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By Will H. Coghill, Fred D. DeVaney, R. G. O’Meara

The size of ball mills in the ore-dressing industry has increased from about 4 ft. in each dimension to 10.5 ft. in diameter by 8 ft. in length. In the cement industry they are as long as 45 ft. Plainly the trend has been toward larger mills. The common belief is that the big mills economize in floor space and first cost, but the advantage with respect to friction has not been clearly defined. In the ball-mill studies at Rolla, Mo., laboratory mills as small as 2 by 2 ft. have been found as efficient in terms of the amount of grinding per unit of net power as mills 6 by 4 or 6 by 12 ft. Therefore, if supremacy can be claimed for the big mills, advantage must come from other sources. The advantage to be considered here is reduced loss by friction or dead load. In the study to determine whether big mills have relatively less loss by friction the commercial catalogs were used for fundamental figures. The dimensions and weights of Table 1 are taken from the catalogs (the last line gives data on a Rolla laboratory mill). The net power—that is, the power expended within the mill for grinding—is calculated by a new formula.' The assumed speeds are each at 70 per cent of their theoretical critical speeds, and the mills are loaded to the center with balls. In other words, they are assumed to be working under analogous conditions and are set for good grinding. In the last column the ratio B: A shows that the larger commercial mills are lighter per unit of net power than the smaller mills, and more of the motor output should therefore go into grinding. The laboratory mill seems to have some advantage over the larger mills; it would be expected to be light because it has no liners.

Jan 1, 1935

By S. R. B. Cooke, H. S. Choi, I. Iwasaki

In magnetic taconite of the East Mesabi range, quartz, magnetite, and cummingtonite (OH}(Fe, Mg)(Si4O11)2) are the three major ore-forming minerals.' Fine grinding and magnetic separation of the taconite leads to a concentrate analyzing 64 to 65 pct Fe and 8 to 9 pct insoluble; the gangue content of the concentrate arising from the presence of locked magnetite-gangue particles, and of occluded gangue particles in magnetic floccules. To further upgrade such taconite concentrates, flotation appears to offer a particularly promising approach. To evaluate its effectiveness in separation, a knowledge of the flotation characteristics of each mineral comprising the concentrate becomes desirable. The floatability of magnetite and of quartz have already been reported in a previous paper,2 in which the flotation characteristics of the two minerals were correlated with the electrokinetic properties at the mineral-solution interfaces. Since cummingtonite has somewhat intermediate chemical composition between the above-named oxides, the electrokinetic and the flotation behavior of this mineral becomes of interest for both practical and fundamental reasons. In the present article the results of electrophoretic mobility measurements and of Hallimond tube flotation tests on cummingtonite are reported and are compared with those of quartz and magnetite. MATERIALS AND METHODS A sample of cummingtonite-rich taconite, received from the Mitchell pit, near Babbitt, Minn., was crushed and ground through 100 mesh, and the product was fractionated magnetically into three portions. First, a drum-type laboratory cobber was used to separate a highly magnetic portion, consisting mainly of free, but also containing locked, magnetite particles. The cobber tailing was then passed in two stages through a Franz isodynamic separator. The middle fraction was isolated, and, upon examination petrographically, was ascertained to be principally cummingtonite. The 100 to 150 mesh fraction of this sample was selected for the present study. It was cleaned by leaching briefly with hot, dilute HC1 and then by rinsing repeatedly with demineralized water. A ground and elutriated fraction of -5p cummingtonite was used for the electrophoretic measurements.

Jan 1, 1961

By H. J. Oberson, J. H. Brown

Comminution devices such as rod or ball mills are characteristically nonselective in their operation in that all material fed to a mill is ground to some extent. This is unfortunate when the object of comminution is liberation, for not only is energy expended in breaking pure gangue particles wasted, but, since gangue is frequently the predominant phase, the opportunity for such loss is great. Ideally, in this application a comminuting device should develop fracture only in locked particles. In the present investigation an attempt was made to use the large density difference which may exist between minerals to prevent comminution of one while grinding the other. EXPERIMENTAL MATERIALS AND EQUIPMENT The two minerals used in this work were -8-mesh galena and -8+ 20-mesh quartz. All tests were made on 2000-g charges containing 50 pct by weight of each mineral. Heavy media suspensions consisted of -100-mesh ferrosilicon in water. For grinding, a laboratory mill 10 in. diam by 12 in. long was filled to about 30 pct of its volume with either a ball or a rod charge. The balls ranged in size from 3/8 to 1 1/8 in. in diam; the rods ranged from 5/8 to 1 1/4 in. diam by 11 in. long. EXPERIMENTAL PROCEDURE In all tests employing heavy media, the ground pulps were treated as follows: 1) After recovering the ground charge from the mill but before drying, the pulp was wet screened on 200 mesh. 2) After drying the + 200-mesh product from step 1, the ferrosilicon was removed magnetically. 3) The nonmagnetic quartz and galena product from step 2 was sized on 10-mesh through 200-mesh screens. 4) The mineralogical composition of each screen fraction was determined by measuring its specific gravity. The specific gravity was related to the weight of galena in the fraction through the relationship

