost continuously during the rest of that year. Operating costs were rather high so a complete test was undertaken early in 1921. The ovens will be referred to as Nos. 1, 2, and 3. Nos. 1 and 2 are electric ovens of different makes while No. 3 is an oil-fired oven made by the firm that made No. 2. The three arc. equipped with similar ventilating devices. They have full-height doors at front and rear. The cores are placed on iron racks, which arc lifted by Cowan trucks and rolled into the ovens through the front doors. When baked, they are removed through the rear doors. Fig. 1 shows the general plan of the elecric ovens while Fig. 2 shows the plan of the oil-fired oven. Oven No. 1 is 91 in. (231 em.) wide, 83 in. (210 em.) high. and 79 in. (200 cm.) deep, inside measurements. The walls are constructed ill panels about 18 in. (46 cm.) wide and 43/16 in. (10.6 em.) thick, packed with blocks of insulating material and covered with sheet iron. Thc concrete floor oil which the owns set is 13 in. (33 cm.) thick. 4 recess about 5 in. (13 cm.) deep was cut in the floor to provide room for the heating units. Channels and plates supported on I beams serve as supports for the trucks and racks. The ventilating system consists of a Sirroco No. 11/2 fan, connected so as to draw air from the top of the oven or from the room, through damper C, and to deliver it either to the bottom of the oven or to the floor; the delivery is controlled by the vane A. The fan 2nd pipes are covered with 1/4-in. (6.3 mm.) asbestos-cement lagging. The heating units have a capacity of 91.4 kw. and arc located along

Jan 1, 1923

By W. A. Backofen, N. E. Paton

DEFORMATION twinning contributes to the plasticity of titanium and has been reported on {10i2}, {1121},{11221}, {1123}, and {1124} planes.' Twinning on {1011} has also been reported in commercially pure titanium as a result of quenching from 1000', but no reports of such twins occurring as a result of deformation alone are known to the authors. In this study single crystals grown from iodide titanium were compressed parallel to the c axis over the temperature range 25 to 800°C. From it has come evidence that (10il) deformation twins are the most common type between 400" and 800°C. Crystals about 6 mm high with a rectangular cross section measuring about 3 by 4 mm were used in the tests. Chemical analyses showed them to contain 220 to 250 ppm 0 arid 10 to 15 ppm N. The long axes of the crystals were made parallel to the c axis while the sides were parallel to (i100) and (1120), as shown in Fig. 1. The crystals were compressed between polished alumina platens with no lubrication, in a vacuum of 2 x lom5 torr. Between 25' and 300°C the predominant twinning mode was (11221, which is to be expected for this stress state. There was an abrupt change at 400°C, however, and from that temperature up to 800°C twinning took place almost entirely on (10i1 ). Twin planes were determined by the two-surface rnethd. Measurement of the crystal orientation within several (10i1) twins by means of polarized-light4 and transmission electron microscopy showed that the lattice reorientations were the same as those previously reported for the (10i l ) twinning in magnesium.' This reorientation is consistent with The calculated (10il ) twinning shear for titanium is 0.10, while that measured from surface offsets on a (1150) face, which is parallel to the shear direction, was about 0.15. Slip within the twins, or in the matrix adjacent to the twins, probably accounts for the larger measurement. Slip in both locations was judged to have occurred from observations on foils in the electron microscope.

Jan 1, 1970

By Paul D. Merica, L. J. Gurevich

In connection with other investigations on the properties and characteristics of chilled-iron car wheels, the question as to the best range of annealing temperatures was raised. Chilled-iron wheels are cast with a chill against the tread and the inside of the flange, the remainder being in sand. The composition of the metal is so chosen that under these conditions the trcad and inside of the flange will show white iron to a depth of from 5/8 to in. (15 to 22 mm.), the remainder of the wheel becoming graph itized or gray. In order to relieve the stresses set up during cooling under such drastic conditions, the wheels are stripped from the mold, while still red hot, piled in a soaking pit, and allowed to cool very slowly from their temperature at stripping. Obviously the most suitable temperature for this annealing is the highest at which no formation of graphite occurs within the white tread and at which the wheels can be stripped from the molds. Inasmuch as no direct determinations have been made of the temperatures at which the formation of graphite takes place in white iron of compositions used in car wheels upon annealing, it was considered worth while to determine these temperatures as a means of establishing the maximum temperatures at which the annealing of the wheels may be carried out. In the course of this work, some incidental observations were made which are of interest in connection with the theory of graphitization in whitc iron. The literature on the subject of the formation of graphite both during cooling at casting and upon annealing is very extensive and can be best studied by reference to a comprehensive discussion of the subject by Hatfie1d.l Perhaps the most important contributions of previous investigation are the following: 1. Whether a given mass of molten iron becomes white, mottled, or gay upon casting, depends on its composition and its rate of cooling through the temperature range of solidification and immediately below. 2. The more rapid the rate of cooling, the less is the extent of graphitization.

