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By Ottar Dragge, C. Danielsson, Bo Kalling

DISCUSSION, T. L. Joseph presiding L. F. Reinartz (Armco Steel Corp., Middletown, Ohio) —I would like to know, in the practical application of the Kalling process, what kind of a lining was used, how thick was the lining, and how much metal was treated at one time? S. Fornander (author's reply)—The rotary furnace is lined with a course of fireclay bricks 6 in. thick. This course is backed by 5 in. of insulation. The furnace has a capacity of about 15 tons. Mr. Reinartz—How was the ladle preheated? Mr. Fornander—As pointed out in the paper, the furnace was heated by a gas flame in the beginning of the experiments. During these first tests, however, the desulphurization was inconsistent. We think that this was due to the fact that iron droplets sticking to the furnace walls were oxidized by the gas flame. Now, the furnace is operated without preheating of any kind, and the results are much better. T. L. Joseph (University of Minnesota, Minneapolis, Minn.)—I might add one comment. This furnace was heated with a flame and for a time they had a little difficulty due to some residual metal in the rotating drum that would oxidize in between treatments and they found therefore, that it was very essential to drain the drum completely of metal so that they would not build up any ferrous oxide between treatments and they eliminated some of their erratic heats by maintaining those more reducing conditions. It was interesting to watch this operation. As soon as the drum started to rotate there was considerable flame, at least, at the time I saw it, that came out around the flanges, indicating there was quite a little pressure on the inside of the drum. W. 0. Philbrook (Carnegie Institute of Technology, Pittsburgh)—Is the reaction slag in the Kalling process liquid or solid, and how is it separated from the metal? Mr. Fornander—In the process there is no slag in the usual sense of the word. The lime powder does not melt during the treatment. After the treatment the lime is still in the form of a fine powder. It is separated from the metal by means of a piece of wood of suitable size placed within the furnace before it is emptied. D. C. Hilty (Union Carbide & Carbon Research Laboratories, Niagara Falls, N. Y.)—Dr. Chipman has given us some of his ideas in connection with a specific effect of silicon and silica on sulphur elimination and how silicon might interfere with desulphuriz- ing in the blast furnace. I wonder if he would like to elaborate on the possibility of a similar effect of silicon in the Kalling process? J. Chipman (Massachusetts Institute of Technology, Cambridge, Mass.)—Silicon does not interfere with the Kalling process. Anything that has strong reducing action is good for desulphurization. In these tests where the temperature was low compared to blast furnace temperatures, the silicon that is in the metal is a better reducing agent than the carbon. At high temperatures, carbon is the better. It is not the silicon in the metal that interferes with desulphurization, it is the silica in the slag. Mr. Joseph—I might add that the metal that was tapped from the drum after desulphurization was really at quite a low temperature. It was not measured, but I think it was well under 1300 °C, probably 1200" or a little above that. That was one of the difficulties, and I think there is no question about the fact that the Kalling process—in that it affects desulphurization between powdered lime, solid and liquid iron— is a reaction definitely between the solid lime and the liquid iron. E. Spire (Canadian Liquid Air, Montreal, Canada) — This Kalling process seems very interesting to us and after all it is only a mixing action that is taking place between the iron and the slag. We have attempted to do the same thing in another way. We have placed at the bottom of the ladle a porous plug through which we injected an inert gas. It can be nitrogen or argon. This plug is placed at the bottom of the conventional ladle and gas injected through the plug. That has appeared in our patent. To define this new type of treatment, I use the word gasometallurgy. I do not know if you like it, but it is a way of defining methods of treating metal using gases. What we do is exactly what is done in the exchange process in another way. We have a porous plug at the bottom with a high lime slag on top of the metal. Using this method, we have very good agitation of metal and slag, and with a small flow of gas, we can achieve a very strong agitation. For instance, in the 500 lb ladle, we use only 5 liters of gas a minute. We have an agitation compared to very rapidly boiling water in a pail. Moreover, the agitation can be controlled to create any amount of mixing desired. In a few minutes, with this method, the sulphur dropped from 0.58 to 0.11. These results have been improved since, and we have obtained results like 0.08

Jan 1, 1952

By B. H. Bergstrom, Will Mitchell, T. G. Kirkland, C. L. Sollenberger

IN a previous article' the authors outlined a study of the variables in rod milling and also reported data from a series of open circuit grinding tests on a massive limestone in a 30-in. x 4-ft end peripheral discharge rod mill. As a second part of the experimental program, an analysis is now presented for the 30-in. x 4-ft overflow rod mill grinding under identical conditions, except that discharge ports on the periphery of the mill shell have been sealed so that the products from the present series overflowed through a 9-in. diam .opening in the center of the end plate. A variance analysis has been made of the combined data for the two experiments, and performances of the two mills are compared here. Included in the first report' were descriptions of feed preparation, rod mill circuit, instrumentation and controls, and techniques used to evaluate data. Dependent and independent variables were defined, and variance analyses were made to test the relative significance of variables and to establish magnitude of error for the experiment. Significant data were plotted in various combinations, and conclusions were drawn from the graphs. The procedure and analysis in this series of tests follows the first tests and is not repeated. Data from the second series are recorded in Table I. Listed in the first three columns are the independent variables of feed rate (1000, 2000, 3000, 4000, and 5000 1b per hr), mill speed (50, 60, 70, 80, and 90 pct of critical), and pulp density (50, 60, 70, and 80 pct). The dependent variables, Pso, P100, reduction ratio, slope of the log-log sieve analysis curve, power demand, and Bond work index follow. Of these, only the reduction ratio and the Bond work index were analyzed for significance. Production of new surface as calculated from sieve analyses has not been included for this series because of the questionable assumptions that have to be made to satisfy the formulas involved. The large number of products obtained during the runs precluded the use of surface measurement techniques by the gas adsorption methods at this time; however, samples of all products have been stored for future reference. To test the consistency of the reporting of the sieved products, an averaged sieve analysis was calculated from the wet-dry plots obtained from the three product samples of each run. The resulting averaged analysis was plotted and the P80, selected. The relative deviations of the P80's from each of the three product samples with respect to the P80 of the averaged analysis were then calculated. In only two sets were the relative deviations (6.2 and 9.9 pct) considered excessive. In each of these two sets, one sieve analysis was obviously out of line; hence that analysis was ignored and new averages were computed. This reduced the relative deviations to 1.2 and 2.7 pct respectively. The relative deviations of the product analyses with respect to their averages ranged from 0.1 to 1.4 pct at 1000 lb per hr, 0.0 to 1.1 pct at 2000 lb per hr, 0.2 to 3.0 pct at 3000 lb per hr, 0.3 to 4.3 pct at 4000 lb per hr, and 0.5 to 5.2 pct at 5000 lb per hr. The relative deviation of the 80 pct passing point for 96 dry sieve analyses of the feed with respect to that of the averaged analysis was 7.6 pct. This slightly higher percentage can probably be attributed to a greater proportion of tramp oversize in a crusher product than is ordinarily found in a rod mill product. The last column on Table I lists the adjusted work index, which has been used as the measure of efficiency for the various combinations of operating conditions investigated. Efficiency increases as the index becomes lower. It was reported in the previous paper that the work indexes for the Waukesha limestone used in these experiments decreased as the product size decreased (as calculated from Bond grindabilities). That is, this limestone becomes easier to grind as the material becomes finer. This is unusual, because the work index for most materials as calculated from the Bond grindability has remained constant as the product size decreased or has increased slightly. Table II lists the results of Bond grindability tests at all mesh sizes from 3 to 200 and the work indexes calculated from them. To remove this variation of work index with product size from the data so that results would apply to any material of constant work index, the work index values shown in Table II were plotted against product size on log-log paper. From this curve (a straight line function in this case), the expected work index for the product size for each of the runs of the experiment was obtained. The work indexes as calculated from the reduction ratio and energy consumption were then divided by the corresponding expected work index. The results obtained are reported in percentages on Table I as adjusted work index and are actually percentages of the work index for the Waukesha limestone at the size in question. Multiplication of the work index value for a material of constant index by these percentages should allow the application of the adjusted work index curves to the material. Only the adjusted work index values, not the actual experimental values, were used for the variance analyses and for the graphs.

