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By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By J. P. Zannaras

Referring to the article by R. J. Charles and P. L. de Bruyn, let us assume that W = weight of glass bar; P = weight of hammer; e = total deformation; K = unit of deformation; K = potential stress energy; E = modulus of elasticity; L = length of the bar; 7 = coefficient of inertia; h = height, ft; V = velocity, fps; and a = cross section area. The portion of kinetic energy which is effective in producing stress energy in a fixed bar struck horizontally is given by the formula" 1 + 1/3 W/P K =1+1/3w/p/(1+1/2w/p)2 P.V2/2g = Ph where 1 + 1/3 W/P ? (1 + 3W/P/(1=1/2w/p)2 [8] Putting e e = W/P =------------ From the above equation it can be seen that the maximum transfer of kinetic energy to stress energy is when e = 0 or W/P = 0 which indicates that the weight of the hammer must be very large as compared with the weight of the impacted rod. Eq. 8 diametrically opposes the conclusions reached by the authors of this article. In fact, if their suggestions were followed to the extreme when e = co when P = 0, there would be no transfer of kinetic energy to stress energy at all, as 7, becomes zero. Eq. 8 presumes that the velocity with which the stress is propagated through the bar is infinite, whereas the authors claim that the compression waves reflected are reaching the struck end of the bar prior to the complete transfer of the kinetic energy to cause such modification of the conditions there as to make them reach the reverse conclusions demonstrated by the above formula. That such interference exists is unquestionably demonstrated by the authors and others. However, if my observations are correct, such interference for this specific experiment and also for practical comminu- tion is insignificant, and the conclusions of the authors are in error and must be reversed to comply with Eq. 8. Eq. 5, w. = AE 2/2, given by the authors on page 51, is derived from the following equation (Eq. 9): K = 1/2Pe, where P = Sa, S = ?E, e = EL, and L = 1. The above formula, Eq. 9, cannot be applied in this case. This formula is applicable for static loads where the load increases from zero up to its final value, P, in such a way that the deformation at different instants is proportional to the loads acting at those instants and actually represents the area of a right triangle in the strain load diagram of base e and height P. The typical photographs shown in Figs. 3 and 4 represent the familiar strain load diagrams, and since the line of the wave marks the existence and intensity of the strain with the unquestionable conclusion that such strain has been caused by the action of a load acting continuously all along the wave until it reached the horizontal axis, the work stored at this point is represented by the area under the wave line and the horizontal axis and not by the area of the fictitious triangle given by the authors. Then if this is correct, even visual estimation of these areas at gage stations given in the typical photographs of Figs. 3 and 4 suffice to contradict the authors' calculation given in Figs. 6a and 6b and Figs. 7a and 7b. The typical photographs presented by Charles and de Bruyn show a considerable variation of the intensity of the strain at different stations but very small variation of areas which actually represent the stress energy at the corresponding stations. And, apparently, by squaring the small quantities, the authors magnified their error tenfold. J. M. Frankland's paperV iscusses the relative strain intensity and not the total energy for different types of impact loading. He states in his paper, "The reader is explicitly warned not to confuse the results in this report with those obtained when the load is applied by a blow as from a hammer. In this case the peak load rises to very large but mostly unknown values. The accompanying large deflections and stresses are the result of high values of P, not of the dynamic load factor n. According to Frankland "the dynamic load factor" is the numerical maximum of the response factor. It therefore appears that the authors followed the same procedure in obtaining the relative strain energy ab-

