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By G. L. Corrigan, J. R. Coleman

FOUR factors that largely determine the end product obtained when cement and water are mixed are the chemical composi-tion of the cement, the fineness to which the cement is ground, the amount of mixing water used, and the procedure used in mixing the slurry. In order to study the effect of two of these factors, the other two, the chemical composition and method of mixing, are kept constant, leaving as variables fineness and water-cement ratio. Through years of experimentation and usage it has been well established that the reduction of cement clinker to a very fine state by grinding improves its cementing value, principally because the reaction between cement and water takes place only at the surface of the cement particle, further action being hindered by the accumulation of reaction products that tend to coat the unreacted material. Therefore, the moE finety-ground %cement, and the greater the surface exposed in propor-tion to its mass, the more rapid is the The work covered by this paper was done by the authors in pursuance of graduate studies in the ~epartment of petroleum Engineering at The University of Texas. The present paper is a condensation of a more detailed DaDer from which, for brevity, information such as the procedure for determining permeability and some of the discussion and procedure relevant to slurry volumes and fineness of cement have bsen deleted. It was felt that this information would not be of interest to most readers and should be outside the scope of this paper. Manuscript received at the office of the Institute Aug. 16, 1940. Issued as T.P. 1266 in Petroleum Technology. January 1941. hydration rate. Similarly, the greater the surface exposed, the greater the proportion of the cement that reacts, and the lower the amount of residue that is unreacted and so remains inert and unable to make any contribution to cementing properties.] Although it is more expensive to manufacture a finely ground cement than a coarser cement, the advantages of fine grinding of construction cements were so marked that the fineness of cement has increased considerably in past years. When the manufacture of special oil-well cements was begun, it was necessary to consider properties of cement that previously were considered of minor importance, foremost of these being pumpability. Since, with the water-cement ratio constant, a finely ground cement will produce a neat slurry of higher consistency and shorter time of pumpability than will a coarser cement, there is a tendency to grind oil-well cements to a lower degree of fineness. Therefore, in order to produce oil-well cements that will make a slurry having a long time of pumpability, cements are manufactured that because of their coarseness generally are Poorer cementing materials than some of the more finely ground construction and on the other hand, often it is necessary to Pump oil-well cements for long periods of time to place them properly behind a string of casing in a deep well, and since high strength is not of foremost importance for such cement, there is room for considerable controversy concening the proper fineness range.

Jan 1, 1941

By H. L. Washburn, W. A. McConnell

There is not a coal preparation engineer alive who would equip his newest and most modern preparation plant with all new and untried pieces of equipment. He may install a few such pieces in which he has great confidence, but he will also spend much time assuring himself that everything else in the plant has been subjected to the test of time. He may arrange these units in a unique manner, but in general, they must have proven themselves elsewhere before he will approve them for his plant. A modern coal preparation plant contains only a few items of processing equipment, the remaining items being devices for materials handling. Consequently, pure automation is almost impossible. Although the main aim of automation is conservation of manpower, it generally results in uniformity of product, lower maintenance costs, and the many intangibles that make a better plant.

Jan 8, 1961

By E. H. Greener, M. E. Fine

The internal friction (1/Q) of cobalt Fig. 1 (measured by an electrostatic dynamic method1) near 250°C begins to increase rapidly on heating and continues to increase until 560°C, the highest tem- perature measured. In the plot of log of 1/Q against 1/T, above 250°C the data nearly fit two straight lines intersecting at 400°C, approximately the temperature of the hexagonal-cubic transformation. The higher temperature branch corresponds to an effective activation energy of 6500 cal per mole, the lower temperature branch 11,000 cal per mole. Young's modulus Fig. 2 increases approximately

