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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 Our calculation (much less precise than the present work) was based on: (1) Unpublished work on the sulphur content of molten iron (1.5 pct at 1500°C) in equilibrium with graphite and an iron sulphide slag; (2) the distribution coefficient of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 J. B. Hess, C. S. Barrett

TO reach equilibrium between different phases in cobalt-rich alloys requires prohibitively long annealing cobalt-richalloystimes when temperatures are below about 700°C. The fact that a transformation from face-centered cubic to close-packed hexagonal readily tered takes place at temperatures below this in the cobalt-rich solid solutions is not an indication that thermally activated processes occur at an appreciable rate, for the transformation is well established as martensitic in nature. Wide divergence between heating and cooling experiments and high sensitivity to prior heat treatment make it difficult to judge temperatures of equilibrium between the phases.&apos; One object of the present work was to see if the information object of on the relative stability of phases could be gained by substituting plastic deformation for thermal agitation. Procedures were worked out that led to the determination of a diffusionless type of phase diagram, which represents the temperature of of phase equal stability for phases of the same composition, and the technique was applied to the Co-Ni system. Another object of the work was to see whether or not deformation would generate frequent stacking faults when these were thin lamellae of quentstackingfaultsa phase having higher free energy than the parent phase. The alloys were prepared in 80 to 100 g melts from cobalt (with metallic impurities estimated spectrochemically as follows: Ni, 0.05 pct; Fe, 0.001 pct.; Mg, Si, Cu, Cr, Al, < 0.001 pct) and Mond Car-bony1 nickel (with Fe, 0.05 pct; Si, 0.003 pct; C, 0.61 pct.; Cu, 0.001 pct; Co, Cr not detected, < 0.01 pct). The metals were melted in pure Al2O3 crucibles. An atmosphere of argon, that had been purified by passing over hot magnesium chips, was used for the alloys that, by analysis of the portions of the ingots actually used, were found to contain 15.3, 25.7, and 35.0 pct Ni, and vacuum melting (after degassing) was used for those containing 29.4 and 31.5 pct Ni. After induction melting the alloys were allowed to solidify in the crucible, and slices % in. thick x ½ in. in diam were annealed 12 hr at 1350°C for homogenization. These same specimens were used throughout the series of experiments, with annealing treatments of 4 hr at 900°C in purified hydrogen followed by furnace cooling, alternating with the deformation and X-ray tests discussed below. Results Spontaneous transformation was observed on cooling to room temperature in all alloys containing 29.4 pct Ni or less and by cooling the 31.5 pct alloy to — 195°C but was not observed in the 35 pct alloys after cooling to —195°C. These results are in satisfactory agreement with the cooling experiments of Masimoto.4 From these data it is clear that the temperature of beginning transformation M,,, drops to 20°C with the addition of about 30 pct Ni. The test for spontaneous transformation was metallographic. Specimens were thermally polished by annealing 10 hr in hydrogen at 1350°C, then furnace cooled; if trans- formation had occurred there were relief effects visible on the thermally polished surfaces. These markings were narrow straight lines, usually resolvable at high magnification as clusters of fine lines that resembled slip lines. It was concluded that they resulted from displacements on (111) planes, for the number of directions in individual grains often reached but never exceeded four, and lines could always be found parallel to the thermally etched (111) boundaries of annealing twins. The markings were thus consistent with the idea that the transformation occurs by (111) plane displacements (Shockley partial dislocations moving on (111) planes). This was further confirmed by X-ray tests for stacking disorders. Using an oscillating crystal technique previously employed to detect strain-induced faulting in Cu-Si alloys," streaks indicative of the stacking faults were looked for and found on X-ray films of the spontaneously transformed 25.7 pct Ni alloys, as expected by analogy with Edwards and Lipson&apos;s results on pure cobalt." The streaks were much intensified after rolling at room temperature. Transformation induced by plastic strain was investigated as a function of alloy composition and temperature of deformation. A series of tests was made to determine suitable straining and X-raying techniques. Filing was found inferior to abrasion in converting cubic samples to hexagonal, and abrasion was less effective than peening in producing smooth unspotty Debye rings in the X-ray patterns. Because the diffraction lines were broad, Geiger-counter spectrometer records of filings were less sensitive in revealing small amounts of transformed material than X-ray patterns recorded on films in a small diameter camera. After these exploratory tests the following methods were adopted. Specimens that had been annealed at least 4 hr at 900°C and furnace cooled were mounted in a block of aluminum, brought to temperature, and peened thoroughly with a mullite pestle preheated to the same temperature. The specimens were then quenched to room temperature. In peening, a circular area of % in. diam was given 500 blows. A few control tests showed that an additional 1000 blows did not detectably change the proportions of the phases present. The amount of transformation was judged by X-ray reflection patterns from the peened surface, using the innermost four lines of the cubic and the hexagonal patterns with filtered CoKa radiation,

