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By Galen Clevenger

TELLURIUM has had the rare and unpleasant distinction of having fewer uses than any of the other common elements; indeed, it has had no regular or important uses. It is not only a useless and disagreeable element, but to the metallurgist it has been a nightmare. The interference of tellurium with the recovery of gold and silver from certain ores, as well as the difficulties it occasions in the electrolytic refining of copper, lead, and silver, are too well known to require enumeration here. In fact, the potential producers of the element have been quite willing to throw the crude products from which it might be recovered into the sewer. First isolated from a gold ore in 1782, although remaining a practically unknown element to the general public, it has long been a guiding star to the Western prospector in his quest for the "golden fleece." On the burning sands of the desert, in the frigid North, and in the mountain fastnesses of the more temperate regions of this and other continents, he has endured all manner of hardships in his search for tellurides; for where he found indications of tellurium, there he expected also to find rich deposits of gold and silver. This peculiar superstition, that tellurium, associates itself in nature with the precious metals, is not with-out reason, for witness the fabulously rich gold tellu-ride deposits of Cripple Creek and Western Australia. Selenium, the sister element of tellurium, which it resembles in many of its properties, was first isolated in 1817. It is found in small amounts in certain native sulfur deposits, and in many copper and silver ores from all parts of the world. Until a few years ago, selenium, like tellurium, was practically useless. The, unusual property of having its electrical conductivity influenced by light has long been known. Selenium cells, which for many years have been a scientific curios-ity, in the future may come to be of more commercial importance, but this use could scarcely absorb any large amount of the element.

Jan 1, 1923

By John B. Miles

The present period of depression in the iron industry, with the resultant close approximation of the cost of production to the selling-price of pig-iron, should make the discussion of this subject at this time not only interesting but profitable. It is possible that it may result in the collection of data, which when added to those set forth in this paper will clearly indicate the influence of the rate of driving a blast-furnace upon its fuel-economy. The question is not whether an excessive rate of driving (meaning a rate definitely in excess of the average rate in the United States) results in an increase in the fuel-consumption, but rather whether a rate considerably less than the average does not result in a marked decrease in the fuel-consumption. The commercial and metallurgical sides of this problem must be kept separate. With a large margin between the cost of production and the selling-price of pig-iron, and a plant already in operation, it is frequently profitable to sacrifice a couple of hundred pounds of fuel per ton of pig-iron in order to secure the greater gross profits resulting from a larger tonnage. When this margin decreases with a plant in operation, or when the determination of the size of a new furnace intended for a certain production confronts the management, a clear understanding of the effect of the rate of driving upon fuel-economy becomes necessary. If it be granted that slow driving leads to low fuel-consumption, it should be borne in mind in the design of a new plant for the production of a certain tonnage that the adoption of an economical rate of driving increases only the size of the furnace proper, whereas, due to the decrease in quantity of blast required on account of the decrease in fuel burned, the stoves, the blowing-engines, the boilers and the piping may be de-

Jan 1, 1909

By O. G. Bell

Development and exploratory work continued vigorously during 1933 in Southwest Texas and, from a standpoint of new production discovered, the new discoveries probably equal or may even exceed any previous year in the history of this area. While exploratory work has been general over the area much of it has been confined to two localities; namely, northern Duval-southern McMullen counties and Starr County. A persistent effort has been made to open new productive areas in northern Duval County and the adjacent part of McMullen County, occasioned primarily by the fact that the Government Wells group of fields in northern Duval County have been by far the best fields so far developed in this part of the state, and the possibility of being able to find new fields in the same general area under similar geologic conditions. This has resulted in the discovery of three new fields, Northwest Government Wells, Moca and Eagle Hill. In addition encouraging shows have been found in North Kohler, Hoffman, Smith-Hunter and the Ignacio areas with the prospect that continued exploratory work will develop some of these prospects into fields. Wildcatting in Starr County was the continued result of an effort to find a trend of production through the county southwestward from the older Reynosa Escarpment fields in Webb, Jim Hogg and Zapata counties. This same impulse has caused most of the wildcatting in Zapata County. New Fields Cuevitas.—During the early part of the year the Sun Oil Co. began some core-drill work on a block of leases in northern Starr County, then in June drilled a test which found production in the McElroy member of the Jackson at 2200 ft. Drilling in this field has continued but the productive area has not as yet been clearly defined. Seven oil wells and three gas wells have been completed. There is as yet no pipe line outlet for this production, hence no productive history of the wells has been