Jan 1, 1956

By Terkel Rosenqvist

THE desulphurization of molten iron and steel is a very complicated process. One way to arrive at a better understanding of this process is to break it down into several simpler chemical processes that can be studied individually in the laboratory. For a study of the different factors that influence the equilibrium distribution of sulphur between liquid metals and slags, several simpler equilibria may be investigated. One very important subject is the determination of the escaping tendency of sulphur in the liquid metal and its dependency on temperature and composition of the melt. Several papers in this field have recently been published.', ' Another subject is the study of the sulphur capacity of the slag. A molten slag is indeed complex, and even if sulphur distribution data for a large variety of molten slags may give empirical data about their desulphurizing power, the importance of the individual components is still not quite clear. It is accepted generally that lime is the most important desulphurizing component in the slag. The present investigation has as its purpose to study the desulphurizing power of lime in its standard state, and to provide a basis for thermodynamic calculations of the desulphurizing power of various lime-containing slags. The standard state of lime at steelmaking temperatures is solid calcium oxide, CaO. It can react with sulphur to form solid calcium sulphide, CaS. The relative stability of calcium oxide and calcium sulphide is expressed by the free energy of the reaction: 2Ca0 (s) + S1 (g) = 2CaS (s) + O2 (g) The existing free energy data for this reaction, listed by Kelley5 nd Osborn,' are uncertain to about 10 kcal and are of limited value for a calculation of equilibrium constants. Under the conditions prevailing in a melting furnace, the sulphur pressure may be expressed conveniently by the ratio H,S/H2 and the oxygen pressure by the ratio H,O/H, (or CO,/CO). The desulphurizing power of calcium oxide may, therefore, be studied by the reaction CaO + HIS = CaS + H2O. A study of this reaction may be complicated by certain side reactions: Water vapor and hydrogen sulphide may react. to form sulphur dioxide, and calcium sulphide may be oxidized to calcium sulphate. A thermodynamic calculation shows that these side reactions will be suppressed to insignificance if the equilibrium is studied in the presence of an excess of hydrogen. The apparatus used is shown in Fig. 1. About 10 g calcium oxide and 20 g calcium sulphide (laboratory qualities) were intimately mixed, and some water was added to make a thick paste. The paste was put into a thimble of zirconium silicate, which was placed within the constant temperature zone of a furnace, and capillary refractory tubes were attached in both ends. After the mixture had been heated in dry hydrogen at 1000°C for several hours all Ca(OH), and CaCO, had decomposed and CaSO, was reduced, so only CaO and CaS remained in the thimble forming a porous plug. The mixture was examined by X-ray diffraction after the initial reduction in dry hydrogen as well as after the subsequent experimental runs up to 1425 °C. It was shown that crystalline calcium oxide and calcium sulphide were always present together in about equal amounts. The unit cell edges were found to be 4.80A for CaO and 5.68A for CaS in good agreement with existing literature values." This shows that the mutual solid solubility is very small, and that the compounds are present in their standard states. Purified hydrogen was passed through water sat-urators kept at constant temperature in a thermostat bath. The amount of water vapor saturation was checked by means of a dew point method, not shown on Fig. 1. The gas mixture was passed through the capillary inlet into the furnace, where it was sifted through the porous plug of calcium oxide and calcium sulphide. The hydrogen sulphide present in the outgoing gas was absorbed in a zinc acetate solution and the hydrogen was collected over water. When one liter of hydrogen had been collected, the amount of hydrogen sulphide was determined by iodometric titration. As one molecule of H,O is used for the formation of each molecule of H,S, the equilibrium ratio H,S/H,O would be , where (H,O) is the molar concentration in the ingoing gas, and (H,S) the molar concentration in the outgoing gas. In the present work (H,S) was always very small compared to (H20). In order for the observed H,S/H20 ratio to represent the true equilibrium ratio the gas flow has to be: 1—Sufficiently slow to give a complete establishment of equilibrium, and 2—sufficiently fast to counteract thermal diffusion. Incomplete reaction would give a value decreasing with increasing flow rate, and thermal diffusion would give a value increasing with decreasing flow rate. When inlet and outlet tubes of about 2 sq mm cross-section were used, the observed gas ratio was independent of the flow rate between 15 and 125 cc per min, Fig. 2. In this range, therefore, the observed gas ratio represents true equilibrium.* For the rest of the in-

Jan 1, 1952

By S. G. Cupschalk, F. A. Dahlman, J. J. Wert

Studies have been undertaken to determine the indicidual effects of particle size, degree of long-range ovder, antiphase domain size, and root mean square stran on the microhardness and yield strength of ordered alloys. Dnta have been analyzed for Cu3Pt initzally ordered to a value of 0.82 and after deformations of 1 and 6 pct. It was observed that deformation fleatly reduced the degree of long-range order. Furtherrnore, wztkin this range of relatively small deforntntlons, the average particle size changed very little while the antiphase domain size was greatly reduced. Smultaneosly, the mcrohardness changed by a factor of two durzng the deforrtation process. PREVIOUS studies have reported some of the effects of cold work on the broadening of X-ray diffraction peaks. These investigations were performed on powder and wire samples representing both ordered and disordered states; i.e., the specimens were initially studied in a severly cold-worked condition. By comparing the difference in line shape between the annealed and cold-worked peaks, fundamental information was obtained concerning particle size, strain distribution in different crystallographic directions, degree of long-range order, and change in antiphase domain size. Considerable theoretical work has been done concerning the analysis of diffraction data obtained from cold-worked metals. Stokes' expressed the change in diffraction profiles in terms of Fourier coefficients. Much of the work in this area has been summarized by warren2 in an extensive review article concerning the analysis of plastic deformation by X-ray diffraction. Cohen and Bever3 applied these techniques in studying the effects of cold work on alloy systems exhibiting long-range order. They utilized the Fourier coefficients of fundamental peaks in conjunction with those of the superlattice peaks to determine the change in antiphase domain size. Little work of this nature has been reported for ordered systems that have undergone small degrees of plastic deformation. The purpose of this investiga-tion was to determine the effects of small deformations in such a material with respect to particle size, strain distribution in various crystallographic directions, antiphase domain size, degree of long-range order, and hardness. EXPERIMENTAL PROCEDURE CusPt was used for the initial investigation since the order-disorder transformation takes place with- out a change in crystal structure. The transformation is readily detectable via X-ray diffraction techniques due to the large difference in the scattering factors of copper and platinum. Additionally, the alloy is relatively low melting (approximately 1300°C) and is easily deformable in both the ordered and disordered states. 1) Specimen Preparation and Cold Working. A 100-g, 12-in. diam., cylindrical specimen of Cu3Pt was prepared by melting and casting 99.99 pct pure Cu and Pt i.n vacuo. Prior to any mechanical working, the material was homogenized in a vacuum for 60 hr at 100O0C, and surface defects were removed by machining to a depth of approximately 116 of an in. The material was then cold-rolled, with an intermediate anneal, into a strip approximately 12 in. wide by 14 in. thick. Straightening and flattening removed another 0.025 in. from the thickness. After a recrystallization treatment at 750°C for 30 min, the specimen was slow-cooled from 55OoC, at the rate of 6°C per hr, down to 150°C to induce superlattice formation. This treatment yielded an ASTM grain size of 7 and a degree of long-range order equal to 0.83 0.06. After obtaining X-ray and Knoop hardness data, the sample was cold-rolled approximately 0.75 pct in one pass through a hand-operated jewelers' mill. X-ray and hardness data were again obtained and the specimen was reduced an additional 5.41 pct in a single pass through the mill. 2) X-Ray Measurements. The specimen was examined in the ordered condition and after the two degrees of cold working previously mentioned using a General Electric XRD-5 unit equipped with a spectrometer and scintillation counter. Using Mo-Ka radiation with a zirconium filter, six orders of the 100 reflection were obtained. It was anticipated that point counting would be necessary for an accurate determination of the low-intensity peaks and tails: however, it was demonstrated that, by using a scanning speed of 0.2 deg per min and the appropriate time constant, the recorded data were sufficiently accurate. Thus, for ease of experimental procedure, all peaks were recorded on chart paper. Specimen position in the holder was considered to be insignificant after making a series of measurements of the same peak area in different positions with respect to the beam. Since peak overlapping did occur at high values of 20, it was necessary to separate the peaks graphically prior to analyzing the data in order to minimize this source of error. The peak tails were also carefully drawn to obtain the best possible data. Fourier coefficients of the line profiles were calculated on an IBM 7072 computer, and graphical meth-ods2j3 were employed in analyzing the results. For this type of calculation, in which the line profile is represented by intensities taken at set intervals, the intervals selected must be sufficiently small to give an accurate representation of the line profile. It was decided that for 20 = 0.02 deg the line profiles were