Jan 1, 1920

By Walter Stephen

MODERN developments in alloy steels have been so rapid and diverse that engineering practice has scarcely had time to re-appraise all items of past experience wherein conclusions became axiomatic under conditions that no longer apply. A case in point is the wet screening of abrasive ores within the fine range still looked upon as belonging exclusively to classifiers. By virtue of the development of stainless steels (18 per cent Cr, 8 per cent Ni), it is no longer true that classifiers are more economical than screens in closed circuit with large-tonnage ball or rod mills at 65 or even 100 mesh, provided a selective overgrind of the higher-gravity constitu-ents of the ore is not desired. Wet screening of ores is a highly corrosive as well as an abrasive opera-tion, because of the inevitably high degree of oxygenation of the water. Screen cloth of ordinary or spring-steel wire almost at once becomes pitted and covered with a film of corrosion products, which causes rust-blinding of fine-mesh screens and offers high frictional impedance to the flow of ore particles. On the other hand, the well-known properties of stainless steel enable it to maintain a bright surface, and so to eliminate blinding either by ore particles or by corrosion products. That a large-tonnage mill grinding hard ore to 65 mesh can be oper-ated economically in closed circuit with screens instead of classifiers has recently been demonstrated at the Copper Cliff concentrator of the International Nickel Co.; in the test to be described, a 6 ½ by 12 ½ -ft. Marcy rod mill ground from 625 to 820 tons. of ore per day to approxi-mately 65 mesh, operating in closed circuit with three 4 by 5-ft. Hum-mer screens, without any classifier.. Stainless-steel cloth gave an indicated life of several months. There was no difficulty with blinding, yet dilution and spray water could be controlled so that screen undersize averaged not less than 40 per cent solids. In contrast, ordinary carbon steel was pitted and blind with rust within 40 hours.

Jan 1, 1938

By N. Shiarella, Coleman D. Hunter, I. B. Browning

The price of crude oil in Kentucky did not change very much during 1935 until the latter part of the year, but more stabilized conditions and an increase in consumption of natural gas made the year a more profitable one than the year 1934. The production of crude oil in 1935 amounted to 5,175,054 bbl. as compared to 4,814,904 bbl. in 1934. This increase was due to two developments: (1) the application of new recovery methods to old fields, which, although in operation in the state for several years, had never before reached a point where the production obtained, while highly satisfactory where applied, was sufficient to influence the totals in the production figures; (2) the development of the Fordsville area in Ohio County, where shallow depths enabled the operators to complete the development of a large number of wells with a good flush production probably accounted for most of the increase. A map showing the location of oil and gas pools in Kentucky was published last year.' The fields of eastern Kentucky, producing from the Weir sand of the Waverly group of the Lower Mississippian series, in Johnson, Magoffin, Lawrence and Elliott counties, produced 826,001 bbl. of oil in 1935 as compared with 902,247 bbl. in the previous year. No new wells were drilled in this area during the year, except those for use as intake wells in repressuring. What has been done in the few small areas within these fields in the way of repressuring has produced results sufficient to a great extent to offset the natural decline of all the field. It is probable the drilling of new wells and construction of new plants will be such as to cause these fields to produce more in 1936 than was produced in the past year. This is based on the fact that the sand is highly favorable for repressuring and the results obtained show from 300 to 500 per cent increase in production. From the "Corniferous" pools, where production is from the Cornifer-ous limestone of Devonian age, in Lee, Powell, Estill, Wolfe, Morgan and