Jan 1, 1953

By A. A. Collins

H. R. BIANCO*—I should like to ask Mr. Collins if that statement he made about the addition of drosses to the blast furnace slowing down the blast furnace is a result of his own experience or a result of the experience of some older metallurgists; and perhaps I should ask him to define the type of drosses that he means. A. A. COLLINS (author&apos;s reply)— That has been my own personal experience with dross. On various occasions at Chihuahua we attempted to incorporate the dross in our regular blast furnace charge and to shut down the dross re-verberatory to try to save some money. As expected, we had very poor results. I think that Ed Fleming will well remember on one occasion, that was back about 1933, when we attempted the first experiment along this line, and as a result of the sulphur addition to the blast furnace to matte out the copper we ended up with hanging furnaces and mushy slags and abandoned the dross experiment, once again turning to the use of the reverbera-tory for handling dross. H. R. BIANCO—Is that dross you refer to from the drossing kettle ? A. A. COLLINS—Yes, the dross that I am referring to came from drossing kettles. Furthermore, to back up my previous assertion, I had occasion in 1943, while up at Leadville, to once again experience the routing of dross through the blast furnace with its sulphur addition, since they had no dross re-verberatory, and to observe that once thf dross was removed, the furnace was speeded up almost 100 tons a day. All of these are personal experiences and I think that Mr. Feddersen also has had a little experience along this line —in fact, I believe all of us have had some experience. H. R. BIANCO—I know at Trail they recirculate considerable dross through the blast furnaces and we also recirculate dross at Herculaneuin and I am not aware that it has done much towards slowing down the blast furnace. A. A. COLLINS—We have always had very poor results. In the first place you have got to add a sulphur addition to pick up that copper and once you do that, that sulphur is apt to combine with some of the zinc and you are going to form a little mush; before you know it you have furnace hangs and a poor working furnace. Now of course that depends on the amount of zinc you have on charge. But in 1943, Leadville had roughly about 7 pet zinc in their slag and it worked very poorly. Previously when they had 4 or 5 pet zinc in their slag it did not matter. B. L. SACKETT* At Tooele we had a great deal of experience with copper. We have always been able to keep a lead well, however, in spite of the fact we have run as much as 5 pet copper and only 15 pet lead on the charge. But regarding the handling of dross, our dross reverberatory furnace is only 7 or 8 years old. Before that we recirculated the dross through the furnace and thought we were doing a pretty nice job. Of course these things are all more or less relative—in other words you establish a certain condition much better than one of a few years ago and possibly as good as any other of which you know and you think you have pretty good results. When we first took the dross off of the blast furnace and put it through the dross reverberatory furnace we immediately found out that we had gained something very real in furnace speed. Since that time there have been occasions when, because of the dross reverberatory being down, we have had to use dross again through the blast furnace and that has checked our original experience in slowing down the furnace very definitely. So we feel that a dross reverberatory is a very valuable asset at the Tooele Plant. A. A. CENTER*—Mr. Sackett&apos;s being here reminds me of trying to run with a minimum of lead concentrates the maximum of dross producing electrolytic zinc plant residue. He came up from International Smelting Co. to help us get started on that. We took an old copper blast furnace at Great Falls, Montana, and made a lead furnace out of it by putting a lead well on the other long side which of course is a very unorthodox lead blast furnace. Our aim was to treat the residue from the electrolytic zinc plant, as I said, with a minimum of lead concentrates. That meant a maximum amount of dross. At that time selective flotation was not general practice, so our zinc concentrates ran relatively high in copper and other dross-producing elements; and of course these were largely in the zinc plant residue. I think we might call it muscle metallurgy, but we had an interesting, successful experience there and we ran for over a year thanks to Mr. Sackett&apos;s helping us get started. I have the details, but time does not permit. We did well enough so that the A. S. and R. Co. at East Helena kept boosting up the offer to us for the electrolytic zinc plant residue and there was not enough lead concentrate to supply two lead smelters there in Montana, so the matter finally finished up by the A. S. and R. Co. taking all of the residue under long term contracts.

Jan 1, 1950

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&apos;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 &apos;trombone&apos; 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&apos; 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 P. T. Chiang