Jan 1, 1961

By S. S. Cole, D. L. Armant

The smelting of titaniferous ores for the past hundred years has not been successful because of thickening of the slags in the furnaces. The interest in the utilization of these iron bearing materials has been due to the extensive deposits which have been found in various parts of the world.1 The literature contains references2 to the use of titaniferous ores in blast furnaces. None of these attempts has met with success due to the lack of sufficient information as to the cause of failures and to economic considerations. Evaluation of the published information, as to the probable cause of failure in various attempts, has indicated that the presence of lower valence titanium oxide and formation of titanium carbide contributed to the failure. It was apparent that severe reducing conditions which are present in a blast furnace had been one of the factors for such failures. The previous smelting operations had as a primary objective the recovery of iron from these ores. Under these conditions the highest recovery of iron could only occur with severe reducing conditions. Application of Melting Point Data The application of the melting point data, from the system CaO-MgO-Al2O3-TiO2 reported in another paper,3 to the smelting of titaniferous ores required an extrapolation of these melting points due to the presence of other impurities from the ores. Likewise, a study of the effect of varying FeO and reduced titanium content of the slag was necessary before the data could be fully applied. The melting point data under oxidizing conditions had clearly indicated that a fluid slag with a melting point in the temperature range of 1370°C could be obtained. The smelting of a titaniferous ore had as a primary objective the production of a high titanium slag which was fluid and substantially free of iron and low in trivalent titanium. Since production of a high titanium slag was of primary importance, rather than the recovery of iron from these ores, the smelting concept could be altered to a marked degree. The ferrous oxide could be maintained at higher levels than is the normal practice in iron ores slags without being detrimental to the slags, although a loss in iron recovery resulted. During the reduction of the molten mixtures of ferrous iron and titanium, carbon acts as the reductant for changing the valence of the titanium and iron. The reaction can go further than in the case of the reduction which occurs in iron-titanium solutions. The formation of metallic iron removes iron from the system, so that the titanium oxide can be reduced to trivalent and divalent titanium oxides and also form titanium carbide which removes the titanium from the slags as a soluble constituent. The problem, therefore, resolves itself into maintaining the proper amount of ferrous iron in the slag thereby avoiding the formation of appreciable amounts of trivalent titanium and no divalent titanium oxide or titanium carbide. The latter two have been found in the viscous

Jan 1, 1950

By P. Niessen, S. J. Bercovici

RECENTLY Aust et al.&apos; proposed a model of non-equilibrium grain boundary segregation based on a vacancy gradient induced uphill diffusion process of solute to grain boundaries. According to their proposal, solute atoms which are attracted to vacancies would, during the cooling of a sample, migrate with vacancies toward grain boundaries (vacancy sinks) and thereby enhance the solute concentration in boundary regions. For solute atoms which are not attracted to vacancies no such boundary solute enrichment would be expected. These concepts were given experimental support by boundary and bulk mi-crohardness measurements of various thermally treated dilute zinc and lead alloys and also doped aluminum oxide. The underlying assumption in their work was that solute enrichment in boundary regions is indicated by higher microhardness values. From thermodynamic considerations1&apos;2 it appears that the solid-liquid solute distribution coefficient (K), defined as the concentration of solute in the solid to the concentration of solute in the liquid at equilibrium, indicates whether attractive or repulsive forces exist between vacancies and solute atoms. K < 1 indicates a solute-vacancy attractions, K > 1 indicates no, or a repulsive, vacancy-solute interaction. This note is to report on observations of different grain boundary etching phenomena in high purity zinc alloys. These phenomena appear to depend on whether a K < 1 or K > 1 solute is present. In experiments designed to measure directly the solute enrichment in grain boundary regions, samples of zone refined zinc with 0.1 at. pct solute additions were cooled from 380°C at a rate of 3°C per min. A surface layer of approximately 50 was then removed from the samples by chemical polishing. The samples&apos; surfaces were then protected by a chromating treatment such that in a subsequent etch only grain boundary regions were dissolved. The dissolved regions were approximately semicircular troughs, 10 µ in width. The grain boundaries were usually in the middle, i.e., at the bottom, of this trough. For zinc containing the K < 1 solutes cadmium and aluminum, the grain boundary was always indicated by a ridge as shown in Fig. 1. For pure zinc, and zinc containing the K > 1 solutes copper and gold, a groove at the bottom of the trough indicated the boundary position, Fig. 2., To test if these different grain boundary etching behaviors were possibly connected with the segregation process discussed above, the Zn-Cd alloy was again heated to 380°C and quenched into brine. After surface protection and etching it was found that this specimen no longer exhibited the grain boundary ridges but rather grain boundary grooves as shown in Fig. 2. It would appear, therefore, that the boundary ridge is a consequence of a process occurring during the slow cooling of the samples. From absorption spectrophotometric analysis it was found that for slowly cooled samples containing K < 1 solutes, the solute/zinc ratios of the material from the boundary trough were consistently considerably higher than the solute/zinc ratios of the bulk alloys. This boundary region solute enrichment was not observed in quenched samples. It is concluded therefore, that the boundary ridge shown in Fig. 1 is directly or indirectly connected with nonequilibrium solute segregation.