Jan 1, 1934

By Percival Roberts

It may seem necessary to offer an apology for presenting for consideration a process which is conspicuous by its absence in the literature of the Institute, and which may be thought by some to belong to the past in metallurgy, and to have been already superseded. But the large capital invested in puddling calls for a careful consideration of the question whether the time has certainly arrived when the puddling furnace must be replaced by the converter and open-hearth furnace. May there not still be a place for puddled iron alongside of molten iron and steel, and is not the improvement of the puddling process itself worthy the attention of engineers equally with the Bessemer and open-hearth processes? The changes involved in the conversion of pig iron into wrought iron are well understood and need only be briefly alluded to. The patent of Henry Cort bears the date of 1784. Since that time * the improvements in the process have mainly consisted in the replacement of sand by iron bottoms by Samuel Baldwyn Rogers in 1818, and the still more recent substitution of iron oxide for the refractory materials used for the rides and bridge of the hearth, which distinguishes the wet or boiling process from the dry or pnddling process. Chemically, the process consists in the removal of the metalloids from the pig iron, a result effected mainly by the iron oxide. Silicon is first oxidized, then the phosphorus, and finally the carbon. The silicic and phosphoric acids produced pass into the cinder and the carbonic oxide burns as it escapes from the bath of metal. It is interesting in this connection to note the effect of temperature on the removal of the phosphorus from the iron. As is well known, no phosphorus is eliminated under the oxidizing influences prevailing in the Bessemer converter, while from 70 to 80 per cent. is removed in puddling. But we find, if in working cold short irons

Jan 1, 1880

By Kefton H. Teague

Olivine is a mineral containing a mixture of forsterite (Mg,SiO,) and fayalite (Fe,SiO,) in solid solution. The name "olivine" was first applied by Werner in 1790 (Hunter, 1941) because of the olive-green color of the mineral. Uses The principal use of olivine is in various foundry applications. Perhaps the largest tonnage is utilized as a special foundry sand for mold-making in brass, aluminum, and magnesium foundries (Schaller, 1957, 1958; Snyder, 1957). Olivine was first used as an industrial mineral in the early 1930s for a refractory material (Anon., 1970; Hunter, 1941). As a refractory raw material, olivine was first introduced in the United States as handcobbed, selected, shaped blocks of crude olivine. This use met with limited success. More recently, finely ground olivine blended with MgO and pressed into bricks, which are then fired, have found use in glass tank furnaces and open-hearth furnaces. Ramming or gunning mixes for basic furnace linings also utilize olivine. Olivine has been used in ladle linings with varied success. In Europe, substantial tonnages of olivine are utilized in refractory brick for night storage heaters (Anon., 1970). A limited amount of olivine has been used in the past as a fertilizer (magnesium source) and has been fused with rock phosphate to produce a magnesium phosphate as a plant food. The relatively high magnesia (MgO) content of olivine also attracted attention to it as a potential source of both magnesium compounds and as a source of metallic magnesium (Bengston, 1956; Hunter, 1941). Industries using olivine have not standardized specifications. Normally, producers supply a product which meets individual customer requirements. Such requirements involve type and amount of impurities, grain size of product, refractory characteristics, etc. Geology Olivine occurs commonly as accessories in basic igneous and basic metamorphic rocks. Economic deposits are of magmatic origin and are restricted to essentially the dunite variety of periodotite. Dunites are medium to coarse- grained crystalline rocks, generally reddish brown on weathered outcrop, and are composed primarily of olivine.

Jan 1, 1975

By F. E. Marcy

It is not my purpose to write a lengthy article or to attempt the solution of the problem I am presenting, but to call attention to what I believe an important issue, hoping that it may arouse in some one, or in some company, a sense of responsibility. I am referring particularly to modern up-to-date milling plants discharging tailings which under existing conditions cannot be further treated or from which a recovery cannot be made at a profit, but where a quantitative and qualitative analysis shows that a portion of these tailings might profitably be treated at some future time when our metallurgical knowledge has advanced, when we have cheaper power, chemicals, labor or other advantageous conditions. Early Methods In our early milling methods very little attention was given to the care of tailings. It was thought that if the mill could be located by a stream or a river where the final disposal of the tailings could be forever lost sight of, it was highly desirable. We can now see and estimate what was lost in these rich tailings, not recoverable from the streams and the great rivers. Fortunately for many, the depositing of tailings in these streams and rivers was objected to by the farmers and finally prohibited by the State Legislatures. In looking back over our past milling methods, it seems that the first endeavor was to make the maximum amount of money for a particular period, or year, regardless of the future, the depletion of the mine, or the conservation of the total mineral content. The dumping and sluicing of 3000 tons per day, for a period of many years, of a copper tailing carrying 12 lb. of recoverable copper together with its portion of precious metals, the thousands of tons of valuable metal contained in the tailings that have been carried away by Silver Bow Creek, the streams of the Coeur d'Alene, the Flat River and Colorado districts, the enormous quantities