Jan 1, 1959

By J. Eric Belford

The Kidd Creek orebody is a massive base metal sulphide deposit with surface dimensions of 168 m by 670 m. The steeply dipping ore- body has been evaluated to a depth of 1 524 m. To date, the mining of the orebody has been divided into three stages. The first stage of mining was an open pit operation which supplied ore from 1966 until 1977 when a depth of 219 m was reached at the widest or north end of the orebody. The second stage of mining, called No.1 Mine, was the development of an underground mine to replace the pit and recover the ore from the floor of the pit to a depth of 792 m. Development of the underground mine started in 1969. Full production of 273 000 tonnes per month was attained in 1977. The third stage of mining called No.2 Mine, was initiated to accommodate a production expansion and included the development of a second underground mining operation to recover the ore from the bottom of No.1 Mine, i.e. the 2600 level; to a depth as yet undetermined but most likely the 4600 foot level. Development of the No.2 Mine began in 1974. Full production of 91 000 tonnes per month will be achieved in 1982. The choice of the 2800 level as the partition between the two underground mines was based on the following factors: The 2600 level is the lowest level which can conveniently be mined through the No.1 Mine crusher on the 2800 level. Development and trial stoping methods for mining at depth could be evaluated in the 2600 to 2800 block utilizing the existing hoisting facilities of No.1 Mine. Early development of the ore in the 2600 to 2800 block would provide additional mill feed sooner than could be achieved by developing from No.2 Shaft. The underground mining plan then is to mine from the pit floor down to the 2600 foot level and at the same time mine from the 2800 level down to at least the 4600 foot level. PLANNING The Kidd Creek operations are designed to follow a logical mining sequence. The grade of ore supplied to the concentrator is the result of following the mining plan. The philosophy of mining has been to mine downward in horizontal slices. In the case of No.1 Mine the slices are 122 m thick while in No.2 Mine the slices will be 183 m thick. The point of this approach is to mine the pillars immediately after creating them rather than allowing them to stand, deteriorate and generate uncontrollable situations. Two types of ore are handled separately to optimize recoverable values in concentrates. The 'A' type ore is a copper zinc ore with some silver and little lead. The 'C' type ore is a zinc silver ore with some lead and copper

Jan 1, 1981

By H. H. Smith, G. J. Danek, M. R. Achter

Nickel, Type 316 stainless steel, and copper show decreases in reversed bending fatigue strength with increases in air and oxygen pressure. The leveling off of the log N us log p curve for nickel at 300" and 816°C is discussed in terms of a model in which surface saturation gives rise to a critical pressure. A metallographic study indicated that the effect of environment on crack initiation and growth rates is a function of temperature. It is shown that small amounts of impurities in inert gases may substantially reduce the fatigue strength of nickel. 1n a previous investigation1 of high-temperature bending fatigue in air and in vacuum, reversals in the effect of environment with changes in test conditions were atrributed to a mechanism consisting of two processes acting in competition. One of them, oxidation strengthening, has been demonstrated in a straightforward manner by measuring the rupture strength of a sintered nickel-oxide bond.' The detailed operation of the other process, facilitation of crack propagation by surface absorption of oxygen, is uncertain. Thompson et UZ advanced the possibility that the formation of an oxide layer prevented rewelding of cracks during the compressive part of the cycle, while Wads worth and Hutchings4 further proposed that the reaction of oxygen with atoms at the tip of the crack helped the metal to break there, a possibility which has also been advanced to explain crack formation in creep.5 A difficulty in arriving at a mechanism for this lowering of fatigue strength in air is the inconsistency of some of the data. For example, Wadsworth and Hutchings4 found an approximately linear decrease of log cycles-to-fracture (N) with an increase of log pressure-of-air (p) for copper and aluminum at room temperature, while snowden,' although finding a similar linear dependence for aluminum, observed a stepped curve for lead. Between l0-' and 10"1 torr there is a sharp decrease in the fatigue life of lead with increasing air pressure, but plateaus, signifying no dependence on pressure, are found at higher and lower pressures. If, as discussed elsewhere: these environmental effects arise from the adsorption of gas on crack surfaces, it would be expected that as the pressure is increased a saturation point would be observed, beyond which there is no further effect. To gain further information about the form of the log N vs logp curve, it was decided in the present investigation to repeat this type of measurement at elevated temperature on nickel, copper, and Type 316 stainless steel. An additional object of this program was