Jan 1, 1957

By E. Eichen, G. S. Ansell. L. R. Sefton

A Technique for the Preparation of Thin Films of Two-Phase Alloys Suitable for Use in Transmission Electron Microscopy In order to obtain foil sufficiently thin to permit transmission electron microscopy of two-phase alloys, particularly alloys in which the second phase is present as a very finely dispersed second phase, a technique has been developed to permit the thinning of these materials without either the large degree of pitting present in the electrochemical or chemical thinning process or the slicing of the second phase particles accompanying the mi-crotoming process. The metal or alloy to be thinned is first ground to initial sheet thickness of approximately 0.002 in. by hand grinding on 4/0 emery paper. This starting sheet of material is then reduced in cross section by an air-abrasion technique using an S. S. White dental abrasive unit. This unit consists of a supply of compressed CO, which provides a stream of high velocity gas impinging on the metal surface. Incorporated in the machine is a hopper-type of device in which fine abrasive particles of alumina are introduced into this high velocity gas stream. The abrasive particles are irregular in shape, approximately 30 across. If any particles are introduced into the sample during thinning, they should be readily observable. Impingement of this gas abrasive mixture on the metal surface causes rapid removal of the metal in a rather uniform manner. The rates of metal removal may be adjusted by both regulating the gas pressure and amount of abrasive material introduced into the gas stream. By moving the gas stream abrasive mixture slowly about the specimen or sheet surface, a large area of the sheet may be uniformly reduced to a thickness which is then suitable for use in transmission electron microscopy. Attendant with this metal removal is a result in cold working of the metal surface due to the impingement of the abrasive particles. The resultant foils, therefore, if made from materials which are desirable to view in the annealed condition, have to be subsequently heat treated. If the alloy from which these foils have been made has been initially heavily worked or deformed during manufacture, then no subsequent heat treatment is, necessary as in the case of the SAP-type alloys. Following this initial abrasive thinning technique, the material is then given a final electropolish which is just sufficient to remove any oxide which has formed on the metal surface during the abrasion technique. Care has to be exercised in this technique to make sure that none of the attendant problems associated with the chemical or electrochemical removal process are encountered. An electrolyte solution which has been found suitable for aluminum-aluminum oxide SAP materials is one containing 78 ml of perchloric acid, 120 ml of distilled water, 700 ml of ethanol, and 100 ml of butylcellosolve. The current density used is 1 amp per sq cm and a polishing time of 3 sec on each side of the foil is used. Extreme care must be used in choosing the current density and time to ensure that just the oxide film on the specimen is removed. Times or current density which are greater than this will lead to perforation of the film due to the removal of the A1,0, particles or a complete loss of the film due to complete solution of the matrix. Once the oxide film is removed the specimens must be kept in alcohol until placed in the microscope to prevent any further oxidation. Fig. 1 shows -an electron micrograph at X28,000 of a thin foil of an aluminum-aluminum oxide SAP-type alloy prepared using this technique. The alloy consists of a matrix of commercial purity aluminum containing a very finely dispersed second phase of aluminum oxide flakes. In the electron micrograph, the dark patches are the aluminum oxide particles. The lighter portions of the micrograph are the aluminum matrix. Within the aluminum matrix, the dislocation structure of the alloy is clearly visible. This structure consists of intersecting twin boundaries and single dislocations. No abrasive particles were observed in the samples. In addition, thin films of commercial purity aluminum were prepared using this method and observed in the electron microscope. Here also, there was no evidence of abrasive particles introduced in the samples. The technique presented provides a satisfactory method for the production of thin films of aluminum-aluminum oxide SAP-type alloys for use in transmission electron microscopy where none of the usual thinning techniques appears to be satisfactory.

Jan 1, 1962

By J. T. Burwell

THE crystal orientation in silicon iron that has been given a particular treatment described by Goss,1 has been studied by Goss, by Bozorth2 and by Sixtus,3 but their results do not agree and are almost entirely qualita-tive. The following pages give a description of the orientation of the crystals in the material as determined by means of X-rays and expressed graphically in a pole figure. The treatment consists of hot-rolling the material into strip and then giving it two cycles of cold reduction followed by recrystallization. This produces remarkably good properties in the rolling direction, the product having a maximum permeability in this direction of more than 25,000 and a permeability at 16,000 Gauss of more than 2200, It is now generally agreed that Goss's original conclusion' that the crystals in the final product are randomly oriented was incorrect, since Bozorth2 and Sixtus,3 using different methods, have shown that there is a decided preference in orientation. Those writers, however, expressed their results only in terms of one or more ideal orientations with deviations therefrom.