Jan 1, 1918

By Clemens Jones

The peculiar property of the lode-stone has been known for many centuries. It was very early observed that, by contact with it, iron is magnetized; and it is needless to dwell here upon the history allied to this discovery, now constituting the wonderful science of magnetism. Deposits of magnetic ore are distributed all over the world. The ore is sometimes very rich and comparatively free from gangue, sometimes sparsely disseminated in the native rock. These bodies of lean ore often cover large areas. Modern ingenuity has utilized the principle of the magrtet, originally derived from the ore itself, and really belonging to the time of Plato and Theophrastus, to separate the magnetic ore-particles from the rock which has previously been crushed. This process recovers the ore in the state of a concentrate very rich in iron and, in some cases, desirable for Bessemer purposes. The magnetic separator is no longer a novelty. It was only at the last meeting of the Institute in New York, that Messrs. Birkinbine and Edison (Trans., xvii., 728) and Mr. Robert Cook (Ib., 599) described some of' the principal designs of apparatus in use. E. S. Dana classifies the natural magnetic minerals in the following order: magnetite, pyrrhotite, franklinite, almandite and minerals " containing considerable FeO." Of these, ferroso-ferric oxide, or magnetite, is the type. The proportionate amount of ferrous oxide essential to magnetize the compound, is not established. Ferrous oxide appears to be the only magnetic oxide of iron. Tissandier (Compt. Rend., lxxiv., 531) produced ferrous oxide in the anhydrous state, by the action of carbon dioxide on pure iron at a red heat, according to the equation : Fe + CO2 = FeO + CO. This is black, crystalline and magnetic. Heated to redness in the air, it is converted into Fe3O4. Sidot (Compt. Rend., lxvii., 175) has also produced an oxide exhibiting magnetic polarity, by heating ferric oxide to a temperature

Jan 1, 1891

By E. M. Boatman, S. J. Pirson, R. L. Nettle

More than a decade ago some theoretically derived relationships were proposed that permitted the prediction of the relative ability of reservoir fluids (oil, gas, water) to flow simultaneously within the porous structure of rocks. While many successful predictions of fluid-flow ratios from wells have been made from electric logs, the theoretical deductions have not been subjected to laboratory verifications. For the relative permeability to liquid and gas, simultaneous flow of water and air was used, employing the familiar external-gas-drive technique. While in the process or water desaturation, the resistivities of the test cores (Woodbine sand) were measured. Their permeabilities varied from 10 to 280 md. The laboratory results indicated the need for changing the values of exponents in the theoretically derived formulas, whereas the form of the functions of resistivities and saturations remained as derived by theory. For the relative permeability to oil and water, the three-core dynamic technique was used on six cores. In this case it was found that the formulas were valid with but slight modification as derived by theory, to a sufficient degree of accuracy for engineering use and for well productivity prediction from electric logs. INTRODUCTION Conventional quantitative well log interpretation generally stops after reasonable values for formation porosities and water saturations have been established. Yet such in formation is insufficient to predict either the outcome or the desirability of well tests and completion, let alone the economic aspect of producing oil- and gas-bearing sections which might have been established by conventional log analysis. To give an engineering answer to such questions, it is necessary to predict the expected fluid flow ratios, the productivity index and the ultimate recovery from potentially productive wells. This problem requires in part for its solution a determination of the relative permeability characteristics of the potentially producing formations, both to oil (or water) and gas, and to oil and water.