Jan 1, 1936

By I. B. Browning, N. Shiarella, Coleman D. Hunter

The price of crude oil in Kentucky did not change very much during 1935 until the latter part of the year, but more stabilized conditions and an increase in consumption of natural gas made the year a more profitable one than the year 1934. The production of crude oil in 1935 amounted to 5,175,054 bbl. as compared to 4,814,904 bbl. in 1934. This increase was due to two developments: (1) the application of new recovery methods to old fields, which, although in operation in the state for several years, had never before reached a point where the production obtained, while highly satisfactory where applied, was sufficient to influence the totals in the production figures; (2) the development of the Fordsville area in Ohio County, where shallow depths enabled the operators to complete the development of a large number of wells with a good flush production probably accounted for most of the increase. A map showing the location of oil and gas pools in Kentucky was published last year.' The fields of eastern Kentucky, producing from the Weir sand of the Waverly group of the Lower Mississippian series, in Johnson, Magoffin, Lawrence and Elliott counties, produced 826,001 bbl. of oil in 1935 as compared with 902,247 bbl. in the previous year. No new wells were drilled in this area during the year, except those for use as intake wells in repressuring. What has been done in the few small areas within these fields in the way of repressuring has produced results sufficient to a great extent to offset the natural decline of all the field. It is probable the drilling of new wells and construction of new plants will be such as to cause these fields to produce more in 1936 than was produced in the past year. This is based on the fact that the sand is highly favorable for repressuring and the results obtained show from 300 to 500 per cent increase in production. From the "Corniferous" pools, where production is from the Cornifer-ous limestone of Devonian age, in Lee, Powell, Estill, Wolfe, Morgan and

Jan 1, 1936

By Arthur Crowfoot

As a result of work done, on an operating scale, in its experimental mill during 1920 and 1921, the plan here given was proposed for crushing, grinding, and concentrating low-grade sulfide ores at the Morenci Branch of the Phelps Dodge Corpn. The flow sheet is designed to recover 85 per cent. of the total copper and 95 per cent. of the sulfide copper from ore assaying 2.0 per cent. total copper and 0.3 per cent. acid soluble copper. At present, 1400 tons are treated daily under the conditions of this flow sheet, two 700-ton units having been installed in the 4000-ton concentrator. The run-of-mine ore will be reduced, in the dry-crushing plant, to —6 mesh and then delivered to the concentrator feed bins. The crushing will be done in three stages: (I) Run-of-mine ore is reduced to —2 1/2-in. in a gyratory crusher; (2) the 3-in. ore is reduced to 1/2-in. in 72 by 24 in. rolls; (3) the 1/2-in. ore is reduced to — 6-mesh size in 54 by 24 in. rolls. The 54-in. rolls will operate in closed circuit with electrically vibrated screens, "Hum-mer" type, a conveyor circuit being used to carry the roll product to the screens. Sufficient screens will be installed to pass 1.5 tons undersize per hour per square foot. Fine-ore bin chutes will be piped for compressed air, which will be used intermittently to keep the ore feed in motion if it tends to arch or pipe in the bins. The Concentrator The concentration of the copper and iron sulfides in the ore will be effected in three stages: I. By direct treatment of the —6-mesh crushing-plant product on primary tables, which will recover about 35 per cent. of the original copper content of the ore in a clean granular concentrate. The table middling will be dewatered in Dorr rake classifiers and ground to — 48-mesh in ball mills operating in closed circuit with circulating classifiers.

Jan 1, 1923

By Arthur Crowfoot

As a result of work done, on an operating scale, in its experimental mill during 1920 and 1921, the plan here given was proposed for crushing, grinding, and concentrating low-grade sulfide ores at the Morenci Branch of the Phelps Dodge Corpn. The flow sheet is designed to recover 85 per cent. of the total copper and 95 per cent. of the sulfide copper from ore assaying 2.0 per cent. total copper and 0.3 per cent. acid soluble copper. At present, 1400 tons are treated daily under the conditions of this flow sheet, two 700-ton units having been installed in the 4000-ton concentrator. The run-of-mine ore will be reduced, in the dry-crushing plant, to —6 mesh and then delivered to the concentrator feed bins. The crushing will be done in three stages: (I) Run-of-mine ore is reduced to —2 1/2-in. in a gyratory crusher; (2) the 3-in. ore is reduced to 1/2-in. in 72 by 24 in. rolls; (3) the 1/2-in. ore is reduced to — 6-mesh size in 54 by 24 in. rolls. The 54-in. rolls will operate in closed circuit with electrically vibrated screens, "Hum-mer" type, a conveyor circuit being used to carry the roll product to the screens. Sufficient screens will be installed to pass 1.5 tons undersize per hour per square foot. Fine-ore bin chutes will be piped for compressed air, which will be used intermittently to keep the ore feed in motion if it tends to arch or pipe in the bins. The Concentrator The concentration of the copper and iron sulfides in the ore will be effected in three stages: I. By direct treatment of the —6-mesh crushing-plant product on primary tables, which will recover about 35 per cent. of the original copper content of the ore in a clean granular concentrate. The table middling will be dewatered in Dorr rake classifiers and ground to — 48-mesh in ball mills operating in closed circuit with circulating classifiers.

