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By R. E. Grace

WHEN an Fe-C alloy, partially reacted to pearl-ite, is quenched rapidly enough to suddenly stop the growth process, it may be expected that any carbon concentration gradients will freeze in situ in the newly formed martensite. A short tempering treatment should precipitate finely dispersed iron carbide particles in approximately the same distribution in which carbon existed during the growth process. Examination with the electron microscope of areas of tempered martensite adjacent to the pearlite-martensite interface should reveal the existence of preferred areas of precipitation. Fig. 1 is an example of an electron micrograph used in this study. Small sections of a commercial steel (0.9 pct C, 1.6 pct Mn, 0.25 pct Si) were austenitized at 945°C in air for 45 min, furnace-cooled to 697°C, iso-thermally reacted to pearlite for 30 min, and quenched into iced brine. Two samples were tempered 5 and 20 min, respectively, in air at 498°C and were subsequently quenched in water. The samples were mechanically polished and etched in 1 pct picral. Collodion replicas were made by the conventional technique; these were shadowed with chromium at about 650' from the horizontal, and electron micrographs were taken at a linear magnification of 14,200X. Prints were made at a magnification of 44,100X. Lineal analysis of the prints was made as a function of the interlamellar spacing and volume percent of iron carbide. Straight lines were drawn into tempered martensite between adjacent lamellae at a distance S/4 or S/6 apart, where S is defined as the interlamellar spacing measured at the pearlite-martensite interface. The volume percent of iron carbide was measured along each of these traces with a "gage length" of S/10. The results of analyses of two samples are shown in Table I. The area included between the S/2 and S distance from the interface could be considered homogeneous within the sensitivity of the measuring technique. This value was used as a proportionality factor to evaluate an arbitrary carbon content of homogeneous austenite, and also as a parameter by which error in measurements could be evaluated. Results of the lineal analysis are shown in Figs. 2 and 3. Fig. 4 shows the statistical nature of the interlamellar spacing for all lamellae measured in this

Jan 1, 1954

By C. H. Frick

BEFORE presenting the main topic, as indicated by the title, this paper will give some of the high-spot history of the anthracite industry. INTRODUCTION The earliest recorded use of anthracite was in the year 1762, when a company of Connecticut pioneers discovered anthracite near the present city of Wilkes-Barre, Pennsylvania. In the early days of the industry, the coal was used only in large pieces, the smaller pieces being discarded to waste piles. As the output of the industry increased during the nineteenth century, uses were developed for somewhat smaller sizes down to and including pea coal. Early in the twentieth century, sizes smaller than pea, such as No. I and No. 2 buckwheat, were developed and used. Accu¬mulations of piles from current mining operations were still increasing; that is, the smaller sizes were still being discarded. It was not until after World War I that preparation of the so-called steam sizes began commercially. This was largely No. 3 buckwheat, which is screened through a 3f 6-in. round-hole screen and over a 332-in. or a M 6-in. round-hole screen. USE OF SMALLER SIZES The earliest recorded experience in burning the small sizes of anthracite for steam production, other than in a previous small experiment, was in the year 1913, when a size known as "boiler fuel" was introduced-a mixture of rice and barley (No. 2 and No. 3 buckwheat). At that time, there was still no method known for burning the still smaller sizes. TABLE I.-Screen Sizes Used for Small Sizes of Anthracite [ ] At the present time, various sizes down to and including No. 5 buckwheat are being prepared (Table I) from current mining and bank-reclaiming operations. The sizes of No. 4 and No. 5 buckwheat and smaller are generally being used for steam production in large boilers. The Pennsylvania Power and Light Co., following World War I, used 680,000 tons of anthracite per year, burned largely on stokers for production of steam, whereas for the year 1943 it used 1,940,000 tons of stoker sizes and 400,000 tons of smaller sizes for grinding and burning in pulverized form. This company has taken a very large part in the research and development of ways and means of burning these smaller sizes of anthracite, including the grinding of the very small sizes. This paper is particularly concerned with the latter feature.