Chemical etching methods for the simultaneous revealing of the tellurium or selenium Phase and the chalcogenide grain boundaries of the alloy systems are given. A tellurium eutectic was found Present in zone-melted ingots. Similarly, a selenium monotectic was present in ingots. In general, the second phase (tellurium or seleniumn) occubies three different sites; viz., along the chalcogenide grain boundaries, as inclusions within the chalcogenide grain, and on the undersurface of the ingot. The detection limit for the tellurium phase is about 1 u in width. THERMOELECTRIC materials based on Group V (bismuth, antimony) and Group VI (selenium, tellurium) elements have aroused considerable interest in recent years in the practical application of thermoelectric cooling. In many cases, a small amount of excess tellurium (or selenium) was added to the material to optimize its thermoelectric properties. Then the question immediately arises as to the number of phases present in the resultant alloy. In the binary systems of Bi-Te, Sb-Te, and Bi-Se, the congruent melting compositions have been reported to be non-stoichiometric and are represented by Bi~Te respectively. It is to beexpected and known that Bi2Te3 and SbzTe3 crystallize from the melt with an excess of bismuth and antimony in the lattice and that tellurium forms a eutectic.~&apos; The same could be assumed to take place in the pseudo binary systems of (Bi,Sb)zTe3 and Bi2(Se,Te)3 as well as in the system studiedby puotinen5 and other workers. Likewise, BiaSe3 crystallizes from the melt with an excess of bismuth in the lattice and selenium forms a monotectic.~ Therefore, in practice, alloys solidified from the melt often contain a second phase (tellurium or selenium) in one region or another of the solid mass even without the addition of excess tellurium (or selenium). ~u~~recht&apos; studied the thermoelectric properties of (Bi,Sb)2Te3 alloys with excess tellurium and simultaneous additions of selenium. He mentioned that the materials show two phases because of the considerable excess of tellurium or selenium. However, he did not report as to how the tellurium or selenium phase was identified. It is generally believed that the presence of an excessive amount of tellurium or selenium phase in the alloy would adversely affect its thermoelectric properties and its uniformity. Consequently, there is a need for a simple method for the identification of the tellurium and selenium phase. The quantity of the second phase present is usually too small to be detected either by chemical analysis or by normal X-ray techniques. This investigation was therefore carried out, first, to devise a simple metallographic method for the identification of the tellurium or selenium phase coexisting with the chalcogenides and, second, to determine the distribution and specific location of the tellurium or selenium phase in the ingots. EXPERIMENTAL PROCEDURE The starting materials used for the alloy preparations were 99.999 pct pure bismuth, antimony, and tellurium and 99.997 pct pure selenium. The bismuth and antimony were obtained from Consolidated Mining and Smelting Co. of Canada Ltd., while the selenium and tellurium were obtained from Canadian Copper Refiners Ltd. The tellurium was purified further in the laboratory by zone refining. The elements were pulverized in a stainless-steel pestle and mortar. The amounts for the desired composition were weighed out each time on an analytical balance to make up a 100-g sample. Then the sample was introduced into a Vycor ampule (19 by 150 mm), pumped down to a vacuum of 10"5 Torr for 15 min, and sealed off. The ampule was then heated in a horizontal resistance furnace at 800" to 900°C for about 20 hr. During this period the assembly was rocked back and forth several times to ensure good mixing. At the end of the heating period, the ampule was quenched in cold water and then transferred to the zone-melting apparatus described in a previous publications to grow large-size aligned polycrystals. The background and ring-heater temperatures were adjusted to make the freezing solid-liquid interface slightly convex to the liquid. The recorded temperature gradient in the vicinity of the freezing solid-liquid interface was around 15°C per cm. The ampule was moved horizontally at a speed varying from 0.4 to 2 cm per hr so that the ring heater would cover the whole ingot length from end to end. A single zone-melting pass was used for the Bi-Te, Sb-Te, and Bi-Sb-Te ingots. Two passes in the forward and reverse directions were carried out for the Bi-Se and Bi-Se-Te ingots. Six passes in the forward and reverse directions were performed for the Bi-Sb-Se-Te ingot. The zone-melted ingots were found to contain several large crystals, with their basal planes (0001) approximately parallel to the growth axis. Samples of bismuth and antimony tellurides coated with a layer of tellurium, and bismuth selenide coated with a layer of selenium, were prepared for comparison in phase identification. These coatings were made by dropping a piece of the zone-melted ingot into some molten tellurium or selenium under argon atmosphere and allowing them to cool slowly to room temperature. The metallographic specimens were prepared by

Jan 1, 1969

By J. B. Newkirk, A. H. Geisler

INTRODUCTION CERTAIN of the permanent magnet alloys provide ideal systems for the study of the kinetics of the precipitation reaction and the correlation of structure with properties. One such system, Cu-Ni-Fe, was found by Bradley1,2 to exhibit a coherent transition state in the precipitation process analogous to that reported for Al-Cu alloys somewhat earlier.3 The attractiveness of some permanent magnet alloys for study lies in the fact that vertical sections of the ternary phase diagram in certain regions of composition (Fig I) have as their prototype the binary Ni-Au diagram. Alloys of this type decompose into products that have the same crystal lattice type but only slightly different lattice parameters. The advantages that such alloy systems offer for study over the usual in which an intermetallic compound is formed are many: I. Since the precipitate has the same crystal structure as the matrix, complex atomic movements are not required to form the new lattice. 2. Similarly, complex orientation relationships are not involved for both the matrix and the precipitate would be expected to have the same orientation. 3. Small disregistry of the decomposition products at equilibrium (in contrast with Cu-Ag alloys) is conducive to extensive coherent growth in the transient state and thus the transition lattice can be detected by the usual X ray diffraction methods. 4. Finally, the relative quantities of precipitate and depleted matrix can be varied from o to 100 pct* thus permitting wide freedom for the study of the effect of composition on coherent growth and properties. In the Cu-Ni-Fe alloys of appropriate composition, the face-centered cubic precipitate and also the depleted matrix when first formed are coherent with the parent matrix.1,2 The two have the same [ao] parameter as the original matrix but they are both tetragonal; the precipitate has an axial ratio c/a < I while that of the depleted matrix is c/a > I. When coherency is lost they assume the normal face-centered cubic structure with the depleted matrix having a lattice parameter greater than the original matrix and that of the precipitate less. Such a mechanism would also be expected for Cu-Ni-Co alloys because of the similarity in constitution but this had not been demonstrated. The present investigation was conducted on a Cu-Ni-Co alloy. The constitution diagram and magnetic properties of these alloys have been fairly well established 4,5 however, no previous determinations of mechanism of precipitation and no correlation of structure with properties had been made. Thus, an alloy of this system was chosen for a comprehensive investiga-