Jan 1, 1963

By James R. I. Henagan, William O. Crego

The Mamou Field, located in Evangeline Parish, La., is an elongated anticlinal structure on the downthrown side of a major east-west fault with oil and gas production from the upper part of the Wilcox formation. Two main producing horizons occur at an approximate depth of 11,500 ft and have been designated the Morein sand and the Deshotels sand. The upper sand (Morein) had an original bottom-hole pressure of 6,788 psig, which is some 1,500 psi above the normal pressure for that depth. Oil production with high gas/oil ratios and fluid analyses giving high shrinkage factors prompted a study of the reservoir in its early stages. The results indicated a closed reservoir producing under solution gas drive with low expected ultimate oil recovery. The early study of the Wilcox formation and its contained hydrocarbons resulted in the presentation of the data to the Department of Conservation of the State of Louisiana. with suggestions for field rules and a request for permission to inject extraneous water into the Morein sand. The water injection plans eventually led to complete unitization of working and royalty interests, to the construction of a gasoline plant, and to a central tank battery installation for the field. Deshotels sand wells were reworked to the Morein sand, and the water injection system was installed and ready for continuous operation in the early part of 1949. The injection of water was started in time to arrest the bottom-hole pressure decline at or near the bubble-point pressure. This injection was calculated to increase the ultimate recovery and to conserve natural resources. Pressure maintenance, the extraction of LPG products, and the sale of residue gas (early 1950) represent an attempt to produce an oil field with utmost efficiency. INTRODUCTION The two principal Wilcox oil horizons in the Mamou Field are the Morein sand and the Deshotels sand, occurring at 11,500 ft and 11,700 ft subsea, respectively. Minor oil production has been encountered in Morein stringer sands. (Fig. 1.) This paper deals only with the Morein sand, which is blanket over the Mamou Field. It is, however, segmented by the northwest-southeast "E" fault as shown in Fig. 2. The west segment is the area involved in the water injection pressure maintenance program. To date the east segment has operated largely under the influence of a strong water drive. The paper presents (1) a brief history of field development, (2) a resumé of reservoir conditions which led to the decision to inject extraneous salt water into the principal producing horizon, and (3) a summary of reservoir analysis and performance. FIELD DEVELOPMENT The Mamou Field is located in Evangeline Parish, La. The discovery well (J. B. Morein No. 1, now Unit No. 5) was completed in December, 1945, after penetrating 200 ft into the Wilcox formation. Initial production from the Morein

Jan 1, 1951

By James R. I. Henagan, William O. Crego

The Mamou Field, located in Evangeline Parish, La., is an elongated anticlinal structure on the downthrown side of a major east-west fault with oil and gas production from the upper part of the Wilcox formation. Two main producing horizons occur at an approximate depth of 11,500 ft and have been designated the Morein sand and the Deshotels sand. The upper sand (Morein) had an original bottom-hole pressure of 6,788 psig, which is some 1,500 psi above the normal pressure for that depth. Oil production with high gas/oil ratios and fluid analyses giving high shrinkage factors prompted a study of the reservoir in its early stages. The results indicated a closed reservoir producing under solution gas drive with low expected ultimate oil recovery. The early study of the Wilcox formation and its contained hydrocarbons resulted in the presentation of the data to the Department of Conservation of the State of Louisiana. with suggestions for field rules and a request for permission to inject extraneous water into the Morein sand. The water injection plans eventually led to complete unitization of working and royalty interests, to the construction of a gasoline plant, and to a central tank battery installation for the field. Deshotels sand wells were reworked to the Morein sand, and the water injection system was installed and ready for continuous operation in the early part of 1949. The injection of water was started in time to arrest the bottom-hole pressure decline at or near the bubble-point pressure. This injection was calculated to increase the ultimate recovery and to conserve natural resources. Pressure maintenance, the extraction of LPG products, and the sale of residue gas (early 1950) represent an attempt to produce an oil field with utmost efficiency. INTRODUCTION The two principal Wilcox oil horizons in the Mamou Field are the Morein sand and the Deshotels sand, occurring at 11,500 ft and 11,700 ft subsea, respectively. Minor oil production has been encountered in Morein stringer sands. (Fig. 1.) This paper deals only with the Morein sand, which is blanket over the Mamou Field. It is, however, segmented by the northwest-southeast "E" fault as shown in Fig. 2. The west segment is the area involved in the water injection pressure maintenance program. To date the east segment has operated largely under the influence of a strong water drive. The paper presents (1) a brief history of field development, (2) a resumé of reservoir conditions which led to the decision to inject extraneous salt water into the principal producing horizon, and (3) a summary of reservoir analysis and performance. FIELD DEVELOPMENT The Mamou Field is located in Evangeline Parish, La. The discovery well (J. B. Morein No. 1, now Unit No. 5) was completed in December, 1945, after penetrating 200 ft into the Wilcox formation. Initial production from the Morein