Jan 1, 1946

By Clarence E. Sims, Donald W. Williams, George A. Moore

INTRODUCTION DURING the past several years the steel casting industry has made studies of heavy castings in which the test bar has been taken from heavy sections rather than from attached or separately cast coupons. It has been noted that the ductility properties of these heavy sections are often lower than those normally expected. Such lowered ductility is usually accompanied by a spotty test bar fracture and the ductility values often can be greatly improved if a low-temperature aging treatment is given to the casting. A loss of ductility of this type is considered "abnormal," since it is not accompanied by an increase of tensile strength or hardness, and "temporary" when the ductility can be restored by aging. The cure of such an abnormal condition represents a real improvement in the quality of the steel. Some hundreds of previous investigations, largely qualitative in nature, have established beyond reasonable doubt that hydrogen is normally present in newly manufactured steel and that this gas, in small amounts, can cause a temporary abnormal loss of ductility. The commonly observed association of high gas content and low ductility has given strong evidence for the presumption that hydrogen, possibly assisted by other gases, is the primary cause of the low ductility observed. Accordingly, the Steel Founders' Society of America has, since Nov. 1, 1944, sponsored investigations at Battelle Memorial Institute whose primary objective is to obtain quantitative information on the relation of the amount of hydrogen and nitrogen in steel, together with associated methods of steelmaking and treatment, to low-ductility effects and porosity. In order that these investigations might proceed on a quantitative basis, hydrogen analyses of a precision considerably higher than any previously reported in the literature were necessary. The method chosen was selected entirely on the basis that it had the highest probability of freedom from uncontrollable errors and will be described in detail in another publication.1 Briefly, the method used consisted of the hot extraction of solid pieces of steel weighing from 45 to 50 g in a highly evacuated quartz tube maintained at 1050°C (1925°F) for a period 0f approximately 40 hr. The greatest precautions were taken to eliminate blanks and insure a sound gas-collection system. Vacuum and temperature conditions were maintained automatically. Mercury-filled risers were used for the introduction of samples in order that the system could be degassed thoroughly in

Jan 1, 1948

By R. G. Hall

The method of crushing rock by throwing with violence against a resisting surface does not differ in principle from the method of dropping it from a height upon a specially arranged surface. The former is applied in various mills now in use, especially those for reduction of the softer rocks and minerals. The latter is the principle of the Hadsel mill. Probably not new, but Hadsel's method of approach is at least novel, and applicable to hard as well as softer rocks. An elevating wheel raises the rock and drops it from a height upon a hard metal plate or series of plates. The first commercial mill consisted of two units, one on each overhung end of a horizontal shaft, driven by a gear train from a 100-hp. motor. A description of this machine, then in operation at a mine at Georgetown, Calif., was printed in the Engineering and Mining Journal (April, 1932) and need not be repeated here. That machine took mine run ores of a size up to 12 in. and successfully produced a pulp of approximately minus 60 mesh, suitable for flotation. The mills were each 24 ft. in diameter and 3 ft. in width of face. The elevating buckets were formed of steel plates, extending inward from the periphery, inclined to the tangent at 75". This plate was 3 ft., the width of the wheel, and 2 ft. deep. One side of the wheel towards the carrying shaft was closed, the other was open except for the rim plate enclosing the ends of the buckets and extending inward about 5 ft. on the radius. About a 90" arc was immersed in a concrete tank, which acted as a classifier. The breaking plates, seven in number, were carried cantilever through the open side of the wheel. The lowest plate was placed just above the water level. The illustrations (Figs. 1 and 2) make plain the general arrangements. The plant set-up was somewhat crude but worked well enough to demonstrate the weak points of the mill. In this operation it was early apparent that, considering all the factors from delivery of the mine run ore to the final pulp for flotation or cyanida-tion, the Hadsel mill offered possibilities of great simplicity of plant and savings in power. Continuous operation developed mechanical weaknesses, some of which were readily discernible in details of design, some became more apparent as operations progressed.

Jan 1, 1935

By William Hume-Rothery

I need not say how much I appreciate the honor of being asked to lecture to you, and how much I would thank you for your kind invitation. It is encouraging to feel that the abnormal restrictions of the war have been relaxed sufficiently to allow a free exchange of scientists between our two countries, and I hope that some of you will be persuaded to go to England, and to visit us at Oxford, so that I may repay a little of the kindness I have received from you. As this is one of the first lectures since the end of the war, I have thought it well to describe some of the advances in the theory of alloys which were made in the years between the two World Wars. From the point of view of theoretical science, these years have been of outstanding interest in the history of metallurgy. They have been years in which the development of new and more accurate experimental methods has led to the revelation of many new facts. At the same time, the theoretical developments have enabled some of these facts to be viewed as parts of a general scheme, and we may say that some of the foundations of metallurgical science have been laid. This work was naturally stopped by the Second World War, and it can therefore be reviewed as one definite stage in the development of our Science. To illustrate the new advances, I shall deal mainly with the alloys of copper and silver, since these have revealed many of the general principles. taken, and copper had been found to crystallize in the face-centered cubic structure (Fig I). The determination of this structure naturally was of immense importance in the understanding of our subject, because it meant that we were at last in a position to appreciate the underlying structure out of which the crystal of copper was built. The earlier science of metallography had shown that in many alloys of copper the addition of a second metal in not too large proportions resulted in the formation of a homogeneous alloy which was regarded as a solid solution of the one metal in copper. Fig 2, for example, shows the so-called equilibrium diagram of the copper-zinc alloys, and here you will see that there is a solid solution of zinc in copper extending to something of the order 35 to 40 at. pct zinc.