Jan 1, 1923

By M. T. Hepworth, R. Schuhmann

Hydrogen solubility measurements were made on a series of Ti-O alloys, and a portion of the 800°C isotherm for the Ti-O-H system was determined. Activities of oxygen and titanium were calculated from the experimentally measured activities of hydrogen, using ternary Gibbs-Duhem integrations. The data were extrapolated to obtain thermo-dynamic properties of binary Ti-O solutions. The thermodynamic properties of a Ti-0 solid solutions thus determined were compared with properties predicted from interstitial models. The Gibbs-Duhem equation, long used to calculate activities in binary systems, has been extended. recently to ternary systems.1"3 When the activity of one component in a ternary system is known as a function of composition, the activities of the other two components can be calculated from that of the first by an integration procedure, provided certain restrictions are observed on the path of integration. Often the activity of one component in a ternary system can be measured readily, while the activities of the other two components are difficult to measure. Gibbs-Duhem calculations are especially applicable to such systems. The further possibility arises of developing a new and indirect approach to activity measurement in binary systems. The proposed new method for binary systems consists of adding varying amounts of a third component whose thermodynamic behavior is easily measured. The activity of this added component is determined experimentally as a function of composition. Then the activities of the two components of the original binary system are calculated by Gibbs-Duhem integrations and extrapolations. For the work to be described in this paper, the Ti-0-H system was selected as a promising system for evaluating this approach. Investigations of the Ti-H system have shown that the activity of hydrogen can be measured precisely with little difficulty. Pure titanium transforms from a low-temperature a form (hcp) to a higher-temperature /3 firm (bcc) at 882.5"~ The phase transformation occurs in a temperature range where hydrogen pressures are easily measured. Also, the diffusivity of hydrogen at temperatures above 600" or 700"~ is large enough that equilibrium can be reached within a reasonable time.