Jan 1, 1940

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO2 oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO2 can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1 ½ ¢ per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past year that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 ½ million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ¾ ton of high-grade pyrite and results in ½ ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15¢ per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1952

By Pierre R. Hines

These notes are based on observations made while on a recent trip through the West, for the purpose of studying the practical operation of the ball-mill. The writer takes this opportunity to express his thanks for courtesies extended at the many plants visited as well as for the valuable data received: While there are several types of ball-mill on the market, particular attention will here be given to the diaphragm type, as the open-trunnion type, especially the conical mill, has been thoroughly discussed in the Transactions. There is a prevailing impression that the ball-mill is a recent development; however, ball-mills were used extensively in Montana and other western states for crushing ores for concentration as far back as 1808. Its present prominence is due in part to its recent successful application by one of the large copper companies. Without any reference to dry grinding, the first successful ball-mill for wet crushing, which is still in operation, was built 10 years ago.l This mill, designed by Erminio Ferraris for crushing Sardinian ores for concentration, is of more than passing interest. It embodies the peripheral discharge with grates, large forged-steel balls, and the principal features of the modern ball-mill. The results approach present-day practice, the chief differences being that the mechanical construction has been improved in the modern types. Crushing Action of Ball-mill The action of the balls and the principles of crushing have been studied by several investigators.2 Their conclusions are confirmed by results obtained by the writer in experimenting with a small machine built at

Jan 1, 1918

By H. A. Schwartz, A. F. Gorton, H. R. Payne

In a previous paper1 we published what we believed to be a correct stable equilibrium diagram for an iron-carbon alloy containing + 1.20 per cent. silicon. The purpose of the present paper is to record the results of studies concerning the effect of silicon on the location of the eutectoid points of both systems, including also certain incidental observations of general interest. We prepared white cast irons containing roughly 21/2 per cent. carbon, 0.05 per cent. manganese, and 0.03 per cent. each of phosphorus and sulfur, but of accurately known silicon contents which varied from 0.40 to 3.32 per cent. These were completely graphitized, and both A1 points determined, as in our previous work, the specimens being converted from the stable to the metastable condition and vice versa by appropriate heat treatments. Each stable alloy was heated to a few degrees above Arl for 3 to 7 hr., quenched and analyzed for total and graphitic carbon, the difference being the "combined" or, better, agraphitic2 carbon content of the saturated solid solution just above the eutectoid transformation. One specimen was lost through accident, and the 0.40-per cent. silicon material gave an anomalous result ascribed to difficulty in maintaining the stable system in the presence of but little silicon.

Jan 1, 1923

By H. R. Payne, A. F. Gorton, H. A. Schwartz

In a previous paper1 we published what we believed to be a correct stable equilibrium diagram for an iron-carbon alloy containing + 1.20 per cent. silicon. The purpose of the present paper is to record the results of studies concerning the effect of silicon on the location of the eutectoid points of both systems, including also certain incidental observations of general interest. We prepared white cast irons containing roughly 21/2 per cent. carbon, 0.05 per cent. manganese, and 0.03 per cent. each of phosphorus and sulfur, but of accurately known silicon contents which varied from 0.40 to 3.32 per cent. These were completely graphitized, and both A1 points determined, as in our previous work, the specimens being converted from the stable to the metastable condition and vice versa by appropriate heat treatments. Each stable alloy was heated to a few degrees above Arl for 3 to 7 hr., quenched and analyzed for total and graphitic carbon, the difference being the "combined" or, better, agraphitic2 carbon content of the saturated solid solution just above the eutectoid transformation. One specimen was lost through accident, and the 0.40-per cent. silicon material gave an anomalous result ascribed to difficulty in maintaining the stable system in the presence of but little silicon.

Jan 1, 1923

By James R. Cost

The increasing emphasis upon the use of refractory metals has resulted in a closer look at some of the properties of these materials. In the course of a study of the Nb-N system it became necessary to re-examine one such property, the spectral emissivity, e. Experiments were performed in order to determine e for solid niobium as a function of both temperature and nitrogen content. The results of these experiments, which indicate a small but significant correction to the spectral emissivity of pure niobium, are reported herein. In addition, data on the total emissivity, et, of niobium and the spectral emissivity of the nitride Nb2N are given. To this author's knowledge, neither of these emissivities are as yet reported in the literature. Previous studies of the spectral emissivity of niobium have been few. In 1915, Burgess and Waltenburgl obtained a value of 0.49. Whitney? in 1935, used a Mendenhall V to obtain the commonly accepted value of 0.374. A recent investigation of sintering of niobium graphically reported a value of about 0.36= using a solid block of the metal with a blackbody hole drilled into it. In the present experiment a method similar to that of Worthing4 was used. An 8-in. long and 3/16 in. diam hollow tube was formed from smooth surfaced 0.0013-