Jan 1, 1967

By K. R. Carson, J. Weertman

An attempt is made to ascertain the mechanism of tangle and cell formation and its dependence upon dislocation-interstitial carbon interactions. The strain-hardening behavior of single crystals of 3.25 pct Si-Fe was determined at 300° and 425°K and under conditions of both continuous and interrupted tensile strain. Significantly enhanced hardening was observed in crystals deformed at the elevated temperature, and it was further accentuated by interrupted straining. Transmission electron microscopy was used to study the resultant dislocation structures. Strain aging was found to aid tangle and cell formation at 425°K, but at both temperatures embryo tangles formed solely from primary glide dislocations, presumably by a process involving cross slip and "mushrooming". IN the course of plastic deformation all bcc metals and alloys develop a dislocation structure characterized by loose-knit groups of tangled dislocations. With increasing strain the tangles become more tightly knit and grow larger; finally a three-dimensional cellular substructure is formed:1 This process has been observed with the transmission electron microscope.'-l7 However, most investigations were confined to the study of nearly pure polycrystalline metals at relatively low temperatures. At intermediate temperatures, 0.17 to 0.14 Tm where T, is the melting temperature in degrees absolute, the mobility of interstitial impurities such as carbon is high enough to permit migration to nearby glide dislocations but is still low enough so that a significant drag force is exerted.18,19 it is also in this temperature range that a hump occurs in the curve of work-hardening rate vs temperature for iron. Analogous plots for tantalum" and columbiumzo show a definite upward trend in the work-hardening rate. Keh and Weissman1 have pointed out that this behavior may be explained solely on the basis of changes in the dislocation configuration: at low temperatures the dislocations tend to be relatively straight and uniformly distributed, but at intermediate temperatures tightly knit tangles and cellular substructure appear. The interference of these tangles with glide dislocations causes the observed increase in the work-hardening rate. This explanation appears reasonable, yet one might ask what factors cause tangle formation to be so favorable at intermediate temperatures. It seens likely that the strong dislocation-interstitial interactions which are known to occur in this temperature range are at least partly responsible," with the magnitude of the effect being proportional to the interstitial concentration. The purpose of the present work is to study the relationship between tangle formation and strain hardening in a bcc metal in the temperature range 0.17 to 0.4 Tm. Particular emphasis was placed upon a study of the effects of interstitial-dislocation interactions. Single crystals of 3.25 pct Si-Fe containing about 200 ppm of C in solid solution were used in the investigation for the following reasons: 1) The mobility of interstitial carbon in 3.25 pct Si-Fe is negligible at 300°K but increases rapidly at slightly elevated temperature22. Hence, differences between the flow curves and dislocation structures of crystals deformed at 300°K, 0.17 T,, and crystals deformed, say, at 425°K, 0.24 Tm, should be appreciable because of the enhanced dislocation-carbon interactions at the elevated temperature. This effect was accentuated in some samples by interrupted straining, thereby introducing a certain amount of aging. 2) Near room temperature, slip in suitably oriented 3.25 pct Si-Fe single crystals is largely confined to the (110) planes.23'24 Dislocation structures formed under conditions of single glide are the least complicated and their method of formation is the most easily discernable. 3) Dislocations in Si-Fe can be tightly locked with carbon atmospheres by a low-temperature aging treatment. The subsequent thinning of samples to foil thicbess causes little or no rearrangement in the dislocation structure.25 EXPERIMENTAL PROCEDURE Large single-crystal sheets of 3.25 pct Si-Fe were donated by Dr. C. G. Dunn of the General Electric Research Laboratory, Schenectady, N. Y. The orientations of the sheets were determined and slabs 1.0 by 0.25 by 0.05 in. were cut such that the desired tensile axis corresponded to the long dimension. The slabs were mechanically polished and subsequently decar-burized by heating at 1000°C for 3 days in a flowing wet-hydrogen atmosphere. A carbon content of about 200 ppm was introduced by heating at 805°C for 25 min in a flowing atmosphere of dry hydrogen containing heptane vapor. Shaped copper tools were then used to spark-machine at 0.125 by 0.50 in. gage length onto each slab. Vacuum annealing at 1225°C for 2 days followed by a quench into the cold end of the furnace to retain carbon in solid solution concluded the soecimen preparation. Continuous tensile flow curves for crystals of severa1 orientations Were obtained both at 300' and 425°K. A strain rate of 6.67 x 10-4 Per set was used in these and all other tests. Crystals oriented for single glide, B and D in Fig. 1, were subjected to a 3.5 pct plastic elongation to insure uniform slip along the gage length; they were then immediately subjected to interrupted strain cycling as indicated in Fig. 2(a). Each cycle consisted of unloading to 1.5 kg per sq mm, holding

Jan 1, 1969

By W. W. Vaughn, R. H. Barnett, E. E. Wilson

Soon after the search for uranium ores on the Colorado Plateau began in earnest, thousands of feet of drill core ranging from 1 1/8 to 2 1/8 in. diam became available for study. Although significant advances had been made in the technique of quantitative gamma-ray borehole logging, instrumentation was in the development stage, and complete reliance could not be placed on gamma-ray logs alone to interpret quantitatively the meaning of radioactivity in a drillhole. A method that would be faster than chemical analysis and still give reproducible and reliable results for various drill core sizes was desirable to provide additional information on the enormous footage of drill core being accumulated. A solid phosphor scintillation drill core scanner was designed and constructed. Basically the instrument was developed to measure radiation from a drill core which would not be clearly recorded by a gamma-ray logger using a Geiger tube as the sensitive element. Such data would be beneficial in constructing isorad maps to delineate ore-bearing zones. A calibration in the range 0.01 to 0.1 pct eU.,O, was provided; above 0.1 pct eU3O8 gamma-ray logs were available and were being used to calculate grade and tonnage of ore reserves. The core scanner, however, has been used to estimate equivalent uranium content of ore-grade materials containing as much as 2.2 pct eU3O8 with an accuracy of ± 10 pct, the sample being in the form of a BX drill core. Actually, an apparent calibration of eU3O8 vs counts per unit time is a straight line with a slope that is a function of the sensitive element and the geometry of the counting assembly. A true calibration that will show the expected departure from a straight line is due principally to the random nature of the pulse from a radiation source and the nonlinearity of the electron circuitry. Design and Construction: Three methods of detecting radioactivity were considered and applied in developing the core scanner now in use: 1) the Geiger tube, 2) liquid scintillation phosphors, and 3) solid scintillation phosphors. The desired sensitivity and long-term drift characteristics needed for this operation could be attained only by using solid scintillation phosphors. All three methods are discussed. Before scintillation counters were common, nine beta-gamma sensitive Geiger tubes 7/8 in. diam by 12 in. long were used, arranged to surround the drill core with tube axes parallel to the axis of the core. This arrangement of Geiger tubes was en- closed in a lead shield 1 in. thick, and provision was made to slide a 6-ft length of drill core manually into the counting chamber, one foot at a time. A count for each segment was taken with a scaler while the core remained stationary. The equivalent uranium content of the different sections of drill core could then be estimated with the aid of a calibration curve of counts per unit time vs percent equivalent uranium (eU). In rare cases the effects of the radioactivity concentrated in small areas within the core introduced errors in the readings made with the Geiger tube arrangement owing to the geometry of the measurement. The variability of counting rate due to a localized concentration of radioactivity in a spot in the wall of a drill core is illustrated in Fig. 1. This effect and the inherent low efficiency of the Geiger tube were considered major disadvantages of this counting arrangement. When liquid scintillation phosphors became available the core scanner in Fig. 2 was constructed to make a more accurate measurement of the equivalent uranium content of a sample. This instrument contains about 4 liters of liquid phosphor in a stainless steel coaxial cylinder 1 ft long, with inner and outer walls 0.060 in. and 0.125 in. thick, respectively. Four end-window type photomulti-plier tube with cathodes of 2 in. diam, immersed in the solution at right angles to the axis of the core, were used to observe light flashes in the phosphor. The liquid phosphor offered equal sensitivity to radiation originating at any point in the enclosure and represented geometrically the optimum in design. However, providing a semi-permanent leak-proof seal between the glass envelope of the phototube and the metal walls of the container proved to be a serious problem in constructing the equipment. The most effective seals were especially machined O-rings from sections of large tygon tubing. The tygon took a permanent set owing to cold flow characteristics and in most cases sealed completely. The light absorption characteristics of the liquid phosphor changed gradually with time, and after one month the counting rate had decreased to half the original value. The most sensitive liquid phosphor tested proved to be a solution containing 4 g of 2.5-diphenyloxazole and 0.01 g of 2-(1-naphthy1)-5-phenyloxazole per liter of toluene. With fresh solution in the chamber and with all photomultiplier tubes operating in parallel, the counting rate contributed by any one of the four photomultiplier tubes was about 85 pct of the counting rate from a single tube operated individually. From these observations it was concluded that owing to coincident loss and light attenuation within the liquid phosphor, the apparent sensitivity could not have been materially increased by additional phototubes. However, this approach to core