Jan 1, 1962

By W. C. T. Yeh, T. G. Oakwood, A. A. Hendrickson, R. H. Hammar

The objective of the research is to determine whether solid-solution strengthening effects observed in dilute solutions of silver can be accounted for by the influence of the solute addition on the dislocation structure oj- the crystals. The additions of both tin and indium produced only small changes in the dislocation densities and arrangements in silver crystals. However, as found previously, small solute additions have large effects on the tensile properties; the inj-luence of the tin and indium additions on the temperature dependence of the flow stress and the easy-glide range is especially strong; It is concluded that the indirect strengthening effect of the solute due to variations in the dislocation density as proposed by Seeger is of minor importance and that solute atom-dislocation interactions are responsible for the observed strengthenirzg effects. The experimental results were combined with those of Rogausch to test the concenlvatiorz dependence of solute strengthening. Both the first and one-half power dependences of the critical resoleed shear stress on concentratiorz fail in very dilute solutions. THE objective of the research is to determine whether solid-solution strengthening effects observed in dilute solutions of silver can be accounted for, at least in part, by the influence of the solute addition on the dislocation structure of the crystals. It is recognized that the addition of solute atoms may influence the strength properties of a metal through both "direct" and "indirect" effects. The former refer to the strengthening mechanisms that result from the interaction of solute atoms with dislocations; in the latter case, the strengthening effects arise as a result of solute's influence on quantities such as dislocation density, dislocation arrangement, stacking-fault energy, diffusivities, the elastic constants, and so forth. It is clear that the correct interpretation of solid-solution strengthening phenomena cannot be given until the importance of indirect strengthening effects is properly evaluated. In the particular case of close-packed metal crystals, Seeger showed that solute strengthening effects in dilute solutions of copper and silver might be accounted for by an increase in dislocation density due to the addition of the solute. Seeger's argument was that the strengthening effects extrapolated from more concentrated solutions indicate that small concentrations of impurities raise the critical resolved shear stress much more than is predicted by a concentration-independent dislocation density. The above idea was a very reasonable one. The dislocation theories of work hardening of Taylor,2 Cot-trell, 3 Mott, 4 and seeger5 had already associated the increased flow stresses with increased dislocation densities in deformed metals; investigations of the dislocation structure of metal crystals had provided a logical basis for expecting an increased dislocation density in crystals containing impurities (see for example, Ref. 6). The numbers involved seem reasonable, too. It can be expected that the flow stress of the crystal would increase as the one-half power of the dislocation density.' Solute additions of 1 at. pct to metal crystals result in strength increases by factors in the range of three to ten. If one assumes that the strengths of the pure metal crystals are determined by their dislocation densities, then dislocation-density increases of one to two orders of magnitude as a result of solute addition would be required to account for the observed strengthening—not an unreasonable expectation. In addition to the effect of the solute addition on dislocation density, one might also anticipate important strengthening contributions to result from the solute's influence on the dislocation arrangement. Parker and washburns have reviewed a number of experimental evidences which show important strengthening effects due to the presence of subboundaries. Further, lattice strains due to impurity segregation would be expected to influence the distribution as well as the dislocation density of the as-grown crystal. As pointed out in the reviews of Chalmers,6 Elbaum,9 and winegard,lo micro segregation of impurities occurs at all interfaces of crystals in cellular growth; the impurity gradient results in lattice strains which can be reduced with the presence of dislocation arrays in the region of the impurity gradient. Hence, one would expect the presence of a solute to favor the formation of dislocation subgrain structures and that the subgrains would have an important influence on the strength of the crystal. The experimental observations that concern the possibility of an important strengthening contribution through the influence of the solute on the dislocation density or arrangement are not in agreement. Haasen has reviewed the observations of Meakin and Wils-dorf,12 Howie,13 and Bocek 36 and concluded that the solute's influence on dislocation density is not sufficient to account for strengthening effects in concentrated solutions but might, as seegerl suggested, make an important contribution in very dilute solutions. On the other hand, Hendrickson and Fine 14 concluded that changes in the dislocation density and dislocation width accounted for the solid-solution strengthening effects observed in silver-based aluminum solid solutions. Goss et a 1.I5 observed dislocation arrays in Ge-6 at. pct Si, Ge-0.2 at. pct Sn, and Ge-0.2 at. pct B crystals that were not observed in germanium crystals of

Jan 1, 1968

By E. Keller

Some years ago the writer's attention was called to the fact that rolling-mill scales from auriferous copper do not have the gold content proportional to the gold contained in the copper from which they are derived.' So, later, he made a few simple experiments, which are described herewith. In order to test the gold movement on the changing surface of a copper plate under the influence of oxidation, plates 6 in. by 8 in. by 1 in. (15.2 cm. by 20.3 cm. by 2.5 cm.) of converter copper, refined, of 99.29 per cent. purity, and leaded copper with a content of 95.65 per cent, copper, were obtained. Ten cuts were made through the thickness of the plates along the longitudinal center and the silver and gold content ascertained in each of the 10 samples in both plates, as shown in the accompanying illustration. The assay results, recorded in Tables 1 and 2, show that the gold distribution is relatively uniform in both plates, while silver has greater variations, especially in the leaded plate. The plates subjected to oxidation were placed in a red-hot muffle for about 2 hr., two small scarifiers being used as supports. Arranged alongside each other, they nearly filled the width of the muffle, and so were practically under equal conditions of heat and draft. The accurate observance of the time was not possible, however, because of other work, but the time of the exposure of each plate in a pair was identical. At the end of each period, each plate was plunged into water and completely freed of scale, which was carefully collected, dried, weighed, and analyzed. Samples of the plates had also been taken of the molten metal and duly analyzed. Table 3 contains the analytical data for the two plates and their several sheets, or layers, of oxide scales. These figures show that in the oxidations of the converter copper the silver is fairly uniform, while the gold is highly irregular and especially low in the first layer. In the oxidations of the leaded copper plate, the silver and gold contents are variable; however, this irregularity is due to two different causes. The