Jan 1, 1962

By William B. Plank

FOR the third consecutive year, I have collected the data on enrolment and employment of graduates from the schools in. the United States and Canada that grant degrees in mineral technology. The data were collected in the same way as formerly; namely, by questionnaire, dated Jan. 5, 1935, and are therefore comparable with the figures reported for former years. This year, along with the names of the 48 schools in the United States and the six in Canada, are listed the courses offered and the degrees granted by each, school, together with the names of the professors in charge. The names and addresses of the four schools in Mexico are also given. The list is on pages 450-1, which should he valued as a directory by those interested in mineral technology education. So far as is known, this is the most complete and up-to-date list that has ever been published.

Jan 1, 1935

By George B. Clark

COMMERCIAL strip mining of coal was first begun in the state of Illinois in 1911.1 The annual tonnage of coal produced from coal strip mines in the state was very small until 1924, when the strip mines shipped 1.68 pet of the total state shipping mine tonnage. Since then, strip mines have gradually come into their own by producing a sizable percentage of the coal produced in the state. Table 1 shows the coal tonnage shipped from both underground and strip coal mines in the state of Illinois for selected years. The table illustrates the major role that strip mines play in the state's coal industry. Today, strip mining is generally conducted in areas where the coal seams lie so close to the surface of earth that a stripping method of mining the coal seam is cheaper than an underground method. Strip mining not only recovers 100 pct of the coal deposit, but may also be used as a method of mining large acreages of coal abandoned by the closing of underground mines.

Jan 1, 1949

By W. L. Phillips

Nickel alloys containing approximately 0.5, 2.0, and 6.0 at. pct of Os, Pd, Ru, and Rh were Prepared by vacuum melting. Tension tests were carried out at 25°, 500°, 800°, and 1000°C; stress-rupture tests were carried out at 650° and 800°C. The room- and elevated-temperature tensile strength, as well as the stress-rupture life, increased for the same atomic percent in the order palladium, rhodium, ruthenium, osmium. The ductility decreased with increasing concentration or temperature for a given concentration. The results are discussed in terms of solid-solution theory. BOTH solid-solution strengthening and weakening have been observed in alloys. The simple rules which have been suggested to explain alloy strengthening are subject to several exceptions and deviations based on atom size and valency effects. The purpose of the present investigation was to extend the knowledge of the effect of solid-solution alloying on the mechanical properties of several Ni-Pt group metal alloys at room and elevated temperatures. The platinum group metals, osmium, ruthenium, rhodium, and palladium were chosen as the solute elements because of their different valencies and atom sizes. Nickel was chosen as the base material because of the large amount of previous literature available on nickel alloys. EXPERIMENTAL PROCEDURE Grade A Carbonyl nickel powder having the following composition: C, 0.07 pct; Fe, 0.013 pct; Mn, 0.0005 pct; Si, 0.016 pct; Mg, 0.0003 pct; S, 0.003 pct, was purchased from the International Nickel Co. The platinum-group metal powders were purchased from A. D. Mackay Co. The powders were weighed, cone-blended for 4 hr, and vacuum-melted into 1-in.-diam billets, approximately 2 in. long. The ingots were homogenized for 5 hr at 1800°F before being extruded into 1/4-in. rods at 2100°F. The alloy compositions are shown in Table I. Typical impurity values were not significantly different than that of the as-received nickel. All alloys were single phase. The lattice parameter of each alloy after the heat treatment described below is also included in Table I. Each alloy was annealed at 2200°F for 1 hr and air-quenched to comparable grain sizes (ASTM G.S. No. 4-5). The rods were then machined into test samples 0.160-in. diam and 0.750-in. gage length. Tension testing was carried out in an In-stron tension-testing machine at a strain rate of 0.02 in. per in. per min. An extensometer was used to determine the room-temperature yield stress. Stress-rupture testing was carried out in an Arc Weld stress-rupture unit. All testing was performed in air. EXPERIMENTAL RESULTS A) Room-Temperature Mechanical Properties. The 0.2 pct yield stress, ultimate tensile strength, and elongation for all alloys tested at room temperature are plotted as a function of atomic percent solute element in Fig. 1. The ultimate tensile strength for the same atomic percent increases in the order palladium, rhodium, ruthenium, osmium. The yield strength at low concentrations, -0.5 at. pct, increases in the order osmium, palladium, ruthenium, rhodium. At higher concentrations, 3.0 and 10.0 pct, the yield strength increases in the same order as the tensile strength. Both yield strength and tensile strength increase rapidly ini-