Jan 1, 1960

By J. H. Swisher

The rate of internal sulfidation of austenitic Fe-Ti alloys in H2S-H2 gas mixtures is controlled primarily by sulfur diffusion, with counterdiffusion of titanium playing a minor role. At temperatures below 1100°C, enhanced diffusion along grain boundaries becomes important. The rate of internal sulfidation at 1300°C is approximately equal to the rate computed from the sulfur diffusion coefficient. The diffusion coefficient of titanium in y iron has been determined from electron microprobe traces in the base alloy near the subscale interface. The solubility of sulfur in Fe-Ti alloys has been measured in the temperature range from 1150° to 1300°C. The equilibrium sulfur content is found to increase with titanium content, due to the large effect of titanium on the activity coefficient of sulfur. The Ti-S interaction becomes stronger as the temperature decreases. TITANIUM as an alloying element in stainless steels is an effective scavenger for interstitial impurities, carbon in particular. Titanium is known to form stable sulfides; however extensive thermodynamic data on the Ti-S system are not available. Schindlerova and Buzek1 have shown that the Ti-S interaction in liquid iron is moderately strong. There have been no previous studies of the Ti-S interaction in solid iron. Internal sulfidation of Fe-Mn alloys was the subject of a recent investigation by Herrnstein.2 He found the rate of internal sulfidation to be an order of magnitude greater than predicted from available solubility and diffusivity data. A satisfactory explanation for the discrepancy could not be given. In the present study, the solubility of sulfur in austenitic Fe-Ti alloys was measured using a standard gas equilibration technique. Fe-Ti alloy specimens were also internally sulfidized. The rate of internal sulfidation was measured as a function of temperature and alloy composition. Supplementary electron micro-probe measurements were made to provide additional information on the nature of the internal sulfidation process. EXPERIMENTAL The starting materials were alloys containing 0.12, 0.24, 0.38, and 0.54 wt pct Ti. The alloys were made in an induction furnace by adding titanium to electrolytic iron that previously had been vacuum-carbon-deoxidized. The major impurity in the alloys as determined by chemical analysis was carbon. The carbon content of the alloys averaged about 100 ppm; metallic impurities were presented in concentrations of 50 ppm or less. Specimens were made in the form of flat plates, 0.03 by 2 by 4 cm for the equilibrium measurements and 0.5 by 1.5 by 3 cm for the rate measurements. The experiments were performed in a vertical resistance furnace wound with molybdenum wire and containing a recrystallized alumina reaction tube. In the gas train, flow rates of the reacting gases were measured using capillary flow meters. The source of H2S was a mixture of approximately 2 pct H2S in H2, which was obtained in cylinders from the Matheson Co. A chemical analysis was provided with each cylinder. The H2-H2S mixture was diluted with additional hydrogen to obtain the desired ratio of H2S to H2, and the resulting mixture was diluted with 30 pct Ar to minimize thermal segregation of H2S in the furnace. Argon was purified by passage over copper chips at 350°C and subsequently over anhydrone. Hydrogen was purified by passage over platinized asbestos at 450°C and then over anhydrone. The H2-H2S mixture was purified by passage over platinized asbestos and then over Pas. The samples used in the solubility measurements were analyzed for sulfur by combustion and iodometric titration. The subscale thickness in the internally sulfidized samples was measured on a polished cross section, using a microscope with a micrometer stage. Electron microprobe traces for titanium in solution were made on several samples that had been internally sulfidized. A Cambridge microanalyzer was used, and the known titanium content at the center of the specimen was used as a calibration standard. The procedure for the microprobe measurements will be described further when the results are presented. RESULTS AND DISCUSSION Equilibrium Data. Fig. 1 shows the sulfur concentration as a function of gas composition for three alloys equilibrated at 1300°C. The dashed line is based on data published by Turkdogan, Ignatowicz, and pearson3 for pure iron. The breaks in the curves are the saturation points for the alloys. The fact that the initial slope decreases with increasing titanium content indicates that titanium interacts strongly with sulfur in solution. To obtain information on the composition of the precipitating sulfide phase, the measurements described in Fig. 1 were extended to higher sulfur partial pressures. These results are shown in Fig. 2. (The initial portions of the curves are reproduced from Fig. 1.) The highest PH2s /pH2 ratio used is believed to be below the ratio required for the formation of a liquid sulfide phase. Time series experiments were used to study the approach to equilibrium in the samples. It was found that equilibrium with the gas phase was reached in less than 4 hr at 1300°C.

Jan 1, 1969

By R. R. Jones, R. W. Kraft

Zn-Mg2Znn eutectic alloys nzay freeze willr either rodlike or lanzellar rnorphology. Alloys with slighlly more than /he eutectic arrzount of rnagnesillrn usually contain three-cnned dendrjles of MgzZnll in a eutec-lic ttlulris. All three morphologies haue the same cryslallographic orientution relationship: (0UOl) zn - 11 (111) Mg2Znll and (2310)Zn 11(101) Mg2Znll, but u3ith different prej-erred groulth direclions. The lurnellae lo rods transifion in con/rolled ingols qf euleclic cotnposition occurs because lhe large kinelic undercooling due to MgzZnll minirrzizes /he ejj-ecl of the solid-solid inlerface energy. The eutectic morphology is influenced by the presence of lhree-nned dendrites 0-f MgzZn11 which may conlrol /he rricroslrccture by acting as nuclealion sites. In recent years there has been much interest in eutectic solidification and several theories have been proposed. One of the confusing factors is the existence of various morphologies in which the solidified phases may form. The lamellar microstructure seems to be most common in metal eutectics, and it has been claimed' that all regular eutectics should be lamellar if sufficiently pure. However, there still remain eutectic alloys which are not lamellar or which change their morphology as a function of growth conditions. The eutectic between zinc and the intermetallic phase Mg2Znll was chosen for this investigation because it has been found to solidify in more than one morphology. The diagram in anssen' locates the eutectic point at 3.0 wt pct Mg and 367°C. lliott gives 364°C as the eutectic temperature, leaving the phase compositions unaltered. Since the growth conditions determine the micro-structure of the solidified alloy, the factors controlling the transition from one morphology to another could be studied. The lamellae to rods transition is of particular interest. PROCEDURE Alloys were prepared from carefully weighed portions of 99.999 pct Zn and 99.97 pct Mg by melting in Pyrex containers under argon and casting into graphite boats. The resulting ingots were remelted under argon and solidified unidirectionally in a horizontal tube furnace at growth rates ranging from 2.0 to more than 50 cm per hr under a temperature gradient, measured over a 5-cm length, of 9" to 14°C per cm. The solid-liquid interface appeared to be planar at all growth rates although no attempt was made to confirm this by decantation or quenching. A few ingots were allowed to freeze uncontrolled. Most alloys were of the nominal eutectic composition, 3.0 wt pct Mg according to Hansen2 and lliott, but some contained as much as 3.35 wt pct Mg. Chemical analyses were not run since metallographic examination confirmed that the desired composition was achieved. Specimens were cut from the middle portion of the ingot normal to the growth axis, polished mechanically, and etched with 2 pct Nital. Suitable areas were selected for the determination of crystallographic orientation relationships by a tiontechniqueof described previously by one of the authors.4 The (2310) planes of zinc and the (8701, {944}? (1032) planes of Mg2Znll were found suitable for orientation determination; experimental error was on the order of 2 or 3 deg. RESULTS Three different morphologies were found in the unidirectionally solidified alloys: lamellar eutectic, rod-like eutectic, and a structure whose most predominant characteristic was the presence of three-vaned (cellular) dendrites of Mg2Znll. These dendrites were only found in alloys with more than the eutectic amount of magnesium. In some ingots fine hexagonal needles of Mg2Znll surrounding a core of MgZn2 were observed. They were probably due to incomplete alloying and seemed to have no effect on the eutectic morphology. In addition hexagonal spirals like those discussed by Fullman and wood5 and Hunt and acksonh ere observed in some ingots frozen without directional control. Both MgZZn,, and MgZnz were detected by X-ray diffraction in these alloys. Since the morphology could not be grown unidirectionally and no characteristic orientation relationship between the phases was found, further study was limited to the lamellar: rodlike, and three-vaned dendrite morphologies. Alloys of Eutectic Composition, No Dendrites. The mcrostructures of allovs with no three-vaned dendrites were either lamellar or rodlike depending on the growth rate. At rates below 10 cm per hr the morphology was lamellar, consisting of two sets of parallel plates intersecting at about 54 deg like the Mg-MgzSn eutectic described by raft.7 At growth rates faster than 14 cm per hr the microstructure showed rods of zinc in a matrix of MgnZnll, while intermediate rates yielded mixtures of rods and lamellae in small groups. The lamellar "grains" were often several millimeters in cross section, but contained small irregular areas which divided each grain into perfect islands 100 or 200 p in diam. Lamellae were parallel to each other throughout the grain in spite of these defects in the structure, Fig. 1. Rods, on the other hand, could only be produced in small groups arranged like fish scales and separated by irregular areas of appreciable thickness, Fig. 2. Alloys Not of Eutectic Composition, With Dendrites. In alloys with 3.1 to 3.35 wt pct ME,-. three-vaned dendrites bf MgzZnll were usually found surrounded by eutectic. At growth rates slower than about 10 cm per hr the dendrites were separated from each other by small areas of both lamellar and rod eutectic, Fig. 3.