Jan 1, 1965

By R. A. Bowen

THE Montreal mine, operated by the Montreal Mining Co., is four miles west of Ironwood, Mich., in Iron County, Wisconsin. It is the westernmost producing property on the Gogebic Iron Range of the Lake Superior district. Its production is 4000 tons of hematite per day from an average depth of 2300 ft. This tonnage is hoisted through a five-compartment vertical shaft. Several mining methods are in use but in the larger orebodies the method is sublcvel caving. ORE OCCURRENCE AND CHARATERISTICS The main orebodies of the Montreal mine lie in eastward-pitching troughs formed by the intersections of intrusive dikes with impervious members of the iron formation. The formation has an average width of 425 ft., strikes east and west and dips north at approximately 62°. The dikes dip south at 45° and bear north 20° from the strike of the formation, making the intersections with various members pitch east at 16°. The dikes are not continuous across the formation but are interrupted by a longitudinal bedding fault, which has a throw of 380 ft. vertically and 885 ft. horizontally. North of the fault zone the formation is called "hanging," south it is called " footwall" formation. The faulting has created two ore-bearing horizons; viz., at the inter- section of hanging dikes with the slate members and the foot dikes with the quartzite footwall. In each formation the main ore-bearing member consists of wavy beds of hard chert interlaid with hematite. In the most productive part of the mine, where the sublevel caving method is employed, a hanging-formation dike lies in juxtaposition with a footwall dike and the iron formation carries ore in certain sections from the quartzite footwall to the hanging dike. Here a horizontal section of the orebody may be 300 to 400 ft. wide for a length of 150 ft. or more and tail out lengthwise in both the foot and hanging formations. The footwall ore section is much longer than the hanging wall because of the orebody's greater height above the "pitching dike."

Jan 1, 1938

By Basil Prescott

In the science of evolution of the species, there is a law which, simply stated, avers that the history of the individual repeats the history of the race. Similarly, if unassisted by the experience of others, each SUCcessful engineer would perfect and develop his methods and technique from the rough grab sample up to the carefully taken, recorded and studied samples and observations which now determine and decide the expenditure of the large sums required for modern mining. Hence real progress results from the interchange of experience, as in that way the entire experience of the profession can be made available to every engineer. The literature of this subject is already voluminous, but some progress has been made in the gathering of data and in their compilation or presentation. More Romance than Developed Ore in Limestone Deposits In certain types of deposits, methods little known 25 yews ago have been developed, perfected and standardized to a degree that lifts sampling from the field of manual labor to the plane of an applied science; for instance, the modern methods of estimating the so-called porphyry coppers. With some types of orebodies, we have become familiar, and sometimes on too easy speaking terms, with millions of tons of ore in reserve and tens of millions of dollars; while with others even these once marvelous figures sink into insignificance before the array of ciphers that are presented. .In such company, the average limestone replacement orebody is entirely out of place, for its reserves seldom reach a figure to arrest the attention of the modern engineer, and more often are insignificant. It is only when the records of the production of some of the older camps are studied, which show them to have continuously produced over very extended periods, that a realization is borne home that occasionally this type is worthy of consideration even in these days of gigantic tonnage.

Jan 1, 1925

By G. B. Gould

THE softening temperature of coal ash, as determined in the labora-tory, has been used for years as an indication of the tendency of coal to form clinker and slag. It has not, however, provided an index of unfail-ing and complete reliability for forecasting the relative performance of different coals, even in the same plant and under the same conditions of operation. A large amount of investigation has been done in attempts to account for this, but practically all of this has been in the direction of studying the chemical constitution of coal ash. Much has been learned, but without much progress toward a better practical application of knowledge to the selection of coal. There has always seemed to be an important factor that eluded discovery. It has long been recognized that the ash-softening temperature as determined in the laboratory had one weakness, in that it represented the temperature at which a thorough mixture of the whole ash of the coal would melt, the mixing having taken place before the heat was applied. This does not, of course, simulate what takes place in practice. With certain exceptions, which will be noted later, almost all of the investiga-tions that have been made, especially the chemical studies of coal ash, have likewise dealt with a mixture of the whole ash in the coal. The present investigation differs from these fundamentally in its approach to the problem. It started from the observation that some clinkering and slagging obviously takes place in a boiler furnace at tem-peratures below the laboratory-determined ash-softening temperature of the coal. Knowing that coal as it is shipped to market is a hetero-geneous mixture of diverse mineral substances, each of which may act independently of the others in the initial stages of clinker and slag formation, it seemed logical to investigate separately the softening tem-perature of the ash in physically separable portions of the coal, and to determine the relative quantities of ash contributed by each portion.