Jan 1, 1964

By J. L. Workman

The Putnam Coal Mine, at design capacity, will be the third largest underground bituminous coal mine in the world and will feature the world's longest single flight belt conveyor. Construction is now in process; initial mining will commence in early 1969 and design capacity will be reached in 1971. The Putman Coal Mine reserves are located in south-central West Virginia and contain approximately 27,000 acres of prime Pittsburgh No. 8 coal; reserves are sufficient to support a 3.5 million tpy operation for over 30 years. The mine portal area is located some six miles inland from the Kanawha River. The river loading facility is located approximately 156 miles by way of river from the market.

Jan 1, 1969

By Jun Hino, P. G. Shewmon, P. A. Beck

THE investigations of Crussard,' of Guinier and Tennevin,' and of Dunn and Daniels," indicate that the subgrains formed in a cold worked and annealed metal are capable of growing at each other's expense during annealing, if the temperature of annealing is sufficiently high and the time is long enough. The results of Dunn and Daniels are particularly convincing in showing that subgrain growth is essentially the result of the free surface energy associated with the subboundaries. In subgrain growth, as in ordinary grain growth, energy is gained as a result of the decrease in total subboundary surface area per unit volume. Recently, Wood and Scrutton4 found that the rate of subgrain growth upon annealing increased very considerably, if simultaneously a creep strain was applied to the specimen at a low strain rate. Working with 99.98 pct pure fine-grained aluminum, these investigators found that the continuous X-ray diffraction back-reflection rings of material strained at room temperature remained continuous after heating for ten days at 250 °C. However, when heated at the same temperature under a stress of 1000 psi even for only three days, the formerly continuous X-ray diffraction rings broke up into numerous dots, which were fairly clearly separated from each other. The continuous X-ray diffraction rings were interpreted as indications of a very small subgrain size, not resolved by the X-ray diffraction method used. The breaking up of the continuous X-ray diffraction ring during annealing under stress was taken as an indication of a great increase in subgrain size, so that the individual subgrains could then be resolved. The effect of simultaneous strain at a low strain rate in accelerating subgrain growth, discovered by Wood and Scrutton and designated by them as "cell growth," is of fundamental importance. The experiments described in this note were carried out in order to confirm Wood and Scrutton's results by direct metallographic observation. Also, information was sought as to the minimum creep strain necessary to produce this effect. A fine grained high purity aluminum strip was prepared by alternate 33 pct rolling and annealing treatments for 1 hr at 350°C. Specimens cut from this strip were subjected to a relatively fast creep strain of 7.2 pct in 3.5 min at 300°C under a constant load initially corresponding to 1185 psi. The subgrains set up were large enough (about 0.015 to 0.03 mm) to be clearly observed (Fig. 1) at X200 magnification with polarized light, after electrolytic polishing and anodic etching, producing a fine oxide film." The corresponding X-ray diffraction pattern is shown in Fig. la. After the fast creep strain treatment a portion of the specimen was subjected to a creep strain of 8.3 pct in 44 hr at 350°C under a constant load initially corresponding to 320 psi. Another portion, annealed under the same conditions, but not strained, served as control specimen. The subgrain size of a typical area of the specimen annealed under strain and of the one annealed without strain is shown in Figs. 2 and 3 (X-ray diffraction patterns, Figs. 2a and 3a). Annealing without strain produced clearly observable subgrain growth (subgrain size: about 0.03 to 0.05 mm). The effect of simultaneous strain was to increase greatly the rate of subgrain growth (resulting in a subgrain size of approximately 0.05 to 0.13 mm), in accordance with Wood and Scrutton. In another experiment, the effect of the amount of simultaneous strain was studied. The initial subgrain structure was set up by fast creep, as described previously. The specimen was then heated at 350 °C under a constant load initially corresponding to 320 psi, as above, for 2.2 hr (0.24 pct strain), and 8.8 hr (0.58 pct strain). Comparison with the corresponding unstrained control specimens by means of X-ray diffraction showed that 0.58 pct strain definitely had an effect, but the effect of 0.24 pct strain was doubtful.