Jan 1, 1969

By R. C. Buehl, J. P. Morris

F. D. Richardson—The authors are to be congratulated on this further contribution to our knowledge of the thermodynamics of the interaction between sulphur and carbon and silicon in liquid iron. As the authors state, the influence of carbon and silicon on the activity coefficient of sulphur in liquid iron is clearly of great importance in the blast furnace, since it must cause a three to fourfold improvement in the partition of sulphur between slag and metal. The influence of increasing temperature in further increasing the activity coefficient of the sulphur in the metal in the blast furnace by increasing the carbon content is also of interest. This effect, however, is probably only part of the reason for the general observation in blast furnace practice, that the sulphur content of the metal is lowered by increasing temperature. Other contributing factors are the lowering of the oxygen potential in the presence of carbon by increasing temperature and the probable increase in the activity coefficient of the lime in the slag for the same reason. The former of these effects, which works via the (CaO) + [S] = (CaS) + [O] equilibrium, might possibly account for a 70 pct improvement in the sulphur partition and the latter might give a further 50 pct improvement. C. Sherman—I would like to compliment the authors on their very careful research. If I may, I would like to show results of calculations on the carbon-sulphur-iron system similar to the ones that were shown in our paper for the silicon-sulphur-iron system. For Fe-S-C ternary system k=PHgs/PH2 x 1/(f1°) (f2°) (%S) where fs = sulphur activity coefficient fs' = fs for Fe-S system of equal pct S f3° = f2/f2 for Fe-S-C ternary system This same analysis has been used on other systems, but the results shown in fie.- 7 are for carbon and silicon. L. S. Darken—I would like to make two brief comments in addition to complimenting the authors on an apparently very precise and accurate investigation. The first is that the present work is in agreement with a calculation by Larsen and myself." Our calculation (much less precise than the present work) was based on: (1) Unpublished work on the sulphur content of molten iron (1.5 pct at 1500°C) in equilibrium with graphite and an iron sulphide slag; (2) the distribution coefficient of sulphur between slag and carbon-free liquid iron. We expressed the result in a form equivalent to log 7. = 0.18 [%C] which gives an activity coefficient (?s.) of sulphur only slightly higher than the authors find and certainly within the precision of the earlier work. My second comment concerns the correlation of the thermodynamic findings with atomistics. A rough pic- ture of the atomic arrangement in the liquid solution is rather easily conceived for this particular liquid solution containing iron, carbon, and sulphur. Carbon has a very much stronger affinity for iron than for sulphur. Hence we may conclude that a sulphur atom will but seldom be adjacent to a carbon atom—since this would be a position of high energy. From the metallic radii of iron and carbon we know that six iron atoms pack neatly around one carbon atom. Thus each carbon atom in retaining this shell of iron atoms (which latter may not be replaced by sulphur on account of the high energy requirement) decreases the available positions for each sulphur atom by six. Hence each atomic percent of carbon decreases the equilibrium sulphur content by 6 pct (of itself). Or, at low concentration each atomic percent of carbon increases the activity coefficient of sulphur by 6 pct. This is in good agreement with the observed increase (6 or 7 pct at low carbon content). It is indeed gratifying to find a case where, by such simple reasoning, quantitative agreement is found between precise data and the modern picture of the atomistics of the metallic state. J. P. Morris (authors' reply)—We would like to point out that there is an error in the equation on p. 322 of the paper. The third equation should read: ½S2 (gas) + H2 (gas) = H2S (gas) The authors wish to thank everyone for the interest they have shown in the paper. In regard to the general observation in blast furnace practice, that the sulphur content of the metal is lowered by increasing the temperature, Dr. Richardson is correct in stating that the cause can be attributed only in part to the increase in activity coefficient of sulphur resulting from the rise in carbon plus silicon content of the metal with rise in temperature. However, this factor is probably an important one. The results of one experiment, performed since this report was written, indicate that at a constant temperature the addition of silicon to a melt saturated with carbon causes an increase in the activity coefficient of sulphur even though the carbon solubility is lowered. In this test, 2.5 pct silicon was added to a melt saturated with carbon and maintained at 1400°C. Although the carbon content dropped from 4.85 to 4.1 pct, the activity coefficient of sulphur was increased by about 20 pct.

Jan 1, 1951

By Sven Fornander

L. F. Reinartz (Armco Steel Corp., Middletown, Ohio) —I would like to know, in the practical application of the Kalling process, what kind of a lining was used, how thick was the lining, and how much metal was treated at one time? S. Fornander (author's reply)—The rotary furnace is lined with a course of fireclay bricks 6 in. thick. This course is backed by 5 in. of insulation. The furnace has a capacity of about 15 tons. Mr. Reinartz—How was the ladle preheated? Mr. Fornander—As pointed out in the paper, the furnace was heated by a gas flame in the beginning of the experiments. During these first tests, however, the desulphurization was inconsistent. We think that this was due to the fact that iron droplets sticking to the furnace walls were oxidized by the gas flame. Now, the furnace is operated without preheating of any kind, and the results are much better. T. L. Joseph (University of Minnesota, Minneapolis, Minn.)—I might add one comment. This furnace was heated with a flame and for a time they had a little difficulty due to some residual metal in the rotating drum that would oxidize in between treatments and they found therefore, that it was very essential to drain the drum completely of metal so that they would not build up any ferrous oxide between treatments and they eliminated some of their erratic heats by maintaining those more reducing conditions. It was interesting to watch this operation. As soon as the drum started to rotate there was considerable flame, at least, at the time I saw it, that came out around the flanges, indicating there was quite a little pressure on the inside of the drum. W. 0. Philbrook (Carnegie Institute of Technology, Pittsburgh)—Is the reaction slag in the Kalling process liquid or solid, and how is it separated from the metal? Mr. Fornander—In the process there is no slag in the usual sense of the word. The lime powder does not melt during the treatment. After the treatment the lime is still in the form of a fine powder. It is separated from the metal by means of a piece of wood of suitable size placed within the furnace before it is emptied. D. C. Hilty (Union Carbide & Carbon Research Laboratories, Niagara Falls, N. Y.)—Dr. Chipman has given us some of his ideas in connection with a specific effect of silicon and silica on sulphur elimination and how silicon might interfere with desulphuriz- ing in the blast furnace. I wonder if he would like to elaborate on the possibility of a similar effect of silicon in the Kalling process? J. Chipman (Massachusetts Institute of Technology, Cambridge, Mass.)—Silicon does not interfere with the Kalling process. Anything that has strong reducing action is good for desulphurization. In these tests where the temperature was low compared to blast furnace temperatures, the silicon that is in the metal is a better reducing agent than the carbon. At high temperatures, carbon is the better. It is not the silicon in the metal that interferes with desulphurization, it is the silica in the slag. Mr. Joseph—I might add that the metal that was tapped from the drum after desulphurization was really at quite a low temperature. It was not measured, but I think it was well under 1300 °C, probably 1200" or a little above that. That was one of the difficulties, and I think there is no question about the fact that the Kalling process—in that it affects desulphurization between powdered lime, solid and liquid iron— is a reaction definitely between the solid lime and the liquid iron. E. Spire (Canadian Liquid Air, Montreal, Canada) — This Kalling process seems very interesting to us and after all it is only a mixing action that is taking place between the iron and the slag. We have attempted to do the same thing in another way. We have placed at the bottom of the ladle a porous plug through which we injected an inert gas. It can be nitrogen or argon. This plug is placed at the bottom of the conventional ladle and gas injected through the plug. That has appeared in our patent. To define this new type of treatment, I use the word gasometallurgy. I do not know if you like it, but it is a way of defining methods of treating metal using gases. What we do is exactly what is done in the exchange process in another way. We have a porous plug at the bottom with a high lime slag on top of the metal. Using this method, we have very good agitation of metal and slag, and with a small flow of gas, we can achieve a very strong agitation. For instance, in the 500 lb ladle, we use only 5 liters of gas a minute. We have an agitation compared to very rapidly boiling water in a pail. Moreover, the agitation can be controlled to create any amount of mixing desired. In a few minutes, with this method, the sulphur dropped from 0.58 to 0.11. These results have been improved since, and we have obtained results like 0.08