Jan 1, 1940

By Craig S. Hartley

Expressions are obtained for the energy changes associated with the reaction of (a& (111) slip dislocations on intersecting (110)planes in anisotropic bcc metals. An energy criterion for assessing the likelihood of dissociation of the products of such reactions is also presented. It is found that the "burrier reactions" which form a(100) dislocations at the intersection of two active {110) slip planes are more energetically favorable in metals which exhibit a high value of Zener's anisotropy factor, A, than those which have a low value. The results are presented in a form which permits the stacking fault energy to be obtained from a measurement of the separation between par-tials in a dissociated configuration. However, until accurate calculations or measurements of the stacking fault energies involved are available, it is not possible to assess the physical importance of dissociated dislocations. In a recent paper,' the energy changes associated with several types of reactions between two slip dislocations, (a/2)(111){110), in bcc structures were calculated.* Isotropic elasticity and the approxima- tion v = -3- were employed. The purpose of this work is to present calculations of the energy changes for many of the same reactions using anisotropic elasticity. The problem of dissociation of a(100) and a(110) dislocations is also considered, and maximum fault energies for which dissociation will be energetically favorable are calculated for several bcc metals. Two general types of reactions are considered; those for which the reactant (a/2)(111) dislocations have long-range attractive forces and those for which the reverse is true. An example of the former is: (a/2)[lll] + (a/2)[lll]-a[l00] while the latter are typified by: (a/2)[lll] + (a/2)[111] -a[011] Only reactants lying in different slip planes are considered; therefore, the products must lie along (111) or (100) directions, which are the intersection of two {llO} planes. It will be assumed that the reactants and products are infinitely long parallel dislocations, since in this case the energy change associated with the reactions is a maximum.' THEORY The self-energy per unit length of a straight mixed dislocation in an anisotropic medium can be written? where b is the Burgers vector, K is an appropriate combination of the single-crystal elastic constants, and R and ro are, respectively, outer and inner cut-off radii of the elastic solution. The energy given by Eq. [I] does not account for any variation of the core energy with orientation. This could be manifested by an orientation dependence of the core radius or, equivalently, the Peierls width, of the dislocation. However, the energy contribution due to this source is expected to be small, and current models of the dislocation core are not sufficiently accurate to justify such a refinement. It has already been shown that for the isotropic case the energy contributions due to nonzero tractions across the cores of the reactants and products exactly cancel one another in the reaction.' Accordingly, it will be assumed that this contribution to the total energy change in the anisotropic case is small. In the subsequent discussion it is also assumed that the core radii of the reactant and product dislocation are the same and that, where stacking faults are formed, the faulted region is bounded by the centers of the partials. Consequently only changes in elastic energy due to the reactions will be considered. When the dislocation is parallel to either the (111) or the (100) directions, K may be written:375 K = (Ke sin2 a + Ks cos2 a) [2] where K, and Ks are the combination of elastic constants corresponding to an edge and screw dislocation lying along the same direction as the mixed dislocation, and a is the angle between the direction tangent to the dislocation line and the Burgers vector. Eq. [2] should not be confused with the isotropic approximation to the variation in energy with line Orientation.6 It should be noted that the essentially isotropic expression for K is a result of the characteristic symmetry of the (111) and (100) directions and is not, in general, valid for other dislocation directions in anisotropic cubic metals. The energy* change for a reaction in which the re- actant and product dislocations are parallel perfect dislocations can be written: where Ep and E, refer to the self-energies of the products and reactants, respectively. For dislocations parallel to (100) and (111) directions, Eq. [3] becomes:

Jan 1, 1969

By N. M. Langdon

ON account of its greater economy and cleanliness, and the extent to which iuferior fuels can be utilized for its generation, there has been of late a rapidly increasing tendency to substitute gaseous for solid fuel for metallurgical and other industrial purposes, which has led to various improvements both in the process and apparatus for its manufacture, notably among which may IN mentioned the Tessie Gas Producer* described by Mr. A. L. Holley. A considerable experience by the author with producers of various kinds using anthracite fuel, in connection with the Taylor ore-roast ing furnace at Chester, N. J., with more or less unsatisfactory resnlts, finally led to the erection of the experimental producer described below, which combines in its construction, simplicity, dorability, and accessibility for cleaning and removing the ash and manipulating the fuel. It is adapted to the use of various kinds of fuel, and is capa ble of generating a steady and uniform supply of gas at a minimum cost. It has been in constant use day and night for about five months, and has given very satisfactory results. The producer, as shown in the accompanying sketch, consists of a cylindrieal furnace, preferably inclosed in an iron jacket or casing, having a bosh or inverted cone-shaped base as in a blast furnace. It is provided with a bell and hopper, the bell being hung to a forked lever, and having guide flanges to prevent it from swing ing, and an opening at its apex through which to insert a bar for manipulating the fuel, which when not in use is closed with a plug or stopper. Cleaning is effected through two small doors placed opposite each other at the hearth level, which in order to facilitate the operation is elevated above the floor. The doors are removable, and when closed are by their own weight and by means of bevelled lugs held tightly against their frames, forming a practically tight joint. A blast of commingled steam and air is injected into the fuel through a series of tuyeres underneath the bosh communicating with steam jet injectors. A small flue also connects the door