Jan 1, 1953

By H. S. Geismer

INTRODUCTION. THERE are three general methods available for obtaining commercial brown iron-ore : hand-screening; washing; and washing and concentrating. Hand-screening has produced a large tonnage of ore in the past, but is rapidly falling into disuse, except as a preliminary step, largely because modern furnace-practice requires all foreign material to be separated from ores before they are delivered to the furnace, and hand-screened ores can rarely be made to fulfill that requirement. Another argument against hand-screening is, that the waste-pile often contains as much ore as the screen recovers. The operation known as hand-screening consists of throwing the ore-bearing material against a stationary inclined screen of 0.5- to 0.75-in. mesh, similar in construction to the screens employed in sand-pits. The recently-prospected brown-ore deposits of Cuba seem to offer possibilities for hand-screening not possessed by the average American deposit, since these ores are not associated with worthless clays. Washing, and washing and concentrating, may be discussed together, since concentrating is never practical on brown iron-ore, except to treat ores that have already been washed. (Strictly speaking, washing might be classed as concentrating, but it is never so designated by brown-ore men.) The various steps of removing ore-bearing material and converting it into ore suitable for blast-furnace use, may be grouped under five heads: loading for transportation ; transporting ; feeding into the washer; washing; and concentrating. These five subjects will be considered in consecutive order. Each of these successive steps may be accomplished in a number of different ways, the choice depending entirely upon the character of the ore-bearing material to be treated.

Aug 1, 1911

By W. H. Emmons

Most mineral deposits change as they are followed downward on their dips. Some of these changes are due to primary arrangement; different ores were precipitated at different depths when the deposits were formed. Other changes are due to weathering or secondary enrichment. The chief changes resulting from weathering, in the main, are understood and standard changes are recognized. The processes are relatively simple. They take place at ordinary temperatures and pressures and can be duplicated in the laboratories. Moreover, there are thousands of deposits available for study, because the processes take place near the surface where the development of mines is begun. The changes due to primary arrangement of the deposits of the metals are not so well understood. Deposition has taken place at higher temperatures and our knowledge of the chemistry of the processes leaves much to he desired. Very few experimental data are available. The deposits, moreover, offer fewer examples because many mines are not developed to depths where the changes can be studied and a change to unprofitable material usually results in stopping work. Most students of ore deposits now accept, as confirmed, the hypothesis that nearly all lodes and related deposits were formed by solutions originating in intrusive igneous masses. Outward from these masses, the deposits at many places are grouped in zones and the arrangement noted from the intrusions outward are commonly the reverse of the changes noted from the surface downward in a single lode. A study of the changes in groups of deposits, from the parent intrusion outward, will throw light on the changes to be expected in a single lode downward. Recause the changes in series of deposits from the intrusive outward may all be observed near the surface, where more data are available, it is possible to increase the number of examples many fold and to apply the statisttical method of study with greater confidence of avoiding error. It is not everywhere possible to identify the parent igneous body to which a group of ore deposits are related. Several kinds of criteria are available; certain groups are superior to others. By applying the