Jan 1, 1951

IN A broad sense, the subject of slag control includes not only the adjustment of the composition of the slag but also of its relative weight in terms of percentage of the metal-bath weight. The slag is made up essentially of (1) oxides formed by oxidation of silicon, manganese, phosphorus, and iron in the charge, (2) SiO2, MnO, and other oxides charged as such, and (3) lime charged or fed to the heat to neutralize the “acidic”: oxides present (Si02 and P2O5). Obviously, if practically perfect control were possible over the composition of the charge as well as over the oxidation process during the melting period, both slag weight and composition would be very nearly predetermined, and the control problem would be solved. Unfortunately, we have only a somewhat. imperfect control over these factors, so that beginning at about the lime-boil period in the heat, and later in the early refining period, it becomes necessary for the first helper, as the heat progresses, to make quick estimates of the composition and physical state of development of the slag (as well as of the amounts of phosphorus, sulphur, and carbon in the metal) from which he can judge how much extra lime, roll scale, fluorspar, and other additions will be needed not only to give the desired final slag composition, but also to hasten the complete fusion or shaping to a desired viscosity in the liquid state. The discussion in this chapter is limited essentially to those more or less practical rapid methods of slag estimation that have been used, or at least tried, for these purposes under actual open-hearth shop conditions. Another aspect of importance here is the extent to which the freedom of choice to the operator is limited by available charge materials, product specifications, and the oxidation and other chemical- reaction effects which inevitably occur under the conditions of this furnace process. For a better understanding of these control limitations and of the chemical and physical background of the subject, the reader is urged to read at least part of the discussion in Chapters XVII and XIX. Charge materials are discussed in Chapter IV, and many of the details of charging, melting, and refining practice closely related to slag control are given in Chapters VI and VII.

Jan 1, 1944

By T. S. Spice, H. L. Lovell, R. W. Utley

With the increased demand for lightweight aggregate, such materials have been manufactured from slags, clays, slates and, to a minor extent, the refuse of coal preparation processes. The latter source is attractive due to current interest in eliminating air pollution and improving the unsightly appearance of refuse piles as well as potential economic advantages, if a satisfactory product can be manufactured. A study has been made relating the composition of coal refuse, expressed as washability data, and the properties of a resulting aggregate. Three different refuse samples (bituminous and anthracite) were separated at gravities ranging from 1.4 to 2.96. Their chemical composition was determined. Aggregates were prepared in the laboratory from these fractions before and after ashing, at temperatures between 1250' and 1350°C. The specific gravity, expansion and compressive strength of the products were measured. The response of the refuse to bloat is related to the chemical composition of the ash to yield a glass of desirable viscosity and to the presence of a material which will form a gas in the appropriate temperature range. Carbonaceous material within the particle is a negative factor since after the glassy phase forms, oxygen cannot enter the mass thus inhibiting bloating. Precombustion of the refuse is also detrimental since it removes substances of gas-forming capability. The fraction sinking at 2.60 yielded an aggregate which had the lowest specific gravity, greatest expansion and greatest compressive strength. The purpose of this study was to seek and evaluate possible relationships between the composition of coal refuse, as expressed by washability data and factors affecting the production of a lightweight aggregate from refuse by bloating. The study was related to an intense program in the Department of Mineral Preparation at the Pennsyl- vania State University to study the possible elimination of air pollution from coal refuse fires by producing a low-carbon secondary refuse. Another phase of this program is considering possible uses of such a secondary refuse. Thus one study was to determine if a lightweight aggregate could be produced from this new refuse material and what effect the nature of refuse beneficiation would have upon the characteristics of a commercial aggregate product. It was hoped that these studies would assist in the design of a process and in establishing the optimum specific gravities at which refuse separation could be made in respect to the production of a low grade fuel from the float product and a raw material for lightweight aggregate from the sink product. The combination of producing a saleable low grade fuel and a lightweight aggregate, while at the same time minimizing disposal costs and air pollution contributions would be attractive to the coal mining industry. MECHANISM OF BLOATING Austin, Nunes and Sullivan,1 Ehlers,3 Everhart, 4 Riley7 and others, have stated that bloating is an expansion process in which two essential conditions exist simultaneously. These are: the development of a glassy phase over a relatively wide temperature range (the viscosity of this glassy phase plays a major role in the retention of the evolved gases) and the evolution of gases (SO2, SO3, CO2 and O2) from mineral decomposition in sufficient quantities at appropriate temperatures to form a cellular structure in the mass. Unless these two "type" reactions take place at the same time or in overlapping ranges of time and temperature, cellulation will not occur. In this event, the gas will escape or the melt will be too fluid to retain it; or if retained, the cells may be too large or the walls too thin for good strength in the product. The stages that occur when clays and shales are heated in an oxidizing atmosphere may be summarized as follows: 1) Drying and removal of free water to 200°C. 2) Absorbed water dehydration period (200 to 480°C). 3) Chemically-combined water dehydration period (480 to 700°C).