Jan 1, 1884

By Albert F. Schneider

The peculiar conditions under which lead and silver ores are now smelted in Salt Lake Valley, Utah, render it advantageous to make slags that are siliceous and carry a high percentage of lime. The ores treated come from Utah, Idaho, Montana, and Nevada; nine-tenths coming from Utah. These ores are siliceous, and contain but little fluxing material; the iron present scarcely sufficing for the sulphur and arsenic. Formerly ores high in iron were furnished from Cottonwood Canons, but this supply is meagre. The ores from Idaho, Montana, and Nevada are high grade galenas, more or less impure. They generally require a large amount of iron for the sulphur, arsenic, and antimony. In the winter and spring of 1880—'81, we were treating at the Germania Works principally Bingham Canon lead ores and silver ores from Tintic. The former contained 15 to 30 per cent. silica, and only 4 to 5 per cent. iron (generally as iron pyrites); they ranged high in lead and low in silver, and carried more or less alumina; the lead was present chiefly as carbonate. The silver ores from Tintic contained sulphides, and ran as high as 80 per cent. silica. We had been running on a slag approximating 40 per cent. FeO, 20 per cent. CaO, 10 per cent. other bases, and 30 per cent. SiO2, which had heretofore been very satisfactory; but with the class of ores then coming in it was of the utmost advantage to use less iron and more lime, especially as the iron ore (50-55 per cent. iron and 5-7 per cent. silica) cost four times as much as the limestone (50-55 per cent. lime, 3-5 per cent. silica). The large amount of alumina in the slag influenced them for the worse; the addition of lime to the charge helped them, and after many experiments regarding the proportion of FeO, CaO, and SiO2, a slag approximating 35 per cent. FeO, 25 per cent. CaO, 5 per cent. other bases, and 35 per cent. SiO2 seemed to answer

Jan 1, 1883

By H. Evans, L. A. Hunt

The raw materials used in the blast furnaces at the Geneva Works of U.S. Steel Corp. have a high degree of variability in size-consist and chemical content. To overcome the problems caused by the use of highly variable raw materials in the blast-furnace operation, large blending piles are used. An investigation of the blending system showed that variability was significantly reduced by the use of these blending piles. This paper describes the blending system, the method used to evaluate the effect of the blending system, and the results of the investigation. The remarkable improvement in blast-furnace performance since World War II has been in large measure due to the use of beneficiated burdens to reduce the variation of both chemical analysis and size consist. This practice allows the operator better control of his blast furnace, enabling him to operate more economically and to produce chemically uniform hot metal. Planning for uniformity usually begins during mining by extracting ore from several locations within a pit to produce a final product with a specified analysis. Another approach is to stockpile ores having widely varying analyses in individual piles and then drawing the required amounts from each stockpile to provide a blend with a predetermined analysis. A further improvement in blending procedures is to take ores from several sources, bedding them horizontally in layers and reclaiming by vertical slicing from the blended pile. The last procedure, when properly executed, is one of the best methods for achieving maximum uniformity with highly variable ores. U.S. Steel Corp.'s Geneva Works initially obtained iron ore for their blast furnaces from the Cedar City area of southern Utah. These ores were formed primarily by the replacement of limestone and siltstone formations with solutions containing iron and several other elements. The original iron content of the ores in the ground varied from a high of 70% to practically zero with a minimum acceptable iron content for shipping of 37%. Substantial amounts of silica, calcium, magnesium and alumina were found in the ores along with minor amounts of copper, lead, zinc, nickel, titanium, manganese, phosphorus and sulfur which were sometimes present in detrimental amounts. Iron-to-silica ratios varied considerably throughout