Jan 1, 1970

By E. A. Has, E. Mcl. Tittmann

Double roasting of sinter carrying a high percentage of lead concentrates, gave rise to the problem of removing the sheets of metallic lead formed in the wind boxes. The solution of the problem has been found in the water sealed wind boxes. The metallic lead is granulated and can be removed without seriously interfering with the operation of the machine and eliminates necessity of men going in wind boxes with air guns. THE practice of double sintering the charge for lead blast furnaces, which is now standard throughout the lead smelting industry, has introduced the rather difficult problem of cleaning the wind boxes on sintering machines used for the final pass. Two factors have more recently accentuated this problem: (1) the growing demand of workmen for better working conditions and less arduous labor, and (2) the growing dearth of crude lead ores and the improved grade of lead concentrates demand a charge carrying the maximum percentage of lead. One to 2 pct of such high lead charges (30 to 40) when processed on machines in reasonably good condition will pass through the grates and report in the wind boxes. One-half to three-quarters of this material will be metallic lead reduced by the coke breeze or other fuel mixed with the charge. This lead forms large cakes or chunks which are hard to remove and which substantially increase the time the machine is out of operation for cleaning the wind boxes. A high lead fall can reduce the operating time 30 pct. As a specific example, a 30 ft machine having 63 in. pallets, and equipped with three conventional wind boxes, requires 3 to 4 hr per day for removing the accumulation in the wind boxes when operated as a final pass machine. This same machine when operated on raw charge requires less than an hour per day for cleaning. To eliminate formation of large chunks of lead in the wind boxes of the final pass machines and time lost in cleaning the boxes, R. C. Rutherford of the Chihuahua smelter of the American Smelting and Refining Co., devised and patented* a wind box having a water bath to granulate the lead as it falls from the grate bars. Later, C. H. Harris of the San Luis plant made and patentedt an improvement on Rutherford&apos;s device in which the water bath seals the wind box in a manner permitting the latter to be readily cleaned without destroying the vacuum. The first machine to our knowledge to use the water bath box was at the Chihuahua plant along the lines of the Rutherford patent. In 1942, a machine was placed in operation at the El Paso plant which incorporated the features of both Rutherford and Harris. This machine is 40 ft long equipped with pallets 63 x 24 in., and having 4 wind boxes 10 ft 4 in. x 7 ft 2 in. x 51 ft 3 in. in size (fig. 1). The air from the fourth box is returned to the first three, and the gas from the first three discharged to a baghouse. The machine has conventional drive, discharge hopper,