Jan 1, 1965

By Benny Langston, Frank M. Jr. Stephens

IN recent years much attention has been given to recovery, treatment, and disposal of dusts discharged into the atmosphere from operations of industry. Considerable data has been accumulated on both operational1-4 and design5,6 of dust-collector equipment for commercial installations. On the other hand, there is almost no published data on design and construction of small-scale equipment to handle dust problems that arise in the ore-dressing laboratory. Dust-collection equipment such as multiclones, single-cyclones, scrubbers, chemical and mechanical filters, settling chambers, and electrostatic separators has proved its efficiency for collecting dust in industry. In the laboratory, however, the engineer is faced with the problem of collecting small quantities of dust, inexpensively, without diverting the major effort from the metallurgical problem to the problem of collecting dust produced by the process.

Jan 8, 1954

By Hadyn W. J. Wren

The practical experience obtained from designing and commissioning of fully controlled coal preparation plants, using centralised mini computers is discussed. The importance of defining a workable operating philosophy is highlighted with particular reference to the problems associated with trying to achieve a total 'hands off' concept. These experiences with a centralised computer controlled system have led to a deeper understanding of the interaction between the operation and the control system and the role of local control and management information within the overall control philosophy of a highly automated process plant. The utilisation of an extensive fully distributed microprocessor control system with over 6500 digital and analogue signals split into six autonomous areas with a supervisory system to provide operator interface is described. The important aspects of project definition and design needs are outlined with particular regard for achieving maximum efficiency with minimum complexity. Views of future trends are offered, based on the experience of both centralised computer and distributed processor systems.

Jan 1, 1984

By C. L. Clark, J. Russ

FOR many years attempts have been made to develop a laboratory test that would serve to indicate the proper temperatures to be used in the various hot-working applications to which metals may be subjected as they are being processed from the cast ingot to the desired finished products. That these attempts have not been very successful is indicated by the fact that most of the temperatures in use today in these operations have been developed over a period of years by trial and error. In other words, the selection of the proper temperatures for the hot-working of metals is generally an art rather than a science. Our knowledge with respect to the behavior of metals at elevated temperatures has been greatly extended during the past several years. While most of the work in this field has been done at proposed operating temperatures of equipment, rather than at processing or fabricating temperatures, one outstanding fact has been developed, which is applicable to the entire high-temperature range; that is, that the rate of deformation to which the metal is subjected is of the same order of importance as the temperature. The failure to properly recognize this fact has greatly retarded the development of a suitable laboratory test for determining hot-working temperatures. INFLUENCE OF RATE OF DEFORMATION ON HIGH-TEMPERATURE PROPERTIES It is not the intent of this paper to survey all the work that has been done on the effect of deformation rate on the properties of metals. The importance of this factor, however, was shown by Zay Jeffries1 as early as 1919, when he called attention to the existence of the so-called equicohesive temperature and to the effect of certain factors on the location of this temperature. Fig. I is a diagrammatic sketch showing the effect of temperature on the relative strength of the crystals and grain boundaries and the type of fracture that will result, depending on the location of the test, or working temperature with respect to the equicohesive temperature. Since the type fracture varies, it also follows that at temperatures below the equicohesive temperature the major deformation occurs in the grains; while at temperatures above, the deformation results either from the movement of the grains with respect to each other or recrystallization of the strained material occurs almost instantaneously, producing new crystals at the fracture. As previously indicated, it is an established fact that the position of this equicohesive temperature for any given metal is dependent on the deformation rate. For example, in 0.15 per cent plain carbon steel the location of this temperature under