Jan 1, 1924

By G. B. Craig, H. E. Rosinger, W. J. Bratina

NUMEROUS studies have been reported on how the mechanical properties of a bcc metal are affected by interstitial solute atoms in the temperature range where appreciable diffusion occurs in finite time. Only recently, though, has work on an "interstitial-free" iron been reported, this by Leslie and sober,' on an Fe-0.15 pct Ti alloy. The data was limited to a maximum temperature of 296°K. This work presents a systematic study of the flow stress and the activation volume of a similar material as a function of plastic deformation over the temperature range 253" to 473°K. The material studied was Ferrovac E iron containing 0.01 wt pct C and 0.0002 wt pct N annealed at 870°C for 1 hr, and the same iron with an addition of 0.15 wt pct Ti annealed at 870°C for 80 hr; cooling rates being 75°C per hr. In the latter alloy depletion of interstitials in the a-iron matrix associated with the second phase formation was expected. The temperature dependence of the flow stress was measured at 6 = 3.33 x 10-5 sec-1 from 253" to 473°K as a function of the deformation in 0.01 intervals from 0.03 to 0.12 strain. Results for two strains, 0.03 and 0.09 are shown in Fig. 1. The flow stress increased and reached a maximum in the Ferrovac E iron with increasing temperature, but showed a consistent gradual decrease in the Ferrovac E-0.15 pct Ti alloy. A yield point and Luders region, observed in Ferrovac E, was absent in the Ferrovac E-0.15 pct Ti alloy. Also, the Ferrovac E exhibited a serrated stress-strain behavior in the temperature range 330" to 400°K; dynamic strain-aging characteristics were absent for the titanium alloy. The activation volume was determined over the same temperature range by the method of stress relaxation at constant strain. Results are shown in Fig. 2. At the temperatures of serrated stress-strain behavior in Ferrovac E the activation volume was found to increase toward infinity and, as the temperature rose, to decrease, to about 100 to 200 v/b3. By contrast, the activation volume for the Ferrovac E-0.15 pct Ti alloy tended to increase continuously from the initial value of approximately 300 at 300°K to more than 800 at 473°K. Above 400°K, then, the corresponding activation volume values of Ferrovac E were less than that of the Ferrovac E-0.15 pct Ti alloy. The activation volumes, Fig. 2, appear to depend on dislocation-interstitial interaction in Ferrovac E and on dislocation-dislocation interaction in Ferrovac E-0.15 pct Ti alloy. The gradual increase in the latter metal is associated with thermally activated motion of dislocation segments.' The decrease of the activation volume in Ferrovac E above 400°K is probably related to the diminishing effect of interstitials on dislocations. Here, the interstitials can be moved along with a dislocation without appreciable drag. As indicated in Fig. 1, removal of interstitials

Jan 1, 1970

By FRANK R. VAN

I. INTRODUCTION. IN the summer of 1908, R. B. Cochran, Superintendent of the Compania Metalurgica Mexicana at Sierra Mojada, Mexico, presented to the Department of Geology and Mineralogy at Case School of Applied Science an interesting series of silver-, copper-, and lead-ores from the Veta Rica mine. These ores contained several interesting minerals which were associated in a somewhat peculiar manner, so that I was impelled to make a closer study of their occurrence. The investigation has brought out several new facts concerning the geology of the district and the geological occurrence of the ores, and therefore has seemed worthy of record. Two papers have been published previously about the region, one by Chism1 in 1887, and the other by Malcolmson2 in 1902. These articles, however, treat of the district more from the stand-point of the metallurgist and mining engineer, and not much attention is devoted to its mineralogy and economic geology. II. GEOGRAPHICAL LOCATION OF THE DISTRICT. The Sierra Mojada is situated in the extreme western part of the State of Coahuila, near the Chihuahua boundary. It is reached best via the Mexican Central railroad from El Paso, Tex., to Escalon, Chihuahua, a distance of 494 miles. From this point, a branch line, the Mexican Northern railroad, runs 78 miles in a NE. direction to the mines, which are its northern terminus. The railroad was built about 1890, practically on account of the Sierra Mojada mining-district. 1 Sierra Mojada, Mexico, by Richard E. Chism, Trans., xv., 542 (1886-87). 2 The Sierra Mojada, Coahuila, Mexico, and Its Ore-Deposits, by James W. Malcomson, Tram., xxxii., 100 (1901).

Aug 1, 1912

By Kendall Y. Keuhey

The Mercur Mine, developed and operated by Getty Mining Company, is designed to process 3,000 tons/day of gold ore from the historic Mercur Hill mining district. The mill process involves crushing, grinding, and cyanide carbon-in-leach for recovery of gold from ores ranging from 0.08 to 0.10 oz/ton. During the mining of normal grade ore, a substantial tonnage of material containing 0.02 to 0.05 oz/ton is mined and would be available for heap leaching. Also, Getty Mining is developing methods to test and prepare lower grade material for heap leaching by rejecting a coarse rock fraction from the grinding SAG mill circuit. This lower grade fraction is suitable for heap leaching, and its removal will upgrade the regular mill feed. Other test methods are directed toward leaching techniques on some 4.0 million tons of tailings that resulted from roasting and/or cyanide leaching Mercur ores in the early 1900's. The natural carbons associated with the Mercur deposits require special treatment to reduce their gold adsorption effects when leached with cyanide.

Jan 1, 1984