Jan 1, 1951

By S. F. Reiter, You-Chao Liu, W. R. Hibbard

Cyril Stanley Smith in his classic 1948 Institute of Metals Lecture&apos; noted in the case of annealing twins that: "It is well known that approximately the same number of twins per grain and the same general geometry are maintained in large grains as in small ones, a fact that certainly suggests a mechanism such as the above where boundaries and twins grow proportionately together. If this view is correct, a spherical grain should have no mechanism for keeping its twins." Such a constancy of twins per grain would be of particular interest as a tool for the study of the origin of annealing twins. It has been proposed by Mathewson2 and later by Barrett3 that in face-centered cubic metals slip occurs in the {111} Plane by <112> shears which combine in pairs to produce the macroscopically observed <110> movement. This proposal is schematically inviting since <112> slip permits the atoms of one plane to move along the interatomic valleys of its neighbor rather than up the hills across the atom sites. There is strong experimental evidence? that annealing twins result from the growth of nuclei in the form of "twin faults" produced by single <112> shears. It is possible that more twin faults would form at low deformation temperatures than at high ones, because the more rigid lattice may increase the tendency to "freeze" the twin faults and prevent their dissipation in slip. This is suggested by (1) the greater resistance to slip at low temperatures and (2) greater tendency toward mechanical type twinning at low temperatures (a summary of these factors is presented by Seigle and Brick5). These considerations would suggest the following: a. Constant twin/grain ratio for specimens deformed constant amounts at a constant temperature, but annealed at various temperatures above the recrystalliza-tion temperature. b. Larger twin/grain ratios for re-crystallized specimens previously deformed constant amounts at lower temperatures. The purpose of this investigation is to test these considerations. One half in. rods of phosphorous deoxidized electrolytic copper and 70-30 alpha brass (68.36 pct Cu, 0.005 pct Pb, 0.01 pct Fe, 31.62 pct Zn remainder) were rolled to a thickness of 0.350 in., milled to specimens 0.350 X 0.437 X 2 in., and annealed at 700°C to an average grain size of 0.065 and 0.090 mm, respectively. These specimens of each metal were then rolled 50 pct reduction in thickness and annealed one hour under conditions which were all combinations of the following: Rolling Annealing Temperatures. Temperatures. °C °C -70 400 25 500 200 600 Standard metallographic techniques&apos; were used to prepare specimens. By the method of lineal analysis7 using a minimum of ten samples of fifty grains each for each specimen, the ratios of twins* per total grains, twins per twinned grain, twin families* per total grains and twins per twin family were determined. Recrystallized grain size was determined by A.S.T.M. comparison methods.8 All observations were made on the rolling plane. The results are tabulated in Table 1.

Jan 1, 1950

By Lewis A. Smith

Chromium is one of the new metals, but considerable research has been required to determine an approximate record of its production from 1827 until the present. Its use in the form of pure metal is not extensive; of the chromite mined, about one-half is required for making ferrochro-mium, about two-fifths goes directly into refractories and the remainder is absorbed chiefly by the chemical trades. The industries that demand chromite for such uses are mainly in countries that produce little chromite, while the important chromite-producing countries are not generally consumers. The international flow of chromite is therefore of interest, as well as the political and commercial control of resources of the metal. World Production The first mining of chromite is said to have occurred as early as 1820, at the Roros deposit in Norway, but the first well-established output came from the Reed mine, in Maryland, about 1827. Thereafter the United States led the world in chromite production until 1860, when Turkey took the lead. Russia in turn became the chief producer from 1897 to 1902, inclusive. From 1903 to 1909, New Caledonia held first place, but alternated with Southern Rhodesia from 1910 to 1917, inclusive. In 1918, 1919 and 1920-1921, the United States, India and New Caledonia led, respectively. In 1922 Southern Rhodesia assumed first place and has since continued to be the world&apos;s foremost producer by a large margin. Fig. 1 illustrates the production of the leading countries from 1890 to 1929, inclusive. Fig. 2 shows the world production of chromite from 1890 to 1929, inclusive, in long tons and Table 2 gives the production by countries for the same period. The total world production for the 103 years, 1827 to 1929, ipclusive, is estimated at about 7,100,000 long tons, distributed as shown in Table 1. World Trade in 1928 In reviewing the figures for the flow of chromite in 1928, it is evident that the six countries that produce most consume little chromite and

Jan 1, 1931

By James A. Barr

Ferrophosphorus, an alloy of phosphorus and iron or perhaps a physical mixture of definite compounds of iron and phosphorus, has become of increasing importance as the use of the basic open hearth has extended. Commercial ferrophosphorus is a porous, brittle metal of crystalline structure and often with a bluish metallic luster. A typical analysis is as follows: Per Cent. Phosphorus......................................... 18.0 to 22.0 Iron............................................... 80.0 76.0 Oxygen............................................ 0.2 Sulfur............................................. 0.3 Silicon............................................. 0.1 Carbon............................................ 0.1 Manganese......................................... 0.2 This alloy was first made in an electric furnace, as it was thought that the high heat of the electric arc together with the reductive effect of carbon was necessary. About 1898, J. J. Gray, Jr., was producing a high-phosphorus pig iron in his furnace at Rockdale, Tenn., by using high-phosphorus, local iron ores. As these high-phosphorus ores were difficult to obtain, he conceived the idea of increasing the phosphorus content of the pig iron by the addition of phosphate rock to the furnace burden. When the phosphorus content of the pig was kept low and only small amounts of phosphate rock were added to the burden, no particular troubles were encountered, but when he increased the amount of phosphate rock to produce an 18 per cent. alloy, his furnace troubles rapidly multiplied. After many failures and freeze-ups, he was partly successful in making short campaigns on 18 to 20 per cent. "ferro." At first, when the furnace got in bad shape on ferro, he would switch to a regular pig-iron burden until the furnace was brought around sufficiently to start on ferro again. About this time, Mr. Gray applied for and obtained his first patent. As he grew more skilled in the furnace operations, he was enabled to extend the furnace campaigns and produce ferrophosphorus continuously for the life of the furnace lining. These improvements were made the subject of another patent issued in 1916.