Jan 1, 1952

By R. Bainbridge

THE Consolidated Mining and Smelting CO. of Canada Ltd. has operated a lead smelter at Trail, B. C., for many years. In order to take advantage of metallurgical advances, as well as to improve materials handling methods, this company, commonly known as "Cominco," commenced planning a program of smelter revision and modernization some years ago. The first stage of this program involved the design and construction of a new blast furnace gas cleaning system. The selection of equipment, the design of facilities, and preliminary operating details of this system will be dealt with in this paper. The essential problem was to clean and collect 100 tons of dust daily from 153,000 cfm* (12,225 lb per min) of lead blast furnace gas which varied in temperature from 350º to 1100°F. Because it was desired to collect the dust dry, either a Cottrell or a baghouse cleaning plant was to be selected. Comin-co's many years of experience with both systems provided a background for choosing the most satisfactory installation. All information pertinent to the two methods of dust recovery was carefully investigated, and it was decided to replace the existing equipment with a baghouse. Very briefly, the reasons for this decision were as follows: 1—A baghouse installation would be practical because the SO2 content of the gas was low and corrosion would not be a problem if the baghouse operating temperatures were held sufficiently above the dew point. 2—Variations in the physical characteristics of fume and dust, which are inherent in this blast furnace operation, should not substantially affect the operating efficiency of a baghouse. 3—For the same capital cost, metal losses (stack and water losses) would be appreciably less in a baghouse. 4—A baghouse would be easier to operate, and would not require the use of highly skilled labor. 5—Operating and maintenance costs of a bag-house would be lower. 6—The only available space for reconstruction was relatively small, and not suited to a Cottrell installation. Once the baghouse system was decided upon, detailed design of the installation was begun. Baghouse Design Gas Cooling: Before the required capacity of the baghouse could be determined, the method of cooling the gas to the temperature necessary for bag-house operation had to be chosen. The problem confronting the design engineers was how best to cool 153,000 cfm of gas from a temperature ranging from 350°F to brief peaks of 1100°F, down to 210°F, the maximum safe baghouse inlet temperature. A survey of existing blast furnace gas temperatures in the outlet flue showed that the normal range was as given in Table I. The obvious choices of cooling method were: 1— cool completely by the addition of tempering air; 2—utilize a heat exchanger; 3—cool by radiation; and 4—cool with water spray in conjunction with the admission of tempering air. The advantages and disadvantages of the various cooling methods were: Air Addition: To cool completely by the admission of tempering air involved tremendous volumes, Fig. 1. For example, to cool 1 lb of blast furnace gas at 450°F requires 1.84 lb of air at 80°F or 1.60 lb at 60°F. As it is necessary to design for peak conditions, it can readily be seen that volumes of tempering air in the order of 1,500,000 cfm would have to be handled. Using the normal design figure of 2.5 cu ft per sq ft of bag area, a baghouse installation comprising some 600,000 sq ft of filter cloth would be necessary. Such design requirements would be prohibitive, not only from a standpoint of capital expenditure, but also because of space limitations. Heat Exchanger: The utilization of a heat exchanger was given serious consideration. A horizontal tube unit using air as the medium to cool the required volume of blast furnace gas from 400" to 250°F was investigated. Cooling above 400°F would be done by water spray, and below 250°F by admission of tempering air. The estimated capital cost of such a unit was found to be prohibitive. From an operating standpoint, there was considerable doubt as to whether the soot blowing equipment provided would effectively keep the dust from building up on the tube surface. The performance of heat exchangers operating on dusty gas in other company operations had not been too favorable. Radiation Cooling: Although somewhat cumbersome, gas cooling by radiation from 'trombone' tubes or other similar equipment (cyclones) is employed in many metallurgical operations. Such an installation was also considered. However, calculations showed that an installation much larger than the space available would be required to handle the gas volume involved. For example, to cool 153,000 cfm of blast furnace gas from, say, 600' to 250°F (i.e., remove in the order of 58,500,000 Btu per hr with heat transfer rates varying from 1.1 Btu per sq ft per hr per OF for the higher temperature ranges to 0.88 Btu per sq ft per hr per OF for the lower ranges) would need a cooling area of some 175,000 sq ft.

Jan 1, 1953

By PAUL AUDIBERT

THE downward trend of metal prices seems to act something like a reagent that precipitates selfishness in most business men's hearts; in the same way the upward trend precipitates altruism. Operators do not say: "I continue working my mine without curtailment because I can afford the present prices while my neighbor cannot and thus will give me way." They say: "I continue working my mine because 1 consider it a service for consumers. They will get better prices on a free market. Any curtail¬ment would be an attempt to raise prices artificially, that is to say, bad economics." Copper producers have learned how fallacious is such an opinion. It perhaps was not fallacious in olden times when all mines were worked on a small scale even if operating on a big deposit. At that time the supply and demand law played fair. When prices were too low and profits were vanishing for many mines, they would. shut at once, then little by little the metal scarcity would induce prices to climb up again; finally these prices would reach the point when it became profitable to again work the closed mines. Re-opening and shutting down mines was then a normal custom. Workers and operators weakly deplored this unstability, but there was nothing effective they could, do about it.

Jan 1, 1931

By F. G. BREYER

ZINC metallurgists continue to follow with keen interest reports of successful results from the continuous retort plants at Palmerton, Pa., and Meadowbrook. W. Va. The new process had already demonstrated its ability to reduce the lead content of the metal and to effect savings in labor and fuel. The past year has shown that the retorts themselves can he built to last long enough so that the retort cost per ton of metal is lower than in past practice with small horizontal re- torts. A number of the latest type of retorts are reported to have been in continuous operation for over 18 months without a failure and are said to be still in good condition. This seems a complete answer to the last question which remained before continuous zinc smelting could he pronounced commercially practicable.

Jan 1, 1932

By AIME AIME

METALLIC corrosion, which results from the chemical affinity of different metals for non- metallic elements, should be considered from both the kinetic and static viewpoints. From the stand- point of statics the metals may be arranged in the order of their affinity for oxygen, chlorine or any other acid- forming element. They may also be arranged in the order of the magnitude of the heat of combination of their corresponding compounds. It is misleading to talk about some metals as "noble," because these relations are all relative; and the order varies for oxygen, chlorine or whatever the element may be. A third way to arrange the metals is by their solution pressures. These three methods indicate that one metal is better under one set of conditions, but not so good under other conditions. Considering the problem from the standpoint of kinetics, the thing which saves most metals is what may be called the sluggishness of reactions. Using organic substances as an example, we know they will all oxidize, but fortunately they do so rather slowly. An increase in temperature always means an increase in the reaction rate, and preventing corrosion at high temperatures is a greater problem than at low temperatures. A second consideration is to avoid all sub- stances which might provoke catalytic action, and as it is generally impossible to foresee what substances might do this, the thing is to avoid all substances other than the main constituents of the alloys. A third consideration is to avoid local electrolytic action, which again indicates the advisability of avoiding the contact of differing substances.