Jan 1, 1947

By Perry Olcott

The Texas Gulf Coast had a rather disappointing year in the type of oil fields found. Although 17 fields were found in the area during the year, as against 21 fields in 1938, no field of major proportion was discovered. Of these 17 fields, 11 were oil fields, 4 were gas and distillate fields, and 2 were gas fields. The wildcat activity for the year resulted in about 400 dry holes, although probably 10 per cent of these were edge wells close to proved areas. During the year, 1455 oil and gas wells were completed as compared to 1637 wells in 1938, a decrease of 11 per cent. Production for the Texas Gulf Coast area for the year was 100,253,937 bbl. as against 92,530,467 bbl. produced in 1938, an increase of 8 per cent. The daily average production was 274,668 bbl. for the 365 days in the year, although only 252 producing days were allowed by the Railroad Commission; that is, 69 per cent of the time as compared to 71 days shut in for 80 per cent of the time producing in 1938. The total number of oil and distillate wells producing at the end of 1939 was 9495 as against 8451 at the end of 1938—an increase of 1044 wells, or 12 per cent. Of these wells, 6547 were flowing and 2948 pumping, or 69 per cent flowing. At the end of the previous year, 66 per cent of the wells were flowing. New Fields Angleton, Brazoria County.—The Angleton field was discovered by the completion of Glen McCarthy's No. 1 Carr in October for an initial production of 180 bbl. a day of 50" gravity oil through 11/64-in. choke with 3400 Ib. on tubing. The field is 1½ miles east of the town of Anchor. The discovery undoubtedly can be credited to the reflection seismograph, since several companies worked the area from 1933 onward with this type of exploration. The principal block was taken by the Humble company in 1934. Two deep holes by Humble revealed the presence of structure, although both were dry. The Humble and Sun companies later contributed money and acreage toward the drilling of the discovery well. The production is from a sand in the Frio, about 1700 ft.

Jan 1, 1940

By J. W. Pamplin

SOME 40 miles east of Tampa is the center of the Florida pebble phosphate deposits. These are of Pliocene age and consist of several members of the Bone Valley formation.1 Physically the phosphate-bearing formation, locally called the "matrix," varies in thickness from nothing to about 25 ft., usually averages 10 ft., and consists mainly of clay, quartz and phosphate pebble, the latter particles varying in size from colloidal to one inch diameter. The pebbles range in color from dark gray, through white, to cream, and in hardness from that which must be broken with a hammer to that which can be scratched with the fingernail. They assay from 66 to 80 per cent bone phosphate of lime (called "B.P.L." in the trade) or tricalcium phosphate (Ca3P20s) and are secondary apatites. They form from 15 to 50 per cent by weight of the matrix, Over the matrix lies a varying thickness (5 to 40 ft.) of overburden composed of quartz and clay. Briefly, mining consists of two operations2 in open pits about 200 ft. wide by ½ mile or more in length. In the first step overburden is removed by large electric draglines at the rate of about 700 cu. yd. per hour. This is followed closely by hydraulic mining of the matrix at the rate of about 300 cu. yd. per hour. This matrix yardage is moved from an improvised sump in the floor of the pit to an ore-dressing plant, some-times a mile distant, by one, two or three "pit-cars" (pumping units) working on a 12-in. pipe line called the "rock line" to distinguish it from the high-pressure water line feeding the hydraulic giants, which is known as the "hydraulic line." This 700 yd. of overburden and 300 yd. of matrix-1000 yd. in all-amounts to about 1500 tons of material moved each hour. The over-burden is dumped into the mined-out pit alongside the pit being mined. The matrix moves to the ore-dressing plant at the rate of about 7 ½ long tons in 4000 gal. of water per minute-a ratio of approximately one to two by weight, or one to six by volume. This constitutes the "mill heads," but in the pebble district it is called the "matrix."

Jan 1, 1938

Discussion Of the paper Of GEORGE K. BURGESS, J. J. CROWE, H. S. RAWDON, and R. G. WALTENBERG, presented at the Pittsburgh meeting, October, 1914, and printed in Bulletin NO. 93, September, 1914, pp. 2433 to 2437. ALBERT SAUVEUR, Cambridge, Mass.-The Bureau of Standards has once more rendered a public service. Manufacturers and consumers of steel rails alike should welcome the important results now laid before them by Dr. Burgess and his co-workers. Personally I welcome these results all the more that in the past, as some of the members may remember, I have never failed, when occasion arose, to call the attention of rail manufacturers to the importance of controlling the finishing temperature of their rails. The pioneer metallographic work conducted at the South Works of the Illinois Steel Co. in the early '90s clearly brought out the close relations existing between the manufacture, properties, and finishing temperature of steel rails and I think I may say without fear of being contradicted that the results of this early investigation had considerable influence in drawing the attention of manufacturers to the importance of finishing temperature and of heat treatment in general. The results now before us should be, it seems to me, of great value to rail manufacturers in their efforts to produce a better and, above all, a safer rail. It seems to be demonstrated in this paper that there is no difficulty in installing in rail mills pyrometric outfits, making it possible to record, without interfering in the mill operations, the temperature of the rails as they leave the finishing rolls, or of the ingots as they are withdrawn from the soaking pits, or at any intermediate point in the process of manufacture. The inference necessarily follows that manufacturers should see to it that such pyrometric installations are made, in order that they may actually know, instead of guessing at, the temperature of their rails;, and, for obvious reasons, rail consumers should insist upon such installations. In the rolling of heavy sections especially should care be taken to lower their finishing temperature, seeing that the present tendency is to finish them at higher temperature, whereas, owing to the slower cooling of their central portions, which makes for coarseness of grain, they should, if anything, be finished at a lower temperature than lighter rails.