Jan 1, 1925

By Paul B. Crawford, Bobby L. Landrum

An electrical model study has been mode to estimate the effect of horizontal fractures in thick reservoirs on pattern conductivity. For the patterns studied, it is shown that fracturing all the wells in the pattern in-creases the conductivity considerably more than frac-turing only single wells. For certain conditions one large fracture is more effective in increasing conductivity than several fractures of half the radius. Particularly in thick reservoirs and with large fractures, the center of the pay is found to be more responsive than other locations. For the seven-spot pattern fracturing the center well is more effective in increasing pattern con-ductivity than fracturing all outer wells with fractures of the same size. INTRODUCTION Although commercial fracturing has been used for some time very little quantitative information is known about the effects of fractures on reservoir behavior. The purpose of this work was to estimate the effect of horizontal fractures in thick reservoirs on the con-ductivity of various flooding and cycling patterns.2 Circular fractures originating at the wellbore and extending symmetrically into the pay were considered. ANALYSIS Electrical models were used in making this study. Horizontal fractures were simulated by copper sheets shaped like the fractures and of a proportional size; the wellbore was simulated by a copper wire. A dilute salt solution served as a conducting medium. A con-ductivity bridge was used to measure the conductivities of the various patterns studied. EFFECT OF HORIZONTAL FRACTURES ON CONDUCTIVITY Fig. 1 shows the effect on square and staggered line drive patterns of horizontal fractures located in the center of the pay. The model is shown schematically in the insert. The upper and lower groups of lines show the effect on conductivity when all wells are fractured, and when only one well is fractured, respectively. In general, pattern conductivities are substantially improved by fracturing all wells. Fig. 2 shows the effect of multiple horizontal frac-tures on the conductivity of a line-drive pattern. At

Jan 1, 1958

By Joseph P. Kittredge

At the time this matter first came up in 1912, the National Malleable and Steel Castings Co. had seven basic-bottom open-hearth furnaces in its plant at Sharon, Fa., using fuel oil, then costing about 2½¢ per gallon. The fuel was satisfactory but its cost fluctuated widely, therefore the use of pulverized coal, which had been successful in the cement industry and in some metallurgical furnaces, was suggested. After a thorough investigation, and an actual trial on one of the open-hearth furnaces, to be certain that the siliceous ash in the coal would not seriously affect the basic slags, the management of this company approved its adoption and a complete installation was started. The conclusion had been reached that the success of the operation depended on pulverizing the coal to greater fineness than had thus far been accomplished, and all equipment was chosen with this in mind. A coal drier was deemed essential, and one of ample capacity and efficiency was selected, which has done all that was expected of it, working without failure and with very small maintenance cost, and is still in operation. Pulverizers The company demanded a pulverization of at least 90 per cent through a 240-mesh screen, with consistent results, and a machine was chosen that was grinding other materials than coal to the fineness desired. The control of fineness was obtained through a vacuum separator, and was governed by the size of the separator and the number and construction of the valves necessary to expand and direct the air current. These valves can be set to obtain any degree of fineness desired, and the result is consistent. Two machines were installed, each capable of producing 2½ tons per hour, 95 per cent through 240 mesh. These machines have done everything expected of them, have run continuously, and are still in operation. An examination of some of our pulverized coal by elutriation and microscopic measurement disclosed that a 1-lb. lump of bituminous coal,

Jan 1, 1940

By Joseph P. Kittredge

At the time this matter first came up in 1912, the National Malleable and Steel Castings Co. had seven basic-bottom open-hearth furnaces in its plant at Sharon, Fa., using fuel oil, then costing about 2½¢ per gallon. The fuel was satisfactory but its cost fluctuated widely, therefore the use of pulverized coal, which had been successful in the cement industry and in some metallurgical furnaces, was suggested. After a thorough investigation, and an actual trial on one of the open-hearth furnaces, to be certain that the siliceous ash in the coal would not seriously affect the basic slags, the management of this company approved its adoption and a complete installation was started. The conclusion had been reached that the success of the operation depended on pulverizing the coal to greater fineness than had thus far been accomplished, and all equipment was chosen with this in mind. A coal drier was deemed essential, and one of ample capacity and efficiency was selected, which has done all that was expected of it, working without failure and with very small maintenance cost, and is still in operation. Pulverizers The company demanded a pulverization of at least 90 per cent through a 240-mesh screen, with consistent results, and a machine was chosen that was grinding other materials than coal to the fineness desired. The control of fineness was obtained through a vacuum separator, and was governed by the size of the separator and the number and construction of the valves necessary to expand and direct the air current. These valves can be set to obtain any degree of fineness desired, and the result is consistent. Two machines were installed, each capable of producing 2½ tons per hour, 95 per cent through 240 mesh. These machines have done everything expected of them, have run continuously, and are still in operation. An examination of some of our pulverized coal by elutriation and microscopic measurement disclosed that a 1-lb. lump of bituminous coal,