Jan 1, 1948

By George D. Dub

Owens Lake is at present a source of important nonmetallic minerals, sodium carbonate (soda ash, Na2CO3); sodium sesquicarbonate (trona, Na2CO3.NaHCO3.-2H2O) and borax, (Na2B4O7.10H2O). Owens Lake is a closed basin in the southern part of Inyo County, California, at the southern end of Owens Valley, east of the Sierra Nevada Mountains and west of the Coso and Inyo Mountains. Broadly considered, it is in the Great Basin area, but at no time was it a part of Lake Lahontan. Closed Basins Closed basins are phenomena of arid or semiarid regions where outflow is considerably less than inflow and where accordingly soluble salts concentrate in residual liquid. When inflowing waters originate in areas where the rocks are predominantly marine sediments, the residual basin liquid is primarily a chloride, brine; if the rocks are largely igneous, the residual brine tends to be alkaline containing carbonates, and sometimes borates as well. Since normally, both types of rocks occur in regions contiguous to closed basins, residual liquids do not often fit into the two broad divisions mentioned. No two residual brines are exactly alike, just as no two ore deposits are exactly alike. Even out of the same closed basin, it is possible to get widely different analyses of liquids since underground and surface flows, as well as local evaporation rates and other conditions might have a marked influence on residual-liquid compositions. At Searles Lake, The American Potash and Chemical Corporation is building a plant to process a lower-level brine which is considerably higher in sodium carbonate and borax, and lower in potassium chloride, than that company has processed for many years. F. W. Clarke1 has classified waters of closed basins as follows: I. Chloride type; largely sodium chloride (NaC1) and of oceanic type, such as Great Salt Lake. Related is -the Dead Sea, a bittern residue of magnesium, potassium and sodium chlorides. a. Sulphate type; largely sodium sulphate (Na2SO4) with considerable sodium chloride such as Sevier Lake, Utah; Laramie Lakes, Wyoming; Dale Lake, California. 3. Carbonate type; high in carbonates and fairly high in sulphates, such as Moses Lake, Eastern Washington; and the Nebraska Potash Lakes. 4. Carbonate—chloride type; lower in carbonates than type 3 and about equally rich in sulphates, such as Pyramid Lake, Nevada. 5. Sulphate—Carbonate type; quite high in sulphates and carbonates, such as Pelican Lake, Oregon. 6. Triple type; considerable quantities of carbonates, sulphates and chlorides

Jan 1, 1948

By George D. Dub

Owens Lake is at present a source of important nonmetallic minerals, sodium carbonate (soda ash, Na2CO3); sodium sesquicarbonate (trona, Na2CO3.NaHCO3.-2H2O) and borax, (Na2B4O7.10H2O). Owens Lake is a closed basin in the southern part of Inyo County, California, at the southern end of Owens Valley, east of the Sierra Nevada Mountains and west of the Coso and Inyo Mountains. Broadly considered, it is in the Great Basin area, but at no time was it a part of Lake Lahontan. Closed Basins Closed basins are phenomena of arid or semiarid regions where outflow is considerably less than inflow and where accordingly soluble salts concentrate in residual liquid. When inflowing waters originate in areas where the rocks are predominantly marine sediments, the residual basin liquid is primarily a chloride, brine; if the rocks are largely igneous, the residual brine tends to be alkaline containing carbonates, and sometimes borates as well. Since normally, both types of rocks occur in regions contiguous to closed basins, residual liquids do not often fit into the two broad divisions mentioned. No two residual brines are exactly alike, just as no two ore deposits are exactly alike. Even out of the same closed basin, it is possible to get widely different analyses of liquids since underground and surface flows, as well as local evaporation rates and other conditions might have a marked influence on residual-liquid compositions. At Searles Lake, The American Potash and Chemical Corporation is building a plant to process a lower-level brine which is considerably higher in sodium carbonate and borax, and lower in potassium chloride, than that company has processed for many years. F. W. Clarke1 has classified waters of closed basins as follows: I. Chloride type; largely sodium chloride (NaC1) and of oceanic type, such as Great Salt Lake. Related is -the Dead Sea, a bittern residue of magnesium, potassium and sodium chlorides. a. Sulphate type; largely sodium sulphate (Na2SO4) with considerable sodium chloride such as Sevier Lake, Utah; Laramie Lakes, Wyoming; Dale Lake, California. 3. Carbonate type; high in carbonates and fairly high in sulphates, such as Moses Lake, Eastern Washington; and the Nebraska Potash Lakes. 4. Carbonate—chloride type; lower in carbonates than type 3 and about equally rich in sulphates, such as Pyramid Lake, Nevada. 5. Sulphate—Carbonate type; quite high in sulphates and carbonates, such as Pelican Lake, Oregon. 6. Triple type; considerable quantities of carbonates, sulphates and chlorides

Jan 1, 1948

By George D. Dub

INTRODUCTION OWENS LAKE is at present a source of important nonmetallic minerals, sodium carbonate (soda ash, Na2CO3); sodium sesquicarbonate (trona, Na2CO3.NaHCO3.2H20) and borax, (Na2B407.10H2O). Owens Lake is a closed basin in the southern part of Inyo County, California, at the southern end of Owens Valley, east of the Sierra Nevada Mountains and west of the Coso and Inyo Mountains. Broadly considered, it is in the Great Basin area, but at no time was it a part of Lake Lahontan. CLOSED BASINS Closed basins are phenomena of arid or semiarid regions where outflow is considerably less than inflow and where accordingly soluble salts concentrate in residual liquid. When inflowing waters originate in areas where the rocks are predominantly marine sediments, the residual basin liquid is primarily a chloride brine; if the rocks are largely igneous, the residual brine tends to be alkaline containing carbonates, and sometimes borates as well. Since normally, both types of rocks occur in regions contiguous to closed basins, residual liquids do not often fit into the two broad divisions mentioned. No two residual brines are exactly alike, just as no two ore deposits are exactly alike. Even out of the same closed basin, it is possible to get widely different analyses of liquids since underground and surface flows, as well as local evaporation rates and other conditions might have a marked influence on residual-liquid compositions. At Searles Lake, The American Potash and Chemical Corporation is building a plant to process a lower-level brine which is considerably higher in sodium carbonate and borax, and lower in potassium chloride, than that company has processed for many years. F. W. Clarke1 has classified waters of closed basins as follows: I Chloride type; largely sodium chloride (NaCI) and of oceanic type, such as Great Salt Lake. Related is the Dead Sea, a bittern residue of magnesium, potassium and sodium chlorides. 2. Sulphate type; largely sodium sulphate (Na2S04) with considerable sodium chloride such as Sevier Lake, Utah; Laramie Lakes, Wyoming; Dale Lake, California. 3. Carbonate type; high in carbonates and fairly high in sulphates, such as Moses Lake, Eastern Washington; and the Nebraska Potash Lakes. 4. Carbonate-chloride type; lower in carbonates than type 3 and about equally rich in sulphates, such as Pyramid Lake, Nevada. 5. Sulphate-Carbonate type; quite high in sulphates and carbonates, such as Pelican Lake, Oregon. 6. Triple type; considerable quantities of carbonates, sulphates and chlorides