Jan 4, 1915

By Douglas Bunting

The first presentation of this paper was before the Pennsylvania Anthracite Section of the Institute in May, 1914, after which a committee was selected to verify and add to the data contained in the original paper. The author is indebted to G. W. Engel, H. W. Montz, R. V. Norris, and H. H. Otto, of this committee, for additional information. I also wish to acknowledge the courtesies extended me by a number of the mining companies for information relative to accidents, systems of proving, and mining practice. The paper should bring forth additional criticism and recommendations, which will be of value in establishing future practice relative to this subject. Throughout the northern anthracite field of Pennsylvania glacial and alluvial deposits of sand, gravel, and clay exist in varying quantities. The extent and character of these deposits are described by N. H. Darton in Bulletin 45 of the U. S. Bureau of Mines, entitled, Sand Available for Filling Mine Workings in the Northern Anthracite Basin. It is estimated that the northern field contains 10,000,000,000 tons of these deposits, or a sufficient quantity to cover the 176 square miles of coal area to a depth of about 25 ft. These deposits of sand, etc., overlying the coal measures attain a maximum depth of more than 300 ft., and are usually saturated to such an extent as to render them semi-fluid. It is this condition of fluidity that limits the mining of coal seams cropping in or in close proximity to these deposits, and it is in the interest of safer and more efficient mining methods under these conditions that the study of this subject is here presented. The dangers incident to mining under deep deposits of sand have long been realized by operators in the northern anthracite field, and much has been done to prove the contour of the rock and vein croppings underlying these water-bearing deposits of sand and gravel. Although, as far as we know, the accidents of the past have not resulted in the loss of life, except in two instances, yet nearly all were attended with more or less miraculous escapes of men working in the vicinity of the breaks, so that the probable existence of pot holes, etc.,

Jan 1, 1915

By Howard Eavenson

IN a discussion in the Transactions of the Institute (vol. xl, page 835 et seq.) the writer gave some data about explosions of gas and dust in the coal mines of the United States, Canada, and Mexico, among which was a table showing all explosions in the countries mentioned, in which five or more fatalities occurred, which had been reported from Oct. 2, 1871, to Mar. 28, 1908. Since the elate of publication of this discussion, as opportunity has presented, this table has been extended by the addition of explosions occurring since that time, by including several explosions which occurred prior to 1871, and by adding to the list all explosions of gas or dust in which less than five fatalities occurred, this having been clone by a search through the various mine reports, partly by the author and partly by the Librarian of the United Engineering Societies, and by correspondence with mine inspection officials of the various States, to all of whom acknowledgment of assistance is made: The accompanying list, Table I, is believed to be as nearly correct as such a statement can be made, and, while only extending to Dec. 20, 1912, contains all of the data now available in reports. An analysis of Table I, showing the occurrence by months of all explosions of gas or dust, including those in which less as well as those in which more than five fatalities occurred, is given in Table II. The statements are frequently made, in technical papers, that the number of explosions, particularly the larger ones, is increasing more rapidly than the number of mines, and that the number of fatalities caused by mine explosions, particularly the more disastrous ones, is increasing faster, proportionally, than the production of coal. To investigate the correctness of these statements Table III was prepared, in which the number of accidents and fatalities is taken from Table I, using those in the United States only, on account of lack of other data for Canada and Mexico; the production is from the records of the U.S. Geological Survey, and the number of mines, from 1909, from the same source; previous to 1909, the number of mines, for the decennial periods, was furnished by the U. S. Census records.