Jan 1, 1940

By David Kellogg

THE electrodeposition of iron has been practiced for many years. The earlier work along this line was directed toward the preparation of pure metal, but later applications were the production of thin adherent coatings as used in the electrotyping of Russian banknotes in 1860. A historical discussion of this subject, however, would require too much space and would be inconclusive. Much has been written on the subject but no two experimenters use the same methods or materials. During the war, the British army repair, shops developed a method for building up worn parts of automotive machinery, aero engines, etc., using the cold surfate . bath and low-current density method.1 The method has been used successfully in commercial work for the production of about 6000 repaired parts, therefore experiments along the same lines have since been made at the Westinghouse research laboratory. The apparatus used was essentially that described in the article mentioned except that the current -used for cleaning was obtained from a 3/4-kw. 60-volt, house-lighting generator, direct connected to a 2-h.p., 870-r.p.m., 440-volt, three-phase, type CS induction motor; while the plating current was taken from the storage batteries at any desired voltage. The anodes were made from 0.036 in. (0.9 mm.) Armco iron made into cylinders 5 in. (127 mm.) long by 8 in. (203 mm.) in diameter with a disk of 1/8 -in. (3.17 mm.) micarta, having a 2 in. hole in the center fitted at each end of the cylinder for stirring. The anodes were hung on a wooden rocker frame driven by a wooden connecting I rod directly connected to a small reducing gear, such as is sold by most of the apparatus supply firms; 3 gal. stoneware crocks were used as containers. The solution used by Thomas, 75 gm. of the crystallized ferrous ammonium sulfate per liter, was tried, using the current density recommended by him; namely, 0.33 amp. per sq. dm. Under these conditions the deposit on a ½ -in.

Jan 2, 1922

By R. D. Sprague, C. B. Richardson

The year 1938 was the most active and successful in the history of the oil industry in southern Louisiana. Drilling showed a 16 per cent increase over that of 1937, with a total of 538 wells drilled, which yielded 315 oil wells, 19 distillate wells, 4 gas wells and 200 dry holes. Nineteen new fields were discovered during the year, all but three of which were completed by plugging back to sands previously drilled through. Electrical logging played a large part in these new discoveries as well as in the discovery of new sands during routine development of fields. Fourteen of the new discoveries produced oil, and five produced only gas and distillate. Southern Louisiana continues to lead the world in depth of drilling, which is in part due to the great thickness of comparatively young deltaic sediments near the coast and the existence of numerous sands. The present development is rather evenly distributed over a coastal strip 100 miles wide. New Fields North Tepetate, Acadia Parish.—Atlantic Refining Company&apos;s No. 1 Klumpp was completed for a new field through perforations from 7960 to 7970 ft. The initial production was 65 bbl. of 54" gravity distillate and 4 million cubic feet of gas daily through a 1/4-in. choke with tubing pressure of 2850 Ib. and casing pressure of 2930 lb. The well is in see. 12, 7 S., 2 W., just north of Tepetate field. It was drilled to a depth of 8705 ft. and plugged back for its final completion. Atlantic&apos;s NO. 2 was completed in a new sand at 8300 ft., also for distillate. South Crowley, Acadia Parish.—Humble Oil and Refining Company&apos;s No. 1 Boyd Finch was completed in July 1938, as an oil well through perforations from 7318 to 7325 ft. after plugging from a total depth of 10,222 ft. After making several hundred barrels of oil, the well went to