Jan 1, 1945

By Neil O. Johnson

In the cyanidation of gold, silver and mixed ores, the solids, suspended in the pulp after fine grinding, primary settling and scries agitation arc subjected to a washing step known, generally, as countercurrent decanta-tion, abbreviated C.C.D The conventional C.C.D. plant consists of a series of separate thickeners—three to five in number—arranged in series on slightly different elevations. Solids settled in the first thickener are pumped at maximum density into the second thickener, where they arc diluted with weak solution overflowing the third. After settling in the second thickener they are pumped to the third, where the process of repulping with still weaker solution is repeated, and so on throughout the system, until they reach the last thickener and are repulped with barren solution or water. Recently there has been developed a new type of thickener, known as the Dorr Washing Tray Thickener, consisting of up to five individual thickeners, which performs in a single unit all the functions of a conven-tional C.C.D plant. Fig. 1 illustrates the essential features. It is a multicompartment tray thickener, in which the trays operate in scries instead of in parallel for countercurrent washing. Feed enters the top compartment via a central, semisubmerged feed well. Strong solution overflows into a peripheral launder. Solids settle to the bottom, are raked to the center and flow into the top of the second compartment, directly below, via an inverted cup and scal. Solids pass downward, compartment by compartment, against a counter flow of wash solution and finally of fresh water. Thus, in accord-:trice with the basic C.C.D. principle, the solids transfer the dissolved metal content to the wash solution and finally are removed by a diaphragm pump from the bottom compartment virtually free of valuable material. Similarly, the wash water, as it flows through the system, becomes increasingly enriched in dissolved metal until it overflows from compartment two and returns to the process.

Jan 1, 1939

By Neil O. Johnson

In the cyanidation of gold, silver and mixed ores, the solids, suspended in the pulp after fine grinding, primary settling and scries agitation arc subjected to a washing step known, generally, as countercurrent decanta-tion, abbreviated C.C.D The conventional C.C.D. plant consists of a series of separate thickeners—three to five in number—arranged in series on slightly different elevations. Solids settled in the first thickener are pumped at maximum density into the second thickener, where they arc diluted with weak solution overflowing the third. After settling in the second thickener they are pumped to the third, where the process of repulping with still weaker solution is repeated, and so on throughout the system, until they reach the last thickener and are repulped with barren solution or water. Recently there has been developed a new type of thickener, known as the Dorr Washing Tray Thickener, consisting of up to five individual thickeners, which performs in a single unit all the functions of a conven-tional C.C.D plant. Fig. 1 illustrates the essential features. It is a multicompartment tray thickener, in which the trays operate in scries instead of in parallel for countercurrent washing. Feed enters the top compartment via a central, semisubmerged feed well. Strong solution overflows into a peripheral launder. Solids settle to the bottom, are raked to the center and flow into the top of the second compartment, directly below, via an inverted cup and scal. Solids pass downward, compartment by compartment, against a counter flow of wash solution and finally of fresh water. Thus, in accord-:trice with the basic C.C.D. principle, the solids transfer the dissolved metal content to the wash solution and finally are removed by a diaphragm pump from the bottom compartment virtually free of valuable material. Similarly, the wash water, as it flows through the system, becomes increasingly enriched in dissolved metal until it overflows from compartment two and returns to the process.