Jan 1, 1947

By Andrew Mayer

EARLY in 1942 National Lead Co. was requested by the War Production Board to construct and operate a plant for the Government to produce magnesium by the ferrosilicon process which had been developed by Dr. L. M. Pidgeon in the laboratories of the Canadian National Research Council. A contract with the Defense Plant Corporation was concluded in May 1942 and Magnesium Reduction Co., a wholly owned subsidiary of National Lead Co., was formed to carry out this project. The capacity of the plant was rated at 10,000,000 lb. of magnesium per year, equivalent to an average production of about 14 tons per day. THE PROCESS The process, in brief, is as follows: The raw materials are dolomite and ferrosilicon, 75 per cent grade. The dolomite is calcined, the calcine and ferrosilicon are ground and mixed and the mixture is briquetted. The briquets are charged into tubular retorts of chrome-nickel steel set horizontally in a furnace with the open ends projecting outside the front wall. The retorts are then closed and evacuated. Magnesium is liberated according to the reaction 2(MgO, CaO) + Si = 2Mg + 2CaO, SiO2 It is distilled from the charge and condensed on a removable sleeve in the throat of the retort. Sodium and potassium, if present, are also liberated, distilled and condensed on a "sodium condenser" in the end of the retort. At the expiration of the distillation period the retorts are discharged and the cycle of operations is repeated. For the rated daily capacity of ±. tons of magnesium there are required approximately 170 tons of dolomite, equivalent to about 85 tons of calcine, and 16.5 tons of ferrosilicon. The retort residue, a bulky powder which at present is waste, amounts to about 87 tons. CHOOSING THE SITE The investigation to select the location of the plant was conducted by the Research Laboratories of National Lead Co. This work included geologic and economic surveys of a number of districts, examination of dolomite samples by chemical, spectrographic and petrographic methods and large-scale tests of dolomite from several of the more promising localities. In the last-named tests 2-ton samples were put through the ferrosilicon process in the pilot plant of the Canadian National Research Council, under the guidance of Dr. Pidgeon. The conclusions drawn from the investigation, in regard to the quality of dolomite to be used in the ferrosilicon process, were: First, the dolomite should contain not less than 21 per cent MgO

Jan 1, 1944

By N. Tiner

THE mechanical properties of magnesium-alloy castings are greatly improved by grain refinement, and at present considerable attention is being paid to methods of obtaining fine-grained castings. One method of achieving this end is the use of proper melting and superheating techniques. These have been employed commercially to a great extent, but no comprehensive study of the subject appears to have been published at present. Numerous investigations found in the literature include some reference to superheating techniques, 1-4 and Achenbach, Nipper and Piwowarsky,5 based principally on thermal analysis, ascribe the grain-refining action of superheating to the influence of solid foreign particles or dissolved gases present in the melt upon the course of crystallization. The primary purpose of this paper is to present the principal facts concerning the effect of superheating on grain size of magnesium alloys. Attempts are also made to explain the experimental results in terms of a general hypothesis. The investigations are limited chiefly to common commercial casting alloys. In order to carry on this work, it was necessary to determine the factors influencing grain size and to develop a proper technique for the preparation of samples. The preliminary tests made by the author indicated that the grain size of magnesium alloys resulting from the solidification of a melt is determined by alloy composition, presence of impurities, rate of cooling during solidification, thermal history of the melt and to some extent the structure of the charge used. The rate of cooling is in turn dependent upon the mass of the melt, pouring temperature, and dimensions, properties and temperature of the mold into which the melt is poured. The final grain size of the alloys is still further modified by heat-treatment subsequent to casting. It is not within the scope of the present paper to discuss all of these factors, and it would suffice to note here that when equal amounts of magnesium alloys are prepared from the same alloying materials, melted in graphite crucibles under the same thermal conditions, then cast into graphite molds at a constant temperature, the .resultant solids have the same grain size. By maintaining all other factors constant, it is thus possible to vary thermal conditions of a melt and to examine the grain size of resultant castings. PREPARATION OF SAMPLES AND GRAIN-SIZE TESTING The alloys investigated in this paper were taken from ingots made by Permanente Metals Corporation or prepared by the author with pure carbothermic magnesium and necessary alloying elements. Melting was carried out under a flux in a graphite crucible heated uniformly by an