Jan 1, 1939

By C. B. Richardson, R. D. Sprague

The year 1938 was the most active and successful in the history of the oil industry in southern Louisiana. Drilling showed a 16 per cent increase over that of 1937, with a total of 538 wells drilled, which yielded 315 oil wells, 19 distillate wells, 4 gas wells and 200 dry holes. Nineteen new fields were discovered during the year, all but three of which were completed by plugging back to sands previously drilled through. Electrical logging played a large part in these new discoveries as well as in the discovery of new sands during routine development of fields. Fourteen of the new discoveries produced oil, and five produced only gas and distillate. Southern Louisiana continues to lead the world in depth of drilling, which is in part due to the great thickness of comparatively young deltaic sediments near the coast and the existence of numerous sands. The present development is rather evenly distributed over a coastal strip 100 miles wide. New Fields North Tepetate, Acadia Parish.—Atlantic Refining Company&apos;s No. 1 Klumpp was completed for a new field through perforations from 7960 to 7970 ft. The initial production was 65 bbl. of 54" gravity distillate and 4 million cubic feet of gas daily through a 1/4-in. choke with tubing pressure of 2850 Ib. and casing pressure of 2930 lb. The well is in see. 12, 7 S., 2 W., just north of Tepetate field. It was drilled to a depth of 8705 ft. and plugged back for its final completion. Atlantic&apos;s NO. 2 was completed in a new sand at 8300 ft., also for distillate. South Crowley, Acadia Parish.—Humble Oil and Refining Company&apos;s No. 1 Boyd Finch was completed in July 1938, as an oil well through perforations from 7318 to 7325 ft. after plugging from a total depth of 10,222 ft. After making several hundred barrels of oil, the well went to

Jan 1, 1939

By Sir John Cadnan

The production from the Maidan-i-Naftun area for the year ending March, 1928, was 5,340,000 tons (40,210,200 bbl.), an increase of nearly 13 per cent. over the production for the year ending in March, 1927. The production by years since the opening of the field in 1912 is given in the following table: [1 ton = 7.53 bbl.] Year Ended MaRch Tons Year Ended March Tons 1912 (7 months)............. 43,084 1920....................... 1,354,631 1913........................ 80,800 1921....................... 1,767,070 1914........................ 273,635 1922....................... 2,310,098 1915........................ 375,977 1923....................... 2,913,908 1916........................ 449,394 1924....................... 3,583,000 1917........................ 644,074 1925....................... 4,221,000 1918........................ 897,402 1926....................... 4,470,000 1919....................... 1,081,919 1927....................... 4,735,000 1928....................... 5,340,000 The actual production represents only a small proportion of the potential production. The Masjid-i-Suleiman Field The present producing unit, the Masjid-I-Suleiman, is an anticline some 20 miles in length by 4 miles in width, including three domes as so far proved. The producing horizon is a very open, dolomitic limestone, which, while showing much variation in texture and porosity, is so connected that the lowering of pressure in one section of the field quickly affects the whole area. The topography of the field is broken, with differences of well elevations of as much as 1100 ft. The depths to the limestone vary from about 800 to about 5000 ft. The closed-in pressure of the oil wells varies from zero to 50 lb. per sq. in. and that of the gas wells is about 450 Ib. Although the static oil level may be several hundred feet below ground level, in oil wells of high elevation, all the wells flow continuously or intermittently after they are once started—the majority continuously. The gas is very high in hydrogen sulfide in gas-oil wells of low pressure.

Jan 1, 1929

By O&apos, R. G. Meara, B. W. Gandrud

In 1034 the Southern Experiment Station of the U.S. Bureau of Mines in cooperation with the University of Alabama began an investigation of the availability of kyanite ores as a source of raw material for the manufacture of refractories. Under this program, gravity concentration and flotation tests have been made on two varieties of Georgia kyariite ore, mica schist and massive. Most of the attention has been directed to the mica schist ores, as they are more widely distributed and seem to offer greater economic possibilities than the massive ores. It is evident that the preparation of kyanite for use as a refractory material involves other considerations than the mere elimination of impurities. A concentration process that eliminates impurities satisfactorily may have to be followed by some additional treatment, since the final product must be of a coarseness and size ratio that will make it suitable for refractories manufacture. So far in the investigation at the Southern Experiment Station of the U.S. Bureau of Mines only the concentration problem has been considered, and this paper is intended as an interim report. Ores The schist ores contain 1 to 9 per cent of kyanite crystals embedded in mica and quartz, with smaller amounts of graphite, garnet and iron oxide and minor amounts of other heavy minerals. The garnet content varies from a trace to a few per cent. The minerals are easily released by moderate grinding. The size of the kyanite ranges from fine needles to crystals 1 1/2 in. long, 1/2 in. wide and 1/4 in. thick. The garnet, quartz and iron oxide grains have a smaller mean mesh than most of the kyanite. The maximum size of the garnet crystals corresponds to about 14 mesh and that of the sand grains is considerably less. Most of the iron oxide is still finer, for the minus 100-mesh size is greatly enriched in Fe²O3. The massive ore consists of radiating masses of kyanite interlocked with quartz, mica, corundum and small amounts of other heavy minerals.