Jan 10, 1914

By Richard E. Gray, Robert W. Bruhn

INTRODUCTION Underground coal mining has been practiced in the United States for over 200 years. Much early mining was not as efficient as today and unrecovered coal pillars, often of variable size and spacing, remain to support the overlying strata after mining has ceased. These conditions constitute a potential subsidence problem since the pillars will eventually fail. The possibility of subsidence above an abandoned mine must be anticipated except when total extraction has been achieved, permitting subsidence concurrent with mining, or when large coal pillars which have an adequate factor of safety for long-term support remain in the mine. Coal is found in 37 states and is mined under- ground in 22 states (HRB Singer, Inc. 1980). Underground mining produced 40 percent of the 7.5 x 10" kg. (830,000,000 tons) of U. S. coal mined in 1980 (Keystone, 1981). There are approximately 70,000 abandoned or inactive underground coal mines in the United States and approximately 2000 active underground coal mines (NAS/NAE, 1975). Underground coal mining in the United States will eventually extend beneath 162 x 109m2 (40 million acres). Thirty-two billion m2 (8 million acres) are already undermined (HRB Singer, 1977). Subsidence due to mining has affected more than 8 x 109m2 (2 million acres) in 30 states, 7.7 x 10 9m2 (1.9 million acres) from bituminous coal mining and the remainder primarily from anthracite mining (Johnson and Hiller, 1979). Subsidence from all other underground mines has affected only 6.9 x 10 7m2 (17,000 acres) (Johnson and Miller, 1979). Subsidence appears to be particularly widespread in two states, Pennsylvania and Illinois, although serious subsidence is found in localized areas of numerous other states (HRB- Singer, Inc. 1977). Host underground coal mining has occurred in the eastern United States, where the U. S. Bureau of Mines estimates 1.6 x 109m2 (398,600 acres) in urban areas in 18 states (out of a total 1.7 x 109m2 1418,400 acres]) are threatened by subsidence (Table 1) (Johnson and Miller, 1979). Pennsylvania and West Virginia contain 57 percent of the total threatened areas. The U.S. Bureau of Mines estimates the total cost to stabilize the estimated 1.7 x lo9m2 (418,400 acres) of subsidence prone urban land located above abandoned mines is over $12.5 billion (1978 costs).

Jan 1, 1982

By T. R. Ingraham

The thermal decomposition of Nickel (II) sulfate was examined by determining the total pressure of SO3, SO2, and O2 developed over a sample when it was heated in an evacuated system fitted with a Pyrex-bellows mercury manometer. The equilibrium pressures were combined with previously published data on the oxide and sulfide systems to establish a predominance volume diagranz for the Ni-S-O system over the normal range of roasting temperature and gas composition used in the treatment of nickel minerals. The protective coating of NiSO4 which forms on the oxide during roasting, and which causes the sulfation reaction to almost cease, may be disengaged readily by temperature cycling. When nickel sulfide is roasted in air, either nickel sulfate or nickel oxide may be formed by altering the experimental conditions. Because of the differing stabilities of various metal sulfates and oxides, possibilities exist for separating some of the constituents of mineral mixtures by the technique of selective sulfation and oxidation roasting. The experimental information required to make reasonable predictions of the possibilities of this technique is incomplete and, in some cases, in conflict. As part of a general program of providing some of the required information, a study EXPERIMENTAL Materials, Apparatus, and Procedure. Fisher certified reagent-grade nickel sulfate hexahydrate was used in the source material for both nickel sulfate and nickel oxide. The manufacturer's analysis indicated: 0.001 pct Fe, 0.005 pct Cu, 0.003 pct Co, 0.001 pct C1, other 0.15 pct. The material was dehydrated, before use, by heating at 420°C for 5 hr. The oxide was prepared by roasting the sulfate in an open tray in a muffle furnace for 10 hr at950°C. Preliminary experiments which were made by thermogravimetric analysis and X-ray diffraction analysis established that only one crystal form of nickel oxide and one of nickel sulfate were involved in the equilibrium shown in Eq. [I]. The apparatus used in the experiments was described previously.1,3 The main feature is a corrosion-resistant, flexible Pyrex bellows which is used to protect the mercury in a u-tube manometer. Before a run was begun, the apparatus was washed thoroughly with hot hydrochloric acid and water and dried by heating under vacuum. The mixed sulfate -oxide sample was contained in a platinum bat rest-