Jan 1, 1939

By K. G. Wikle, J. G. Klein, W. W. Beaver

Consolidation of hydride process, electrolytic, calcium reduced, and comminuted thorium powder, as well as saw chips and lathe turnings, by vacuum hot pressing and by cold pressing-vacuum sintering was studied. The mechanical properties of the consolidated material in the extruded form are compared with those of wrought castings. AT present there little little industrial use for thorium metal, although it has some important though small scale applications in electronic equipment. Despite its high inelting point—about 1750°C —a low modulus of elasticity, 11.4xl0 si at 20°C;' relatively low mechanical properties coupled with a high density, 11.7 g per cu cm; and an unusually high chemical activity with normal atmospheres limit any structural applications. The metal is utilized as an alloying element principally in magnesium. Pure thorium finds utility as electrodes in gaseous discharge lamps such as the high intensity mercury lamp' because its low work function and high electron emissivity provide lower starting potentials and more uniform operating characteristics than other available materials. The metal is also found in photoelectric tubes used for the measurement of the ultraviolet spectrum." Thorium metal has been used in germicidal lamps of the cold cathode type as sputtered coatings on nickel in order to provide a low work function surface and a low starting voltage. Other applications have involved the radioactive properties of thorium for the production of ionized particles." The potential value of thorium is much greater than its present use pattern because of possible utility in the field of nuclear power. Th may be converted through nuclear reaction to a fissionable element U which should be capable of acting similarly to U in the g'eneration of atomic power. Thorium has been reported to be about three times as plentiful as uranium in the earth's crust, placing it in the order of abundance of lead and molybdenum." Thus, it is of interest in augmenting the potential supply of fissionable material for nuclear power. Because of its high melting point, thorium is usually produced as a powder through the calcium reduction of its oxide or thermal reduction of halides by sodium, magnesium, and calcium. It may also be produced in flake form by electrolysis of fused alkali or alkaline earth chloricles and fluorides. Therefore, powder metallurgy assumes importance in the fab- rication of thorium metal shapes. Furthermore, it is rather difficult to obtain pure thorium by melting, as the molten metal reacts readily with graphite as well as oxide, carbide, and nitride refractories. These contaminate the melt with oxides, carbides, and metallic impurities." The current investigation was undertaken to examine the fabrication of thorium by powder metallurgy methods which have been used for the commercial production of beryllium and other metals.' A sparcity of data concerning the comparative cold and hot compaction of thorium powders of different derivation existed. Therefore, all commercially available types were examined along with other experimentally produced thorium powders in order to round out the comparison of consolidated thorium powders with melted reguline metal. Review of the Literature By heating a mixture of ThC1, with potassium, Berzelius made the first thorium metal as an impure powder in 1828. Improvements in the basic process, increasing thorium assay to 99 pct, were made by several investigators including Arsem," Lely and Hamberger10 and Von Bolton." Calcium reduction of Tho, to make powders was investigated by Berger," Huppertz,'" Kroll," and Kuzel and Wedekind.'" A thorium powder produced by this method using a CaC1, fluxing agent assayed 99.7 pct, as reported by Marden and Rentschler.'" Compacted and sintered, this product was found to be ductile, and could be fabricated into wire and sheet. Improvements of the calcium reduction process were made later" wherein CaCl, was eliminated from the reaction, producing metal assaying 99.8 pct Th. Further work by Lilliendah118 howed that a coarser metal could be obtained by the substitution of ThC1, or ThOC1, for oxide with consequent advantage of stability to atmospheric reaction. Reports on the technology of thorium developed in Germany during World War II have been made by Espe."' Thorium powder of 99.5 pct Th was obtained by reduction of the oxide by calcium. Screening to —200 mesh, compacting with about 20 tsi, and sintering in vacuo at 1320" to 1360°C for 3 hr resulted in a porous sinter cake. The sinter cake was sufficiently ductile to be worked into bar, wire, and sheet which could be employed as electrode materials.

Jan 1, 1957

By A. Arrott, G. S. Ansell

A model is proposed to account for the observed strengthening behazlior of ferrous martensites. me model is based upon the inheritance of carbon-rich regions by the martensite which were present in the FERROUS martensite is formed by a nucleation and shear mechanism when austenite is cooled, at a rate which is fast enough to suppress the formation of pearlite, below a critical temperature (Ms ). Although the mechanism of the transformation process appears to be well understood,1,2 the mechanism re- austenite prior to transformation. These carbon-rich regions strengthen the martensite in a manner analogous to the strengthening effect due to a finely dispersed second phase. sponsible for the observed high strengths exhibited by ferrous martensites is uncertain. In this paper, a model is proposed to account for the strengthening of ferrous martensite. MODEL The maximum solubility of carbon in austenite is 1.98 wt pct as compared with a maximum solubility of 0.025 wt pct in ferrite. Prior to transformation, the carbon in the austenite is nonuniformly dis-