Jan 1, 1938

By Henry Howe

THE data which are collected in Table 1 show that the ferrite of low-carbon steel and of electrolytic iron, like the network of hypo- and hyper-eutectoid carbon steel, inherits, either absolutely or relatively, the, grain size of the mother austenite from which it is born in cooling through the transformation range. This evidence is so abundant and concordant as to cast doubt on Prof. Jeffries&apos; 1 hypothesis of what might be called reversed inheritance, by which he explains Mr. Ruder&apos;s observation2 that the grains of a "slug of pressed pure electrolytic iron powder"were coarser, perhaps 12,000,000µ2, after a 3-hr. exposure to 1000° C. than after a 3-hr. exposure to 1250-1300°, when they were about 35,000µs. His explanation implies a natural tendency of coarser austenite grains to yield finer ferrite ones than spring from finer austenite grains, so that the inheritance is reversed. What seems to me a simpler explanation is that the enormous grains which followed the 1000° heating were not inherited from the mother austenite, but were formed by the process of natal coarsening, by means of which Stead and Carpenter developed grains of this same order of magnitude in this same kind of material, on heating to 950° or to 1000°, the temperature used by Mr. Ruder, provided the exposure was not unduly prolonged, and it does not seem to have been in Mr. Ruder&apos;s case. That their coarsening was natal and not through inheritance is shown first by the fact that their grain coarseness was out of all proportion to their relatively low temperatures and short periods, and hence was very far in excess of the grain size which the mother austenite must have had, and second and more cogently by the fact that the shape of these coarsened grains was evidently determined by the temperature distribution which existed during and immediately after the Ar3 transformation, that during the passage from the gamma to the non-gamma state.

Jan 9, 1917

By B. W. Gandrud, R. G. Meara, O&apos

In 1034 the Southern Experiment Station of the U.S. Bureau of Mines in cooperation with the University of Alabama began an investigation of the availability of kyanite ores as a source of raw material for the manufacture of refractories. Under this program, gravity concentration and flotation tests have been made on two varieties of Georgia kyariite ore, mica schist and massive. Most of the attention has been directed to the mica schist ores, as they are more widely distributed and seem to offer greater economic possibilities than the massive ores. It is evident that the preparation of kyanite for use as a refractory material involves other considerations than the mere elimination of impurities. A concentration process that eliminates impurities satisfactorily may have to be followed by some additional treatment, since the final product must be of a coarseness and size ratio that will make it suitable for refractories manufacture. So far in the investigation at the Southern Experiment Station of the U.S. Bureau of Mines only the concentration problem has been considered, and this paper is intended as an interim report. Ores The schist ores contain 1 to 9 per cent of kyanite crystals embedded in mica and quartz, with smaller amounts of graphite, garnet and iron oxide and minor amounts of other heavy minerals. The garnet content varies from a trace to a few per cent. The minerals are easily released by moderate grinding. The size of the kyanite ranges from fine needles to crystals 1 1/2 in. long, 1/2 in. wide and 1/4 in. thick. The garnet, quartz and iron oxide grains have a smaller mean mesh than most of the kyanite. The maximum size of the garnet crystals corresponds to about 14 mesh and that of the sand grains is considerably less. Most of the iron oxide is still finer, for the minus 100-mesh size is greatly enriched in Fe²O3. The massive ore consists of radiating masses of kyanite interlocked with quartz, mica, corundum and small amounts of other heavy minerals.

Jan 1, 1938

By N. A. D. Parlee, I. D. Shah

ThIS reports on some work done on the solubility and diffusivity of oxygen in liquid silver. Solubility. Improvements in the measurement of the solubility of oxygen in silver and silver alloys were made because precise data were needed for calculating diffusion coefficients of oxygen in these alloys. The apparatus and the procedure used were similar to those of Parlee and Sacris.&apos; The main improvement was in the temperature-measuring technique. A MV30 mullite protection tube containing a calibrated Pt-Pt, 10 pct Rh thermocouple was inserted in the reaction tube from the top through a vacuum-tight swage-lok fitting. The protection tube was well-immersed in the liquid silver, which was heated by induction. A Leeds and Northrup potentiometer was used in connection with the thermocouple. All runs reported were made at 1 atm. The temperature range investigated varied from 940" to 1200°C. The ternary point in Ag-0 phase diagram2 is at 939°C and 0.32 wt pct O. This point was reinvestigated using the above-described apparatus and was found to be at 939°C and 0.315 wt pct 0. It was determined by observing the sudden evolution of oxygen at this point while the solubility measurements were made. The point was easily reproducible if precautions were taken to keep the liquid silver essentially in equilibrium with oxygen and to lower the temperature very slowly. All of the results of this new solubility investigation are plotted in Fig. 1. They are represented by the "least-squares" line shown in the figure and the following equation for solubility at po2 = 1 atm: log% O =695.4/T - 1.086 This equation differs only slightly from that proposed by Parlee and sacrisl and that of Diaz, Masson, and Richardson.3 The points of Parlee and Sacris, using an optical pyrometer, are shown for comparison.

Jan 1, 1968