Jan 1, 1945

By J. L. Gillson

In a brief summary of the many occurrences of fluorspar in our western states, it is not possible to go into detail in regard to the geology, mining and milling methods, and reserves about individual deposits. Rather a broad review is made of the occurrences and an analysis of how the western deposits may fit into our future domestic economy is presented. General Statement on Fluorspar Situation Fluorspar has been used for years as a flux in the manufacture of basic open-hearth steel, in foundry cupolas for melting pig iron, in the manufacture of opalescent glass and in the preparation of hydrofluoric acid. In 1880 the annual production of fluorspar in the United States was around 4000 tons per year. By 1900 it was up to 21,656 tons and between 1902 and 1908 it averaged about 40,000 tons annually. Prices were then so low, by present standards, that it seems impossible to believe now that mining could have been performed profitably at the price levels then prevailing. In 1890 fluorspar sold at the mines for $6.70 per ton, in 1901 for $5.81, in 1916 for 85.92. When the United States entered World War I, flyorspar production and price skyrocketed; the production reached 263,817 tons in 1918, with prices above $20.00 per ton, and prices were even higher in 1919 and 1920. After 1920, the domestic production fell from the 1918 level to about 125,000 tons annually and the average price dropped back to about $18.00 per ton. By 1935 the average price had fallen to $15.00, but in 1940 with a rising demand the average price of all grades had gone back up to $20.31. Today it is 75 per cent higher than in 1940 and presumably the price of acid grade at .least would be even higher if it were not for price ceilings. The first subsequent year that the domestic production exceeded that of 1918 was 1941, and the 1944 production is expected to exceed 400,000 tons, or over one and a half times that of 1918. In five-year periods, average production and imports were as are shown in Table I. Many people concerned with the consumption of fluorspar believe that the postwar period ahead will show a repetition of price trends of the last such period. They believe that imports will be again an important factor in the available supply and the price will fall back gradually toward the prewar level. The following basic factors are pertinent in analyzing this conclusion:

Jan 1, 1948

By J. L. Gillson

In a brief summary of the many occurrences of fluorspar in our western states, it is not possible to go into detail in regard to the geology, mining and milling methods, and reserves about individual deposits. Rather a broad review is made of the occurrences and an analysis of how the western deposits may fit into our future domestic economy is presented. General Statement on Fluorspar Situation Fluorspar has been used for years as a flux in the manufacture of basic open-hearth steel, in foundry cupolas for melting pig iron, in the manufacture of opalescent glass and in the preparation of hydrofluoric acid. In 1880 the annual production of fluorspar in the United States was around 4000 tons per year. By 1900 it was up to 21,656 tons and between 1902 and 1908 it averaged about 40,000 tons annually. Prices were then so low, by present standards, that it seems impossible to believe now that mining could have been performed profitably at the price levels then prevailing. In 1890 fluorspar sold at the mines for $6.70 per ton, in 1901 for $5.81, in 1916 for 85.92. When the United States entered World War I, flyorspar production and price skyrocketed; the production reached 263,817 tons in 1918, with prices above $20.00 per ton, and prices were even higher in 1919 and 1920. After 1920, the domestic production fell from the 1918 level to about 125,000 tons annually and the average price dropped back to about $18.00 per ton. By 1935 the average price had fallen to $15.00, but in 1940 with a rising demand the average price of all grades had gone back up to $20.31. Today it is 75 per cent higher than in 1940 and presumably the price of acid grade at .least would be even higher if it were not for price ceilings. The first subsequent year that the domestic production exceeded that of 1918 was 1941, and the 1944 production is expected to exceed 400,000 tons, or over one and a half times that of 1918. In five-year periods, average production and imports were as are shown in Table I. Many people concerned with the consumption of fluorspar believe that the postwar period ahead will show a repetition of price trends of the last such period. They believe that imports will be again an important factor in the available supply and the price will fall back gradually toward the prewar level. The following basic factors are pertinent in analyzing this conclusion:

Jan 1, 1948

By P. W. Ramsey, E. J. Boyle, E. A. Anderson

ONLY a few zinc-rich alloys are in commercial use today, and these depend largely upon solid solution effects for their desirable properties. Typical examples are rolled alloys containing about I per cent of copper, with or without additions of elements such as magnesium,1 and die-casting alloys containing about 4 per cent of aluminum with small amounts of magnesium and sometimes copper.2 A number of elements are capable of forming alloys with zinc, which do not enter into solid solution except in limited degree. The present paper considers one such system, zinc-titanium, in which alloys having potentially useful properties in the rolled form were found. ZINC-TITANIUM SYSTEM The zinc-rich portion of the zinc-titanium diagram, as determined by standard methods in this laboratory,3 is shown in Fig. I. The portion of the diagram of significance in the present work extends from 0 to about 0.5 per cent titanium. Gebhardt4 places the eutectic at 0.45 per cent, but the evidence in favor of the composition given in Fig. I is considerable. The solid solubility of titanium in zinc is very low indeed and possibly places these alloys in a somewhat different category from the commercial alloys previously mentioned. ALLOY PREPARATION All of the alloys used in this work were prepared with titanium metal of 98.9 per cent purity. In most cases high-purity zinc (99.99 per cent) was used but sufficient work was done later to show that zinc of about high-grade composition would yield approximately equal results. Best practice in alloy manufacture seemed to be to prepare a zinc-titanium hardener by adding about 4 per cent of titanium metal to molten zinc and holding under a boric acid cover, with frequent stirrings, at 750°C. until the titanium had dissolved. The desired titanium content could readily be obtained in the final alloy by adding the calculated amount of this hardener. The alloys proved to be quite stable in composition, unless overheated, and, with reasonable temperature control, required no cover or flux to prevent loss of titanium content by oxidation. HOT-ROLLING TESTS Three alloys containing, by analysis, 0.05, 0.12, and 0.23 per cent titanium were selected for preliminary study. Slab cross sections revealed a marked grain refinement* (Fig. 2) in the two alloys of highest titanium content. Cast unalloyed zinc has a columnar grain structure similar to that shown for the alloy of lowest titanium content but without the scatter of small equiaxed grains at the center.