Jan 1, 1926

By R. P. Alger, M. P. Tixier, D. R. Tanguy

In the combination induction-electrical log used at present in the field, the induction logging tool is appropriate for the investigation of moderately invaded formations. A new induction sonde with a radius and investigation about twice as large has been developed recently for the case of deep invasion. It has very nearly the same vertical resolution as the present sonde so that thin beds are defined as accurately as before. The characteristics of the new tool are described, the corresponding interpretation charts are given and field examples are discussed. The design of the sonic logging tool has been modified to improve the calibration and the reliability. The fact that porosity can be accurately recorded by means of the sonic log has prompted new interpretation procedures for saturation estimation, wherein the data concerning the various permeable beds in a given well are correlated. One approach consists of plotting transit time vs true resistivity, with an appropriate scale. With this approach, saturations can be estimated conveniently even in cases where formation water resistivity is not well known. In another approach, a comparison is made of the values of the formation waters computed from the re- sistivity and sonic logs. Using the concept of continuity, this procedure makes possible a quick determination of zones of saturation in shaly sands and/or in case of appreciable variations of formation salinities with depth. It has been found that the comparison of porosity from the sonic log with the apparent porosity computed from a short-investigation resistivity log may reveal, in many cases, the presence of residual oil and thus detect potentially productive formations; this procedure is valuable when the true formation resistivity and the resistivity of the formation water are in doubt. INTRODUCTION During the past year, the efficiency of log interpretation has been vastly improved. The improvements have largely resulted from the introduction of a deep-investigation induction device and from the application of new interpretation techniques that utilize sonic vs resistivity readings. Since the new interpretation techniques depend, in part, upon good values of true formation resistivity, the new induction log will be discussed under Part I. The sonic interpretation techniques will be studied under Part 11. Early in 1959, the 6FF40 induction equipment was introduced in the field. This device was designed for a better approach to true formation resistivities in deeply invaded zones. The greatly improved radial investigation of the 6FF40 equipment has been achieved without sacrificing vertical resolution. The first combination induction-electrical log, the 5FF40, was introduced as a standard tool in 1956 for the logging of wells drilled with fresh muds. The tool has received wide industry acceptance in the United States. 'The 5FF40 induction log has a radial investigation sufficient to overcome average depths of mud filtrate invasion. At 5d invasion, for example, the 5FF40 induction log will read about 1.4 R. in a water sand where R., = 10R,. At 10d invasion, such induction log would read 2.45 R. in the same water sand. In either case, the effects of invasion would not be sufficiently great to cause a water sand to be mistaken for a shale-free oil- or gas-producing zone. Some formations, however, invade deeply — in excess of 10d. Such water zones could be mistaken for oil-or gas-saturated sands unless the porosity balance' can clearly make the distinction. It is for these deeply invaded formations that the 6FF40 was developed. CHARACTERISTICS OF THE 6FF40 Radial Investigation Characteristics To describe the comparative responses of the 5FF40

By P. F. Gnirk, J. B. Cheatham

Single bit-tooth penetration experiments under static load were conducted on six rocks at confining pressures of O to 5,000 psi using sharp wedge-shaped teeth with included angles ranging from 30 to 120° In general, the force-displacement curves for all rocks exhibit an increasingly nonlinear and discontinuous behavior with decreasing confining pressure. The confining pressure at which a rock exhibits a macroscopic transition from predominantly ductile to predominantly brittle behavior during penetration vmies from about 500 to 1,000 psi for the limestones to greater than 5,000 psi for dolomite. The correlation between calculated values of force per unit penetration based on plasticity theory and experimental values is quite encouraging, even at confining pressures as low as 1,000 psi A qualitative correlation between volume of fragmented rock per unit energy input for a single bit-tooth and drilling rate for microbits appears to exist over a confining pressure range of 0 to 5,000 psi. INTRODUCTION Laboratory experiments utilizing a small-scale drilling apparatus have demonstrated that penetration rates are reduced considerably as a result of increasing the confining pressure from atmospheric to a few thousand psi?-3 This undesirable situation can, in general, be attributed to a combination of decreased efficiency of chip removal at the bottom of the borehole, increased rock-failure strength, and a possible change in the mechanism of chip generation and rock fragmentation with increasing confining pressure. To more fully understand the principles underlying the last circumstance, it is the purpose of this investigation to experimentally study the mechanism of single bit-tooth penetration into dry rock at low confining pressures and, in particular, to establish the confining pressure at which the penetration mechanism may undergo a brittle to ductile transition for various rock types commonly encountered in drilling. Confining pressure as used here refers to the differential pressure between the borehole fluid pressure and the formation-pore fluid pressure. EXPERIMENTAL PROCEDURE Using an experimental apparatus previously described: a single, sharp wedge-shaped tool was forced under a "statically" applied load into an effectively semi-in finite dry rock sample subjected to a prescribed confining pressure. To prevent the invasion of the con fining-pressure fluid into the pores of the rock sample during penetration, the exposed surface of the rock was jacketed with a layer of silicon putty.* Electrical instrumentation incorporated into the apparatus yielded a graphical plot of force on the tool as a function of penetration or displacement of the tool into the rock during an experiment. During the course of the experimentation the following conditions were maintained constant: (1) pore pressure — atmospheric (i.e., the rock was dry); (2) temperature — 75F; (3) rate of loading — essentially static (approximately 0.002 in./sec); (4) bit tooth — a sharp wedge-shaped tool loaded normal to the rock surface; (5) rock surface — smooth and flat; (6) drilling fluid — hydraulic oil; and (7) maximum depth of penetration — approximateiy 0.1 in. In addition, each experiment was performed on a different rock sample so the rock surface is free of a layer of cuttings and of any previous indentation craters. The influence of the corners of a borehole was neglected, since each rock sample was cemented inta a section of aLuminum tubing to simulate a semi-infinite body. Indentation experiments were made on a sandstone, two limestones, a marble, a dolomite and a schist with a 60' bit-tooth over a range of confining pressures from atmospheric to 5,000 psi and at a constant confining pressure of 1,000 psi for a variety of bit-teeth with angles ranging from 30 to

Jan 1, 1966

By G. R. Speich, D. J. Mack

The transformation of was studied by isothermal methods. At all temperatures, the ß transforms quickly to fine grained ß" which develops silver-rich striations. At higher temperatures the striations disappear, the final structure being Widmanstatten a in ß'. At lower temperatures, the striated ß" is consumed by pearlite nodules of a + ß' which in turn form the Widmanstatten a in ß'. The transformation is unlike any previously reported for a eutectoid. AS part of a general program of study on the eutectoid reaction, the Ag-Cd eutectoid seemed particularly attractive because the ß phase which undergoes the reaction is another of the typical "electron compounds," AgCd, having the disordered body-centered cubic structure. This work deals with the eutectoid which occurs at 50.5 wt pct Ag, and not with the eutectoid which has been reported at 42.9 pct Ag.' There has been no work reported specifically on the Ag-Cd eutectoid which was studied. The only information available on this eutectoid is that work done in the determination of the phase diagram. For the most part, the terminal portions of the diagram, Fig. 1, are well established; however, the region from 40 to 60 pct Cd is still somewhat in doubt. The X-ray evidence'..' in this region is in conflict with the results of thermal analysis. X-ray analysis3 does not show a polymorphic transformation at about 240°C as indicated by thermal analysis.1,5 Even the X-ray evidence, within itself, is confusing in this region of the diagram. One investigator2 reports the ß phase as hexagonal close-packed and has found three phases in the one-phase ß field. Other X-ray investigators"' ' report the ß phase as body-centered cubic. The authors believe that much of the confusion resulted from incomplete homogenization of the alloys. It was found that it takes almost a week to produce a completely homogeneous two-phase structure at 370°C. Another possible source of error is the volatilization of cadmium from the surface and cracks of the specimen during annealing with the possible formation of a cadmium-poor phase. The only metallographic data available on this eutectoid alloy are those of Durrant1 and Fraenkel and Wolff.5 his information is very limited. The phase diagram chosen as most nearly correct is that in the Metals Handbook," which is almost identical with the phase diagram suggested by Durrant.1 Materials and Procedure The alloy was prepared from high purity silver (999.5 plus fine) and high purity cadmium (0.02 Pb, 0.005 Cu, trace Fe). The metals were melted in carbon under a nitrogen atmosphere. Some cadmium was lost by volatilization but by remelting several times and adding cadmium a 200 g ingot of approximately eutectoid composition was obtained. Because the CdO fumes are quite toxic, the exit gas from the furnace was led through a water trap to condense the CdO. The ingot was homogenized at 650°C for 24 hr, no attempt was made to prevent decadmiumization. This resulted in a narrow decadmiumized rim on the ingot which was removed. Filings for analysis were obtained from cross sections of the ingot. Silver was determined by the Volhard thiocyanate method,? cadmium by difference. The ingot analyzed 50.8 pct Ag at the top of the ingot and 50.7 pct at the middle and at the bottom of the ingot. This alloy was only slightly hypo-eutectoid compared to the accepted value of 50.5 pct Ag and was used for the isothermal studies. It was felt that further attempts to adjust the composition might be futile because of the volatility of the cadmium. The ingot was sectioned to specimens approximately 3/4 x 3/4 x 1/8 in. These specimens were "austeni-tized" for 3 hr in air at 650°C to yield homogeneous ß and then quenched into a salt bath held at constant subeutectoid temperatures where the transformation