Jan 1, 1967

By R. D. Carpenter, R. R. Smith, E. E. Sougstad

The Union Carbide Nuclear Co., in treating low-grade tungsten and molybdenum-copper ores at the Pine Creek mill near Bishop, Calif., employs flotation for the initial extraction of the valuable mineral constituents. These minerals consist of scheelite, molybdenite, and several of the copper sulfides including chalcopyrite, bornite, and possibly chalcocite. They occur in a gangue consisting mainly of quartz, pyroxene, and garnet, with minor amounts of calcite and fluorite. This assemblage of minerals presents an unusual grinding problem because of their diverse physical characteristics. Feed for the circuit is prepared from -6-in. ore by screening on a rod deck screen to remove the -5/8 - in. material. The+5/8-in. fraction is crushed in open circuit with a 5 1/2-ft Symons shorthead crusher operated at a closed- side setting of approximately 3/8 in. The screen undersize and the conecrusher discharge are then combined to constitute the feed to the grinding circuit. The ore is ground in 6 by 5-ft grate discharge ball mills operated in closed circuit with 54-in. spiral classifiers. The classifier overflow goes directly to the flotation section of the plant. Studies of the flotation circuits have shown that the maximum recovery of the tungsten, molybdenum, and copper minerals occurs in the particle-sue range from 100 mesh to approximately 400 mesh. Further, practice has shown that the ore minerals are effectively reduced to this size range if the total ore is ground to -48 mesh. The general objective of the grinding circuit then is to reduce the feed to -48 mesh with the production of a minimum amount of slimes. The purpose of this investigation was to ascertain how well the grinding circuit was carrying out this objective. The method used was to sample the various products in the grinding section, size these products accurately, and analyze the resulting sue distributions. SAMPLE PREPARATION AND SIZING TECHNIQUES Representative samples of the different grinding circuit products, namely, new feed to the ball mill, ball mill discharge, classifier overflow, and classifier return, were taken over a 24-hr period. The samples were then dried without being filtered to avoid loss of slimes. The percentages of solids in the different products were the following: new ball-mill feed, 95%; ball mill discharge, 82%; classifier overflow, 52%; and classifier return, 84% solids. Particle sizes were determined on each product by wet-screening an

Jan 1, 1962

By George Comstock

THESE notes should be considered as a further discussion of Mr. S. W. Miller's paper on "Some Structures in Steel Fusion Welds."1 In that paper and the resulting discussion; several conflicting opinions were expressed as to the identity of the needles or small crystals present in oxyacetylene and electric welds. An interesting specimen of an electric weld examined by the writer seems to afford new evidence regarding the needles, that may be worth presenting at this time. The specimen consisted of a piece of 3/8 -in. (9.5-mm.) steel plate, about 2 ¼ by 1 ¾ in. (5.7 by 4.5 cm.) in size, on which some iron had been deposited by arc-welding to a depth of about 3/16 in. (4.7 mm.). The weld appeared to be a very good one, with no noticeable break or boundary between the plate and the deposited iron. Sections for microscopic examination were cut in three planes at right angles to each other through the deposited metal, two of these sections also extending through the steel base. After polishing in the usual way, the microscope showed that the deposited metal was full of very small, round, gray, oxide spots, while the steel plate contained quite a number of alumina inclusions and a little slag. The sections were etched with nitric acid in ordinary alcohol, and structures similar to those described by Mr. Miller were then displayed. Fig. 1 shows the typical appearance of the deposited metal, which was quite uniform in all three of the planes examined. The round dark-gray spots are oxides, and the pale angular crystals are evidently what, Mr. Miller called cementite and Mr. Jeffries martensite, while Dr. Ruder and Prof. Boylston suggested that they might be nitride. The boundary between this structure and the steel base was sharply defined under the microscope. Fig. 2 shows the structure of the plate just under the weld; this is almost martensitic and resembles a coarse-grained casting. This coarse cast structure soon became finer as the distance from the weld increased, finally, becoming very fine, with numerous small particles of sorbite. This very fine structure, shown in Fig. 3, became coarser in turn and finally merged into the original normal structure of the steel plate, shown

Jan 1, 1919

By John Chipman

IN the manufacture of rimming steel-which constitutes the bulk of the world's production of mild steel-the factors that determine the quality of the product are very closely associated with the pouring and solidification of the steel in the ingot mold. The behavior of the metal in the mold is, on the one hand, the resultant of the complex processes through which the liquid metal has passed, and, on the other, it is an indication of the probable behavior of the ingot during subsequent rolling and finishing operations. A detailed study of the mechanism of solidifica-tion of rimming ingots ought to yield information that will be useful iii controlling the structure of the ingot and the quality of the ultimate product. The first step in such a study is to determine the rate at which the solid shell of the ingot forms from the liquid metal in the mold. A knowledge of the rate of solidification will make it possible to correlate observations of rimming action and gas evolution with studies of the structure, porosity, and segregation within the ingot. Thus, if the metal in the mold is observed to fall, let us say one minute after pouring, the portion of the structure associated with this fall will lie located in the part of the ingot that is solidifying at that particular moment. The knowledge that the ingot solidifies to a depth of ¾ in. in the first minute enables us to locate in the ingot the structural features corresponding to the observed fall. Conversely, the blowhole condition associated with a rising or growing ingot will be found in the part of the ingot that is solidifying at the time the phenomenon occurs. The time required for the complete solidification of an ingot has long been a matter of speculation. It is of considerable practical importance in connection with charging and heating in the soaking pits. Too rapid handling of the ingot through heating and rolling processes has been thought to be associated-with excessive pipe and other internal defects, so that a further object of this investigation was to attempt to estimate the time required for complete solidification, or the minimum practical time within which the subsequent operations can be performed with the minimum hazard.