Jan 1, 1968

By Carl O. Thieme

INDUSTRIAL specialization has rapidly created a demand for new and better alloys. A more thorough understanding of the requirements of specific industries and the discovery of processes by which it has been made possible to utilize certain metallic elements in larger quantities, together with greater purity in most virgin metals, has resulted in a great increase in the number and variety of alloys. This increase in the number of elements entering into non-ferrous alloy manufacture has increased the contamination in secondary metals, as received by the smelter. In order to successfully refine scrap of a much more complex nature it has been necessary for the smelter to install new equipment and to constantly search for new methods of refining as well as to improve the methods already in use. In some instances it is largely responsible for specialization in the secondary metal industry. It is usually necessary to remove the contaminating impurities in scrap or eliminate to a degree not objectionable, no matter how introduced. In order to know what metals or elements are of an objectionable nature, the smelter must make a study of the effect of single elements and combinations of elements on many of the physical and mechanical properties, such as strength, ductility, color, hardness, casting soundness, etc. The smelter of secondary metals must be able to utilize scrap materials of all kinds, for, sooner or later, most fabricated metal parts and the waste products formed during the process of manufacture find their way to the smelter. In addition, the smelter, in contracting for salvaged material from the producer, gets some undesirable and unusual materials. The value of secondary metals and the importance of their recovery has increased in recent years and it is no longer sufficient to recover the base metal only, such as copper from copper-bearing material, but it is essential that other elements present be recovered and converted into useful form. This paper will deal not with the smelting and refining or separation of impurities from secondary metals but an attempt will be made to show the increased value of metal scrap when properly segregated and in a less contaminated condition. There are five principal reasons for contamination in secondary metals: New and unusual alloys, productive methods,

Jan 1, 1928

By F. A. Harvey

DURING the last few years, there has been a constant tendency toward increasingly high temperatures in many lines of industry. The necessity for increased production of coke gave a 16-hr. coking period where we used to have 24 or even 30-hr. periods. Mechanical stokers have increased the temperatures in boiler furnaces. The use of a chain-grate stoker with coke-breeze fuel necessitates a low arch in the furnace and temperatures run much higher than firebrick were formerly called upon to stand. These higher temperatures have necessitated more rigid tests and the separation of even high-grade firebrick into separate classes. More severe tests mean higher temperatures and closer control, and this, in turn, means better pyrometer tubes. Platinum couples do not have a very long life when used at temperatures exceeding, say, 1200° C., but the proper selection of a high-grade brick to be used in a-boiler setting will save the cost of several couples; and with freight rates continually increasing it becomes increasingly important to know where brick may be bought, nearest the job, that will prove satisfactory. Couples that will stand continuous use at higher temperatures, of tungsten, molybdenum, or other metal, will doubtless be forthcoming as soon as the need becomes sufficiently urgent. There are on the market several tubes that are entirely satisfactory for temperatures up to 1300° C., if properly protected from heat shock. The Semet-Solvay Co. makes a practice, however, of testing silica and clay firebricks at 1450° C. for 72 hr., and so far we have been unable to find a tube that- is entirely satisfactory for this purpose. A graphic record of the temperature during the entire test is extremely desirable and hence we have been trying out anything that gave promise of success. Several years ago we were using insulating tubes, made purely for electrical work, of vitrified porcelain. These tubes stood up under any temperature we were able to reach and were apparently impervious to gases. When this stock gave out, due to mechanical breakage, we tried other insulation tubes only to have them melt down like glass. Marquardt Masse tubes were tried next under a statement from the distributors