Jan 1, 1919

By Walter M. Stephen

Modern developments in alloy steels have been so rapid and diverse that engineering practice has scarcely had time to re-appraise all items of past experience wherein conclusions became axiomatic under conditions that no longer apply. A case in point is the wet screening of abrasive ores within the fine range still looked upon as belonging exclusively to classifiers. By virtue of the development of stainless steels (18 per cent Cr, 8 per cent Ni), it is no longer true that classifiers are more economical than screens in closed circuit with large-tonnage ball or rod mills at 65 or even 100 mesh, provided a selective overgrind of the higher-gravity constituents of the ore is not desired. Wet screening of ores is a highly corrosive as well as an abrasive operation, because of the inevitably high degree of oxygenation of the water. Screen cloth of ordinary or spring-steel wire almost at once becomes pitted and covered with a film of corrosion products, which causes rust-blinding of fine-mesh screens and offers high frictional impedance to the flow of ore particles. On the other hand, the well-known properties of stainless steel enable it to maintain a bright surface, and so to eliminate blinding either by ore particles or by corrosion products. That a large-tonnage mill grinding hard ore to 65 mesh can be operated economically in closed circuit with screens instead of classifiers has recently been demonstrated at the Copper Cliff concentrator of the International Nickel Co.; in the test to be described, a 6½ by 12½-ft. Marcy rod mill ground from 625 to 820 tons of ore per day to approximately 65 mesh, operating in closed circuit with three 4 by 5-ft. Hum-mer screens, without any classifier. Stainless-steel cloth gave an indicated life of several months. There was no difficulty with blinding, yet dilution and spray water could be controlled so that screen undersize averaged not less than 40 per cent solids. In contrast, ordinary carbon steel was pitted and blind with rust within 40 hours.

Jan 1, 1939

By Walter M. Stephen

Modern developments in alloy steels have been so rapid and diverse that engineering practice has scarcely had time to re-appraise all items of past experience wherein conclusions became axiomatic under conditions that no longer apply. A case in point is the wet screening of abrasive ores within the fine range still looked upon as belonging exclusively to classifiers. By virtue of the development of stainless steels (18 per cent Cr, 8 per cent Ni), it is no longer true that classifiers are more economical than screens in closed circuit with large-tonnage ball or rod mills at 65 or even 100 mesh, provided a selective overgrind of the higher-gravity constituents of the ore is not desired. Wet screening of ores is a highly corrosive as well as an abrasive operation, because of the inevitably high degree of oxygenation of the water. Screen cloth of ordinary or spring-steel wire almost at once becomes pitted and covered with a film of corrosion products, which causes rust-blinding of fine-mesh screens and offers high frictional impedance to the flow of ore particles. On the other hand, the well-known properties of stainless steel enable it to maintain a bright surface, and so to eliminate blinding either by ore particles or by corrosion products. That a large-tonnage mill grinding hard ore to 65 mesh can be operated economically in closed circuit with screens instead of classifiers has recently been demonstrated at the Copper Cliff concentrator of the International Nickel Co.; in the test to be described, a 6½ by 12½-ft. Marcy rod mill ground from 625 to 820 tons of ore per day to approximately 65 mesh, operating in closed circuit with three 4 by 5-ft. Hum-mer screens, without any classifier. Stainless-steel cloth gave an indicated life of several months. There was no difficulty with blinding, yet dilution and spray water could be controlled so that screen undersize averaged not less than 40 per cent solids. In contrast, ordinary carbon steel was pitted and blind with rust within 40 hours.

Jan 1, 1939

Prof. J. W. Richards, South Bethlehem, Pa. (in reply to several questions):—The cut representing some of the round briquettes which were under test shows that, at the brcaking-stresses, they break into lumps, and do not disintegrate into powder. The process is simplicity itself. I have followed the different processes for sintering iron-ores and flue-dust, am interested in all of them and wish them all success; but this process appealed to me strongly by its simplicity. It is in line with a method which I once tried to work out myself, but which never came to anything, namely, the agglomeration of fine iron-ores by means of Portland cement. As small an amount as 5 or 6 per cent. of Portland cement will enable fine ore to agglomerate and set into a briquette, but when I came to estimate the cost, I found it was a little too high. The Schumacher process, curiously enough, does not work well on iron-ore, but flue-dust thus treated does harden and give a strong briquette. In my attempts to find a reason for this difference, it has occurred to me that possibly the particles of flue-dust had been coated over with some sublimed silica, which might be an agent in the setting. But I found by experiment that neither sublimed silica, nor ordinary silica, nor ordinary alumina, would briquette. Chemically pure iron oxide briquetted quite hard; but the application of pressure is a necessary part of the operation. If simply moistened with the solution and allowed to stand it does not set hard, but the more pressure applied to it the hardcr it sets. Evidently the flue-dust particles must be pressed very close together in order to get the small proportion of active material wet, so that the particles shall act at their points of contact; for it is evident that only at the points of contact does the cementing action take place. The reactions are facilitated by considerable pressure, bringing the acting particles close together, and may go on with