Jan 1, 1944

By E. P. Klier

THE work of Klier and Lyman1 on the bainite reaction has led to the full description of this reaction for medium-carbon low-alloy steels. Certain experimental data reported by Klier and Lyman appear, however, to be improbable of obtaining in very high-carbon steels. The study herein reported consists of an analysis of the transformation characteristics of a 3 per cent chromium, 1 per cent carbon steel by means of X-ray and metallographic methods. From the results obtained it is concluded that the bainite reaction in this steel is more complicated than that reported by Klier and Lyman for the low-carbon low-alloy steels. However, the mechanism of bainite formation in this steel appears to be essentially the same as proposed for the low-carbon low-alloy steels with the modification that the first volumes to transform are carbon-rich rather than carbon-poor. EXPERIMENTAL PROCEDURE The analysis of the steel studied was: Cr, 2.89; C, 1.02; Mn, 0.33; Si, 0.35; S, 0.012; P, 0.020. The stock was received from the Carpenter Steel Co. in rolled bars. Specimen Preparation Specimens were austenitized at 1200°C. (2192°F.) for 25 min. in purified nitrogen. The resultant A.S.T.M. austenite grain size was 0-1. Isothermal transformation was allowed by quenching the specimens into liquid metal baths controlled to ±5°C. at 540°C. (1004°°F.) and above, and to ±2°C. at 490°C. (914°°F.) and below. Following the desired degree of transformation all specimens were water quenched. Metallographic Examination Specimens 1/4 in. round by 1/4 in. were used for metallographic examination. Etchants used were Vilella&apos;s reagent (1 gram picric acid, 10 ml. hydrochloric acid to make l00 ml. ethyl alcohol solution) and alkaline sodium picrate (electrolytic). Vilella&apos;s reagent was effective in revealing all microconstituents, while the alkaline sodium picrate developed the carbon-rich structures or carbides only. X-ray Examination Two X-ray techniques were used-the Debye-Scherrer method and a back-reflection method. The small specimens necessary for the Debye-Scherrer method were austenitized and then allowed to transform, or were cut from the larger metallographic specimens. The metallographic specimens were used for the back-reflection X-ray analysis. When it was desirable to determine the structure of the carbide present, sliver specimens were electrolytically etched to develop a carbide coating, then were exposed in the cameras. This technique was frequently necessary to allow positive identification of the carbide.

Jan 1, 1945

By L. D. Jaffe, J. H. Hollomon

The effect of tempering temperature and time upon the properties of quenched steel is clearly a subject of great practical importance, as well as of considerable theoretical interest. It would be very desirable to be able to predict the properties of a quenched steel when given any selected tempering cycle, with a minimum of experimental work upon that heat or even type of steel. Two problems are involved: one consists of relating the structure obtained to the physical properties, and will not be discussed here; the other, with which this paper deals, involves the relations of time, temperature and composition with the structure obtained 011 tempering. Indentation hardness is perhaps the most convenient quantitative measure of the tempered structure, or of what might be called the "degree of tempering." Undoubtedly, it is not an exact measure, and in some cases (as when temper-brittleness occurs) pieces of the same quenched and tempered steel may have widely differing properties at the same hardness. However, hardness has the advantage of being simple to determine and is certainly the most widely used indication of the degree of tempering. Moreover, for fully quenched steels that are not temper-brittle and do not contain retained austenite when quenched, it appears that, in general, if the hardness of the steel is the same, the tensile and impact properties will also be identical, whether a high or a low tempering temperature is used. The tempering charts available in the metallurgical literature usually give the hardness as a function of tempering temperature for only one tempering time. However, in practice, tempering times frequently vary, and a method of converting tempering curves for one time to tempering curves for some other time would be very valuable. It is also advantageous, in determining the tempering treatment necessary for a given part, not to have to temper samples at several temperatures for the time used for the commercial tempering operation. A more satisfactory procedure would be to employ times of the order of 10 to 30 min., rather than 5 or 10 hr., and if possible to use only one temperature. Furthermore, it is frequently desirable, for reasons such as temper-brittleness,* to obtain a given hardness with short times at relatively high temperatures, and it would be convenient to be able to compute without tests the time required to produce this hardness. Remarkably few quantitative studies of the hardness of quenched steel tempered over wide ranges of temperature and time