Jan 9, 1919

By E. L. Estabrook

Oil companies operating in foreign countries have made increased use during 1930 of cooperative agreements in prospecting unproven territory and in developing proven territory. A considerable proportion of the existing foreign fields are held under single ownership and have been developed as units. The operating companies seek to maintain this condition by acquiring land in large blocks or by merging smaller concessions under unit operating agreements. Your committee has secured a number of reports on foreign non-competitive operations and a valuable formal paper which deals with the Golden Lane of Mexico.l 0. B. Knight has drawn some valuable comparisons concerning costs of development and ultimate production between a certain competitive area and another noncompetitive area along that strip of rich producing territory. After discussing details of production, numbers of wells and cost of drilling and production, he concludes that the 93,700,000 bbl. of competitive oil will have cost $0.193 per barrel as compared to 85,700,000 bbl. of noncompetitive oil at $0.059 per barrel. The saving of $0.134 per barrel amounts in round figures to $11,500,000, which is 21/2 times the total cost of developing the non-competitive block. The cost for drilling and development of the two areas of 3550 acres which Mr. Knight compares was $18,150,000 under competitive and $4,760,000 under noncompetitive conditions. The estimated ultimate production of the competitive 3550 acres is expected to reach 93,700,000 bbl., or 8,000,000 bbl. above the noncompetitive block. Under ordinary production methods it is to be expected that a noncompetitive area will yield less oil per unit area than will a similar competitive area, because of closer spacing and more rapid extraction in the latter case. The advantage of noncompetitive conditions, in this regard, is that the undivided ownership makes it possible to apply repressuring as a production method fairly early in the life of the field and thus, in the end, to vastly increase the yield. In the particular region described by Mr. Knight, repressuring is not thought to be a factor of

Jan 1, 1931

By L. G. E. Bignell, E. A. Hanson, J. O. Lewis

There was a general upward trend of all petroleum activities in Oklahoma during 1933, especially in the last quarter, though the average price for crude oil for the year was below 1932 and overproduction in East Texas demoralized conditions for several months. As the Federal Agencies established under the National Industry Recovery Act began to function conditions in Oklahoma were visibly affected and production of "hot oil" from Oklahoma City field practically ceased. More oil wells were completed, six new fields discovered, extensive geophysical prospecting carried on in many parts of the state, and an increased amount of clean-out and reconditioning work done in older pools during the past year. To increase quantity and rate of recovery of oil many wells producing from limestone formations were treated with hydrochloric acid. The Oklahoma City field had a reported production of 66,984,570 bbl. of oil during 1933, or almost exactly twice the amount reported for 1932. Potential tests and other indicators show that the wells are rapidly approaching the stage where mechanical lift must be applied to all of them, including the wells in the Wilcox sand zone. The major pipe-line companies serving the oil fields of Oklahoma have been anxious to make connections to all wells during 1933 and have indicated a willingness to accept all oil available. The amount offered has been restricted by proration orders, but the state did produce 178,356,-449 bbl. during the year, an increase of about 18.6 per cent over 1932. This increase was practically all due to the greater demand for Oklahoma City oil, as all other pools of any importance showed loss in production or very small gains last year. The U. S. Bureau of Mines' laboratory reports the lubricating distillate content of Oklahoma City crude from Arbuckle and Lower Simpson formations about 17 per cent by volume. East Texas oil is reported to contain about 5.6 per cent lubricating distillates. The desire of the transportation companies to keep their equipment operating at maximum capacity and the need of some refineries for more

Jan 1, 1934

G. McMEANS*—This paper is a very good demonstration of the use of a new tool for the solution of industrial problems of a physical nature. To have solved this problem without the use of radioactive tracers would have required a far greater amount of investigation and work. The authors have described their work in such detail and in such a manner as to suggest the possibilities of the use of such tracer studies in the solution of many perplexing problems of industry. R. E. BREWER †—The authors are to be congratulated on this very excellent paper, which describes the results obtained by utilizing radioactive tracers in investigating the complicated reactions in the coking process. The paper has especial merit because the study was made under typical operating conditions in a commercial coke oven. Considering all the difficulties involved, the results obtained are excellent. My brief comments will relate primarily to the method used for determining the sulphur in the gas and to the value of the data obtained in stimulating further research. In discussing their results, the authors state (p. 17) that "Of the total sulphur charged to the oven in the coal 95 pct was accounted for in the coke and gas. Most of the sulphur unaccounted for is believed to have been lost in the gas, since the gas sampling system removed all the H2S but probably only part of the organic sulphur compounds present in the gas. Although sulphur in the acid-insoluble tar residue in the gas scrubbing bottles was determined and added to the total sulphur determination of each sample, it is doubtful whether an aliquot portion of tar for each sample was obtained." The reviewer believes that the authors might have obtained a better sulphur balance had they used an improved procedure for determining sulphur in the gas. The absorption train should have included a trap to remove the tar and am-moniacal liquor and a wash bottle containing 5 pct sulphuric acid to remove ammonia in the coke-oven gas before it entered the gas scrubbing bottles containing the ammoniacal cadmium chloride solution. Sulphur in the tar and

Jan 1, 1950