Jan 1, 1913

By R. E. Cech, T. D. Tiemann

It has been shown that hydrogen may be made to serve as a rapid and eflicient reducing agent for Cu, Ni, Co, and Fe sulfides if a scavenging agent for hydrogen sulfide is intimately mixed with the sulfide particles being reduced. Accelerated reduction kinetics are demonstrated for nickel sulfide. Copper, nickel, and cobalt sulfides, when treated at certain temperatures in a combined reducing agent-scavenging agent system, are converted to voluminous masses of fibrous metal product. Studies have been carried out to determine the conditions which lead, on the one hand, to irregular poly crystalline fibers and, on the other, to long single crystal filaments a few microns in diameter. A mechanism is proposed to account for the formation of single crystal filuments. The sulfide minerals of Cu, Ni, Co, and Fe are an important source of these metals yet there has been comparatively little scientific effort devoted towards understanding reduction mechanisms of these minerals. This may be, in part, due to the fact that the most convenient reducing agents for carrying out such studies, viz., hydrogen and carbon, do not react appreciably with sulfides. We have found that the reaction of hydrogen with metal sulfides can be markedly accelerated by placing a scavenging agent for hydrogen sulfide in close proximity to the metal sulfide. A brief series of experiments demonstrating relative reduction rates is reported in this paper to illustrate the effect. With the reduction process thus accelerated we have observed an unusual type of reduction behavior on some of the sulfides investigated. Under certain conditions the metallic product of the reduction reaction takes the form of filaments growing outward from the sulfide particles. The present paper deals largely with efforts to classify the various types of growth forms observed. This study has shown that filamentary growths from sulfides take a much greater variety of forms than has heretofore been reported by Ercker,1 Hardy,2 and Nabarro and Jackson3 in their reviews of metallic growths from copper and silver sulfides. THERMODYNAMIC CONSIDERATIONS The thermodynamics for hydrogen reduction of metal sulfides is quite unfavorable. For the sulfides considered here equilibrium constants typically range from 10-3 to 10-5. These low equilibrium constants impose severe kinetic limitations on reduction since hydrogen sulfide must be transported out of the system at concentrations of only a few hundred ppm. Unless extremely high gas flow rates are employed the atmosphere surrounding any sulfide particle will always be essentially in equilibrium with the sulfide. If, however, one places an efficient scavenging agent for hydrogen sulfide in close proximity to the metal sulfide particles the concentration of H2S near the metal sulfide will be held to a very low value. This would permit the reduction reaction to proceed with little or no inhibition from a buildup of reaction product gas. It is well known that calcium oxide is capable of removing hydrogen sulfide from a hydrogen gas stream of low dew point.4 If a sufficient quantity of calcium oxide is mixed with the metal sulfide particles the reaction: CaO+H2S=CaS+ H2O [l] will substitute moisture in place of hydrogen sulfide in the gas stream and this will not affect, in a direct manner, the reaction: MeS +H2=Me + H2S [2] A convenient method of considering the thermodynamics of the combined reducing agent-scavenging agent system is to consider the atmosphere when the partial pressure of hydrogen sulfide is the same over both the metal sulfide and the scavenging agent, i.e., pH2S (1) =pH2S (2). As a consequence: pH2O (1) pH2(2) =K1K2 The chemical driving force for reduction will depend inversely upon the moisture content of the gas and will be 0 when, in the system, pH2O = pH2.K1K2. Table I lists values of the equilibrium constants for reduction and H2S scavenging reactions for a number of sulfides at several temperatures. Data are taken from Rosenqvist4,5 and Kelly.6 The equilibrium constant products calculated from this data show that the limiting level of gaseous reaction product has been increased by a factor of 10' to l04 as a result of substituting a reducing agent-scavenging agent system for a simple reducing agent system. One possible side effect which must be considered is the possibility that the moisture evolved in the scavenging reaction might cause the atmosphere in the system to be sufficiently oxidizing to favor the formation of oxide rather than metal. This possibility was examined by comparing the equilibrium constant products listed in Table I with equilibrium constants for hydrogen reduction of the respective metal oxides. It was found that for copper, nickel, and cobalt the combined reduction-scavenging reactions could not develop a sufficiently high oxidizing potential in the

Jan 1, 1970