Jan 1, 1963

By Eugene A. Thiers, William V. Morris

INTRODUCTION Direct reduction processes reduce the various commercial forms of iron oxide (pellets, concentrate, fines, etc.) to metallic iron at temperatures lower than that of molten iron. Thus, this technology includes practically, all iron reduction processes other than blast furnaces and electric pig iron furnaces (whose output in terms of world production is negligible). The product of these processes which is known as direct reduced iron (DRI), or sponge iron, is primarily used as a source of metallic iron in steel-making operations. Interest in DRI, which has been significant since the early 1960s, increased significantly in recent years with the rapid growth of DRI installed capacity throughout the world. The importance of the subject for the Circum-Pacific region stems directly from the influence that DRI has on iron ore consumption and on future steel development for this region. Although there is widespread agreement that electric furnaces will continue to increase their share of global steel output, and especially so in the countries of the Pacific Steel community, some doubts exist about future scrap supplies being adequate to support growth at past rates. The authors believe that such doubts are soundly based. As this paper points out, the total supply of all metallics used in electric furnaces may not be adequate to support the extrapolated rapid growth in electric furnace steel production. This paper seeks to provide perspective on the global and Circum-Pacific prospects for DRI in light of recent energy price developments and the current recession. In this regard, the demand for DRI within the context of recent evolutionary patterns in steel-making, the outlook of DRI supply in terms of prevailing production costs, and the prospects of new technology are discussed. THE DEMAND FOR DRI Although several reports published in the last 10 years predict high rates of growth in DRI, the subject remains a controversial one. Significant growth has indeed occurred, but not to the extent anticipated in the studies summarized in Table 1. The substantial difference between previous expectations and present reality can be ascribed primarily to: (1) lower growth in steel production than formerly anticipated; (2) numerous cancellations of DRI facilities that were previously announced; and (3) a fundamental and probably irreversible change in the economics of DRI production. Note that DRI capacity at the end of 1980 was about 16 million tonnes, a significantly lower figure than any of the projections above. In addition, DRI production was only about half of capacity, reflecting the abnormally low rates of capacity utilization in this industry. [ ] Before examining the current outlook in steel, it is pertinent to note that the market for DRI is usually different in the industrialized countries of the West from that in developing countries. In the former, the available infrastructure and industry's diversification extends DRI's potential markets to numerous steelmaking, foundry, and other industrial applications, although competition from scrap and other forms of metallic iron is constant. Scrap is generally available in these countries and, therefore, DRI competes with it in electric furnace steelmaking, basic oxygen steelmaking (as a coolant), cupola foundry operations, or as an additive in the metallic charge for open hearth and blast furnaces. On the other hand, DRI in developing countries is often allocated exclusively to domestic electric furnace steelmaking or, when capacity exceeds domestic captive requirements, to export. Notwithstanding quality considerations, DRI is being and is likely to continue to be used predominantly as a source of metallics in iron and steel-making. Other uses of DRI, such as in copper cementation represent a marginal market in terms of overall tonnage and can be ignored at this point. Therefore, DRI demand is-determined by the overall availability of metallic scrap in its various forms--a function of steel production and its probable evolution. The Global Steel Outlook Given the present recession, an objective appraisal of the long-term outlook for steel is particularly difficult. On the one hand, historical trends and, especially, the inertial forces associated with such a basic industry as steel must be recognized; such trends suggest that the current stagnation with

Jan 1, 1982

By Charles H. Jacoby, Leo E. Read

IN diamond coring salt beds to evaluate deposits, special techniques are applied to standard slim hole drilling to obtain a representative sample of the water soluble sodium chloride. Industrial consumers generally require a run-of-the-mine material of 98.0 pct sodium chloride. Since no commercial beneficiation has yet been devised to upgrade the lower quality salt beds, the only processing is crushing and screening to the size required by the customer. Coring salt sections has recently become important not only in dry mining, but also in the development of brine cavities by hydraulic fracturing and the formation of jugs for storing liquid petroleum products. Drilling Fluid: The fluid used in coring salt beds in New York, Ohio, and Michigan is a thoroughly saturated brine made from 3.0 lb rock salt per gal of make-up water. This mined rock salt, -4 mesh +10 mesh, is dissolved in a lixator, a double-bottomed cone hopper that turns fresh water to fully saturated brine.

Jan 5, 1957

By George D. Dub

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

Jan 1, 1947

By George D. Dub

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

Jan 1, 1948