Jan 1, 1937

By J. Spotts McDowell

Preparation and Use.-Magnesite is an important refractory in open-hearth, heating, and electric furnaces for steel-making and in many of those employed in the metallurgy of copper and lead. It is sold in the form of brick, finely ground "furnace magnesite" for brick-laying, and dead-burned grains for making and repairing furnace bottoms. The latter are a mixture of granules varying in size from pieces of about 5/8 in. (1.6 cm.) diameter to very fine but sandy particles. Dead-burned magnesite results from calcining 'the crude or lightly burned mineral at a temperature that will not merely drive off practically all the CO2, but will also cause sintering of the particles. During this process the pieces shrink considerably and become hard, dense, and inert to atmospheric moisture and C02; under-burned material, on the other hand, will hydrate on exposure to the air. A small percentage of ferric oxide seems to be necessary for the production of a satisfactory sinter; from 4.5 to 8 per cent. in the dead-burned grains is considered the most desirable amount. Dolomite has been little used for brick-making in the United States, but it is prepared for use in the granular condition calcined or "double-burned" and is the principal ingredient of several materials offered for sale under various trade names for refractory purposes. Dolomitic refractories are almost wholly confined to the open-hearth and electric furnaces, where they are used for fettling and as substitutes for magnesite. Much more magnesite and dolomite are used for basic open-hearth steel-making than for all other refractory purposes. The hearth of the furnace is usually built up of magnesite brick and dead-burned grain magnesite so laid that the brick base is protected by a working bottom of the granular material. The latter is sintered into place in layers 1 to 2 in. (2.5 to 5 cm.) thick to a total depth of 12 to 18 in. (30 to 45 cm.) at the center of the furnace. After each heat, burned dolomite is thrown against the banks as high as it will stick, and all holes in the bottom are filled. At most plants such holes, at the end of each week, are also

Jan 2, 1919

By J. W. Pamplin

Some 40 miles east of Tampa is the center of the Florida pebble phosphate deposits. These are of Pliocene age and consist of several members of the Bone Valley formation.1 Physically the phosphate-bearing formation, locally called the "matrix," varies in thickness from nothing to about 25 ft., usually averages 10 ft., and consists mainly of clay, quartz and phosphate pebble, the latter particles varying in size from colloidal to one inch diameter. The pebbles range in color from dark gray, through white, to cream, and in hardness from that which must be broken with a hammer to that which can be scratched with the fingernail. They assay from 66 to 80 per cent bone phosphate of lime (called "B.P.L." in the trade) or tricalcium phosphate (Ca³P2O8) and are secondary apatites. They form from 15 to 50 per cent by weight of the matrix. Over the matrix lies a varying thickness (5 to 40 ft.) of overburden composed of quartz and clay. Briefly, mining consists of two operations2 in open pits about 200 ft. wide by 1/2 mile or more in length. In the first step overburden is removed by large electric draglines at the rate of about 700 cu. yd. per hour. This is followed closely by hydraulic mining of the matrix at the rate of about 300 cu. yd. per hour. This matrix yardage is moved from an improvised sump in the floor of the pit to an ore-dressing plant, sometimes a mile distant, by one, two or three "pit-cars" (pumping units) working on a 12-in. pipe line called the "rock line" to distinguish it from the high-pressure water line feeding the hydraulic giants, which is known as the "hydraulic line." This 700 yd. of overburden and 300 yd. of matrix—1000 yd. in all-amounts to about 1500 tons of material moved each hour. The overburden is dumped into the mined-out pit alongside the pit being mined. The matrix moves to the ore-dressing plant at the rate of about 7% long tons in 4000 gal. of water per minute—a ratio of approximately one to two by weight, or one to six by volume. This constitutes the "mill heads," but in the pebble district it is called the "matrix."