Jan 1, 1954

By J. T. McGrath, R. C. A. Thurston

Poly crystalline specimens of copper, and copper with various additions of zinc, were tested in plane-bending fatigue. In tests performed at a constant stress, the fatigue life of copper increased slightly with the addition of up to fifteen per cent zinc. There was a sharp increase in fatigue life as further additions of zinc of up to 35 pct were made. Surface observations showed that the coarse slip bands that are characteristic of fatigue became narrower as the concentration of zinc increased. The rate of crack propagation also decreased with increasing zinc content. It is suggested that the fatigue behavior is controlled by cross slip processes. CROSS slip occurs when dislocations move from one primary slip plane to another along a connecting or cross slip plane. The primary and cross slip planes have a common slip direction. Wide stacking faults, having low energy, tend to make cross slip more difficult. According to recent theories, the ability of a material to cross slip is important to the formation and propagation of fatigue cracks. McEvily and Machlin2 have examined this concept experimentally by fatiguing single crys tals of lithium fluoride, sodium chloride, silver chloride, and KRS-5.* The latter two materials ex- hibit easy cross slip, while the former two do not. Normal rapid fatigue failure was obtained in the silver chloride and KRS-5 specimens, whereas it was much more difficult to achieve in sodium chloride and lithium fluoride crystals. Furthermore, numerous slip band extrusions and intrusions were found in silver chloride and KRS-5 crystals, whereas only one isolated extrusion was found in lithium fluoride and none was found in sodium chloride. McEvily and Machlin also consider that there is no clear-cut demarcation between crack initiation and propagation. The intrusion-extrusion process, which initiates crack formation, continues ahead of the cracks to allow propagation of these cracks along slip bands. The process of cross slip is therefore important not only to crack initiation, but also to crack propagation. Holden4 postulates that fatigue crack growth is dependent upon cross slip and thus upon stacking-fault energy. Using an X-ray microbeam technique, he found a cyclic subgrain structure in fatigued aluminum and a-iron. Holden suggests that micro-cracks develop in the densely-packed sub-boundaries ahead of the main crack front, and when sufficiently numerous join up to promote crack growth. The ease with which a subgrain structure could form would depend upon the ability of the dislocations to cross slip, and thus upon the stacking-fault energy. Hence in aluminum, which has a high stack-ing-fault energy and in which cross slip is easy, Holden found that the relative rate of fatigue crack propagation was considerably greater than in stainless steel, a low stacking-fault energy material. It has been shown by a number of investigators that the probability of forming deformation stacking faults in copper increases as zinc is added.5-7 Cross slip should, therefore, become more difficult as the zinc concentration is increased. In order to check the validity of the foregoing theories, it was decided to test copper and copper with various additions of zinc in plane bending fatigue, and determine to what extent their fatigue properties are dependent upon their ability to cross slip. EXPERIMENTAL PROCEDURE Fatigue tests were carried out on polycrystalline OFHC copper and various copper-zinc alloys with 5, 15, 20, 30, and 35 wt pct Zn. The Cu-Zn alloys were obtained commercially in sheet form. Test specimens, as shown in Fig. 1, were machined from sheet that had been reduced to an approximate grain size of 0.040 mm, the heat-treatment procedure being similar to that of Feltham anc copley.8

Jan 1, 1963

By J. Kostuik, M. J. de Bastiani

The method of mining the uranium orebody at Denison mine in the Elliot Lake district of Ontario is discussed. Details of the development of haulage-ways and rooms are giuen, as well as methods of ore haulage. The milling and beneficiation stages are noted briefly. Located in the Elliot Lake mining district of Canada, the Denison mine lies between Sudbury to the east and Sault Ste. Marie to the west, Fig. 1. This area is overlain by Huronian sediments filling a basin in the Keewatin basement complex. Subsequent folding of these sediments developed a synclinal structure, known locally as the Quirke Lake Trough, which forms the northern portion of the now famous big Z structure, the geological trademark of this region. Plunging to the west and extending nine miles from the granite outcropping in the north to the Keewatin lava contact to the south, the syncline has dips of 10" to 45" on the north limb and 10" to 20" on the south limb, Fig. 2. The Denison orebody, located on the north limb of the syncline, lies between 800 and 3000 ft below the surface. The main ore zone is within 50 ft of the Missisagi-Keewatin contact (see Fig. 3), and it varies in thickness from 6 to 32 ft. This ore zone has a composite dip of 19" to the south and is composed primarily of two pyritized conglomerate beds separated by a band of pebbly quartzite stratum. The upper conglomerate bed of the main zone varies from 2 to 15 ft in thickness, while the lower conglomerate varies from 2 to 12 ft. Over extended areas, mining is carried out over the entire section comprising the two conglomerate layers and intercalated quartzite; in other areas, the conglomerates are selectively mined, depending on the thickness of the quartz-ite. The ore zone contains approximately 6 pct sulfides and 80 pct silica. The predominant uranium-bear ing minerals are brannerite, uraninite, and an unidentified mineral (probably thucolite). Some of the uneconomic metallic minerals identified to date are pyrite, pyrrhotite, and chalcopyrite. MINE LAYOUT Shafts: Underground exploration and development (Fig. 4) were performed through two shafts 1/2 mile apart. The No. 1 shaft, sunk to a depth of 1856 ft with the main station established at the 1600-ft ore level, is a 250-sq ft, 5-compartment shaft capable of handling 4500 tpd. Two 5x5-1/2-ft compartments serviced by a 12-ft diam double-drum hoist, are used for hoisting 190-cu-ft bottom-dump skips in balance. A 30-man cage, offset by a counterbalance, operates within a 5-1/2x11-1/2-ft compartment. The fifth compartment is used as a manway, pipe, and cable-way. This shaft is also provided with an intermediate pumping station used during the initial mining phase as well as loading-pocket and spill-pocket installations and pumping facilities. The 480-sq ft No. 2 shaft, located down-dip from No. 1 shaft, has a 2776-ft depth. Serving as the main hoisting shaft with a maximum capacity of 9000 tpd, it has eight compartments, four of which are used for hoisting ore in 10-ton capacity bottom-dump skips. These four compartments, measuring 5x6-ft in section, are serviced by two 12-ft diam double-drum semi-automatic hoists. A 50-man cage is operated in a 15x8-ft compartment, and it is counterbalanced by a 12-ton weight. Two compartments are used for ventilation, manway, electric cables, and pipes. A station was constructed at the 1325-ft horizon of the No. 2 shaft for relay pumping. The main sta-

Jan 1, 1961

By J. R. Dawson, J. A. Glunt

VESTA No. 4 and 5 mines supply most of the high volatile coal required for Jones & Laughlin's byproduct coke plants. Until 1944 all coal produced in these mines was loaded by hand. Pressure to mechanize at Vesta was brought on by the manpower shortage in World War 11, and in 1943 a decision was made to change to mechanical loading. It was the opinion of those concerned that by use of selective mining the coal would be clean enough to be processed by the existing preparation plants. At that time coal cleaning at Jones & Laughlin consisted of hand picking the larger sizes at the mine tipples and wet washing at plants located at the Pittsburgh and Aliquippa Works. A typical section of the Pittsburgh seam as it occurs at Vesta is shown in Fig. 1. The 14 to 16 in. of drawslate immediately over the coal is too soft to be used for roof and must be taken down, leaving the rooster coal just above for roof. In selective mining an attempt is made to mine the coal under the drawslate first and then load out the drawslate separately as refuse. In full seam mining slate and coal are shot down at the same time and loaded out. Mechanical loading at Vesta created new problems. It was found that selective mining, even when carefully applied, increased the refuse in the raw coal to such an extent that the washers at the mills were unable to handle it. Large tonnages of refuse were being transported, a reject disposal problem was created, and good metallurgical coal was being loaded out and lost. In addition, the selective method handicapped operation of the loading machines. Fig. 2 shows how the ash content of the raw coal increased as mechanization proceeded at No. 5 mine. New Preparation Plant: After an exhaustive study of this situation by J & L management, is was decided that a preparation plant should be constructed at the mines, capable of handling full seam mined coal and large enough to process full production of both No. 4 and No. 5 mines. It was decided that from the standpoint of inside operation and haulage the new plant should be located as near as possible to the old No. 5 tipple on the Monongahela River near Vestaburg. Full seam mining at Vesta could be expected to produce a tremendous volume of refuse along with the coal, since 14 to 16 in. of slate would be mined with an average 78 in. of coal. Disposal of this refuse required serious consideration in selecting a plant site. The only area near Vestaburg large enough for disposal of the plant refuse was across the Monongahela River at LaBelle. Ample space was available here, and construction at this location would not interfere with operation at the No. 5 tipple. It had been noted over the years that the sulfur content of the coal produced at Vesta was gradually increasing, and there was every indication that this increase could be expected to continue as the mines developed further from the river. It was decided to construct a three-part separation plant with complete blending facilities to produce both metallurgical and steam coal. In this system an attempt is made to reduce the sulfur in the metallurgical coal in a secondary cleaning process. Construction of the new preparation plant got under way early in 1947 and continued until the spring of 1950. The plant was designed to handle 2400 tph of raw coal containing 30 to 35 pct refuse. Cost of outside buildings and equipment was about $14 million. Total cost of the plant, together with improvements and changes made at this time, was approximately $21 million.

Jan 1, 1957