Jan 1, 1945

By Michel T. Halbouty

South Texas, one of the world&apos;s important oil and gas-producing areas, continued to hold interest during 1939 as another scries of new fields and new producing levels was developed within its boundaries at a fairly continuous rate throughout the year. There were 1801 wells drilled in this area in 1939, compared to 1678 in 1938; of these, 1515 were drilled in established fields, or as extensions to proven areas (semi-wildcat), and 286 were drilled as rank wildcats, in search of new production. Among these rank wildcats, 22 produced oil, 7 produced distillate and gas and 257 were abandoned as dry holes (see Table 5). Of the 1515 field and cxtension wells, 1213 produced oil, or distillate and gas, 50 produced only gas and 252 were dry; 38 found new productive sand levels in proven ficlds. Jim Wells County led this district in discoveries, having had 9 new fields arid 9 new productive sands found within its border during 1939. Most activc of the 1939 discoveries were Reynold?; and Ben Bolt fields in Jim Wells County and Minnie Bock field in Nueccs County; among producing fields the most activc were: East White Point and Aransas Pass (McCampbell). Production from South Texas in 1939 totaled 55,105,506 bbl. against a total of 56,152,122 bbl. produced in 1938. South Texas should continue its present rate of developrnent and exploration and this should result in the discovery of many fields in its numerous oil-bearing zones, and should also lead to more knowledge of the possibilities of its deeper horizons, such as the Mt. Selman and Wilcox formations; with the latter holding particular interest at the present time. Tide Water Associated Oil Company&apos;s A. M. Bruni No. I, Cole-Bruni field, Webb County, the deepest well drilled along the Mirando trend, was carried to 8959 ft., where it started to blow out from a Wilcox sand topped at 8954 ft., making gas, distillate and mud under heavy pressure. After the well was killed, drill stem beeame stuck and was cemented in place in order to test a Queen City (Mt. Selman) sand logged from 5620 to 5675 ft. This and other shows found in other wells along this trend, but out of this district, have served to increase the importance of this but little explored Sparta-Wilcox trend of South Texas during the year, and more activity along it is to be expected in the near future.

Jan 1, 1940

By Michel T. Halbouty

South Texas, one of the world&apos;s important oil and gas-producing areas, continued to hold interest during 1939 as another scries of new fields and new producing levels was developed within its boundaries at a fairly continuous rate throughout the year. There were 1801 wells drilled in this area in 1939, compared to 1678 in 1938; of these, 1515 were drilled in established fields, or as extensions to proven areas (semi-wildcat), and 286 were drilled as rank wildcats, in search of new production. Among these rank wildcats, 22 produced oil, 7 produced distillate and gas and 257 were abandoned as dry holes (see Table 5). Of the 1515 field and cxtension wells, 1213 produced oil, or distillate and gas, 50 produced only gas and 252 were dry; 38 found new productive sand levels in proven ficlds. Jim Wells County led this district in discoveries, having had 9 new fields arid 9 new productive sands found within its border during 1939. Most activc of the 1939 discoveries were Reynold?; and Ben Bolt fields in Jim Wells County and Minnie Bock field in Nueccs County; among producing fields the most activc were: East White Point and Aransas Pass (McCampbell). Production from South Texas in 1939 totaled 55,105,506 bbl. against a total of 56,152,122 bbl. produced in 1938. South Texas should continue its present rate of developrnent and exploration and this should result in the discovery of many fields in its numerous oil-bearing zones, and should also lead to more knowledge of the possibilities of its deeper horizons, such as the Mt. Selman and Wilcox formations; with the latter holding particular interest at the present time. Tide Water Associated Oil Company&apos;s A. M. Bruni No. I, Cole-Bruni field, Webb County, the deepest well drilled along the Mirando trend, was carried to 8959 ft., where it started to blow out from a Wilcox sand topped at 8954 ft., making gas, distillate and mud under heavy pressure. After the well was killed, drill stem beeame stuck and was cemented in place in order to test a Queen City (Mt. Selman) sand logged from 5620 to 5675 ft. This and other shows found in other wells along this trend, but out of this district, have served to increase the importance of this but little explored Sparta-Wilcox trend of South Texas during the year, and more activity along it is to be expected in the near future.

Jan 1, 1940

By William J. Craig, Harold L. Walker

INDUSTRY has for many years recognized the dependence of certain mechanical and physical properties, as well as workability, upon grain size variations in brass. Although the dependence of properties upon grain size has been recognized there are not many instances where a planned program of testing has been conducted and where the only intentional variable has been grain size. In particular the literature is limited with respect to the effect of grain- size on the endurance limit, though there are many industrial applications where the brasses are subjected to cycles of repetitive stresses. LITERATURE A number of the physical and mechanical properties of a given brass sheet, in the annealed condition, will depend upon: (I) The ready-to-finish grain size before final rolling, (2) the percentage deformation by cold working, and (3) the time and temperature of annealing, The first and second variables affect the recrystallized grain size, and the third variable affects the coalesced grain size. Gibbs1 has shown the effect of two different ready-to-finish grain sizes upon the tensile strength and hardness of deep drawing brass strip after a definite amount of cold rolling. The smaller the ready-to-finish grain size, the higher the tensile strength and hardness, but with lowered ductility, Gibbs also demonstrated, for a given time and temperature of anneal, that the tensile strength, hardness, and elastic limit are higher for the smaller ready-to-finish grain size, but the elongation is lowered, Townsend and Greenall2 have reported data which show that cold working increases the tensile strength, hardness, and load carrying ability of deep drawing brass sheet under reversed stress. Greenall and Gohn3 have reported data which show that the smaller grain sizes have higher tensile strength and hardness properties, and the data also show that increased cold working increases the tensile strength and hardness properties of deep drawing brass sheet. Kommers4 has reported data for tests of brass of known ready-to-finish grain size, a specific amount of cold reduction, and annealed at different temperatures above and below the recrystallization temperature. Kommers&apos; results indicate that annealing below the recrystallization temperature does not change the tensile strength and hardness properties, but the endurance limit may be increased from 22,000 to 28,000 psi. The increased endurance limit is explained as resulting from the reduction of internal stresses by a stress-relief anneal. Burghoff5 and associates have reported data on creep tests, over

Jan 1, 1948