Jan 1, 1938

By Charles H. Jacoby, Stanley J. LeFond

Salt, or halite, has had a long and most varied history. While we know the Chinese were producing salt as early as 3000 B.C., the first written reference to salt appears in the book of Job recorded about 2250 B.C., which reads: "Can nothing which is unsavory be eaten without salt?" We do not know when man first used or realized that he needed salt. Perhaps one of our early Eolithic ancestors noted a piece of salt shining along the shore of a salt lake or a bay. He may have been attracted by its shining luster and tasted it. He found that it had a tart, zesty, distinctive flavor and perhaps chewing a large piece, became ill. However, by experimentation he soon learned that taken in the proper proportions salt was a very pleasant additive to his food. He may have noticed, too, that it increased his feeling of well-being. Thus, when the curtain of recorded history rises, we find man using salt for his health's sake, seasoning his food with it, feeding it to his animals, and using it in many superstitious and religious ceremonies. These uses, however, are a far cry from the uses for salt today. The first attempts of white man to make salt in continental United States were reported in the year 1614, while the first commercial production was in 1790. Practically all of this salt was for human consumption, preserving food, and curing hides. Prior to the coming of the "white man" Indians are known to have made salt from such salt springs as those that existed at Avery Island, La., Charleston, W.Va., Saltville, Va., and Onondaga, N.Y. In 1862 the first rock salt was mined in the U.S. from the Avery Island salt dome. Almost immediately thereafter, rock salt was mined at Saltville, Va., and then at Retsof, N.Y., in 1885. Of the more than 40 million tons produced annually in the United States, about 68% or 27 million tons are consumed by the chemical industry in the production of chlorine, caustic soda, hydrochloric acid, sodium metal, and some 30 other basic chemicals. These basic chemicals, in turn, are used in the preparation of approximately 14,000 chemicals which range from soap and detergents, to such chlorinated hydrocarbons as DDT, BHC, carbon tetrachloride, etc.

Jan 1, 1975

By J. W. Pamplin

SOME 40 miles east of Tampa is the center of the Florida pebble phosphate deposits. These are of Pliocene age and consist of several members of the Bone Valley formation.' Physically the phosphate-bearing formation, locally called the “matrix," varies in thickness from nothing to about 25 ft., usually averages 10 ft., and consists mainly of clay, quartz and phosphate pebble, the latter particles varying in size from colloidal to one inch diameter. The pebbles range in color from dark gray, through white, to cream, and in hardness from that which must be broken with a hammer to that which can be scratched with the fingernail. They assay from 66 to 80 per cent bone phosphate of lime (called "B.P.L." in the trade) or tricalcium phosphate (Ca3P2O8) and are secondary apatites. They form from 15 to 50 per cent by weight of the matrix. Over the matrix lies a varying thickness (5 to 40 ft.) of overburden composed of quartz and clay. Briefly, mining consists of two operations2 in open pits about 200 ft. wide by ½ mile or more in length. In the first step overburden is removed by large electric draglines at the rate of about 700 cu. yd. per hour. This is followed closely by hydraulic mining of the matrix at the rate of about 300 cu. yd. per hour. This matrix yardage is moved from an improvised sump in the floor of the pit to an ore-dressing plant, some- times a mile distant, by one, two or three "pit-cars" (pumping units) working on a 12-in. pipe line called the "rock line" to distinguish it from the high-pressure water line feeding the hydraulic giants, which is known as the “hydraulic line." This 700 yd. of overburden and 300 yd. of matrix-1000 yd. in all- amounts to about 1500 tons of material moved each hour. The over- burden is dumped into the mined-out pit alongside the pit being mined. The matrix moves to the ore-dressing plant at the rate of about 7 ½ long tons in 4000 gal. of water per minute-a ratio of approximately one to two by weight, or one to six by volume. This constitutes the "mill heads," but in the pebble district it is called the "matrix."

Jan 1, 1938

By L. F. Bishop

The ore bodies of the Butte district1 are found in many different vein systems having many different structural characteristics; some are narrow with self-supporting ore but with weak walls; some are wide with weak or strong ore and weak walls; some of the veins dip steeply and others do not; some have well defined walls, and others, in the "horsetail" area, are mined to sample limits. Consequently, many types of stoping methods are used in mining these ores. The stoping methods used in the Anaconda Copper Mining Company's properties are the square-set and cut-and-fill. The square-set method is the principal stoping method used because, generally speaking, the ore and walls are weak, the ore is high grade, and good extraction is required. These square-set stopes may be classified as conventional, vertical slice, slot, timber rill, and Mitchell slice types. The conventional type square sets are those of various shapes and sizes, mined by horizontal slicing to the boundaries of the stopes. The rock is slid to the chute and the filling is moved by gravity. The ore and waste may be handled with a scraper when local conditions require it. The vertical-slice square-set system is the type used in narrow veins that are mined in vertical panels. The slot stopes are stopes in wide veins that employ a slot across the vein instead of a con- ventional chute to handle the ore. The timber rill is a type of square set in which the ore is mined as in inclined cut and fill but the square-set timber is used. The Mitchell slice is the type used for some broken or heavy ore blocks, or for pillar extraction between two converging slot stopes. The cut-and-fill method is used where the ore will stand without much support although the walls may be weak. The principal type of this method is the horizontal cut-and-fill, with mechanical scraping. The inclined cut-and-fill, or rill, stopes are used in some veins too narrow for scraping. The production from the various methods in use for the second six months of 1944 are shown in Table I. This paper deals mainly with the vertical slice and slot types of square-set stoping, because of their effect on production locally and their potential possibilities here and in other districts, with sufficient reference to conventional square-setting for comparison.

Jan 1, 1946