Search Documents

By E. R. Morgan

A series of manganese-rich ferromagnetic alloys has been studied. The alloys are based upon an interstitial solution of carbon in manganese. In each of the systems investigated, the ferromagnetic phase was found to be face-centered-cubic in structure. This magnetic phase is based upon the composition (Mn, X)4C; where X is a metallic element which has both a positive size factor with respect to manganese and a high positive valence. Typical elements of this type are aluminum, indium, and tin. Preliminary magnetic measurements indicate that the effective magnetic moment of manganese in the alloys is at least 1.0 Bohr magneton per atom. It is emphasized that the critical atomic separation beyond which ferromagnetism can occur will be governed by the electronic configuration of the manganese atoms. Lattice parameter measurements show that the ferromagnetic condition in the present alloys is associated with an increased atomic separation of only 3 to 6 pct compared with a hypothetical face-centered-cubic lattice of pure manganese. MANGANESE which is paramagnetic and not ferromagnetic in the pure state can become ferromagnetic if suitably alloyed. Under favorable conditions manganese can have an effective ferromagnetic moment of the order of 3.41,2 or even 4.0 compared with a moment of 2.2 Bohr magnetons for iron. The incidence of ferromagnetism is believed to be controlled by the degree of atomic separation.4,5 This is consistent with the fact that the atomic separation, when considered in terms of the ratio of the atomic diameter R to the diameter r of the 3d-shell, is larger in the ferromagnetic metals. The values of R/r calculated by Slater,' for metals of the iron group are given in Table I, and Slater has suggested that the critical ratio of R/r is 1.5. Consideration of the criterion of atomic separation alone would suggest that there are three classes of possible ferromagnetic manganese alloys: 1— Ordered structures, of either solution or compound types, such that manganese atoms are not near neighbors; 2—Structures in which the separation of near-neighbor manganese atoms is sufficiently increased by the presence of suitable solute atoms; and 3—Dilute solutions of manganese. Typical ferro- magnetic alloys of types 1 and 2 are the Heusler alloys and Mn4N, respectively. There are no recognized ferromagnetic alloys of type 3. In view of the proximity of the ratio R/r for manganese to the critical ratio of 1.5, the present work was concentrated on manganese-rich alloys rather than on ferromagnetic alloys of low manganese content, several examples of which are already known. Calculation of the distance of closest approach of manganese atoms in each of its allotropic forms, Table 11, indicates that the high temperature ? phase provides the most suitable basis for ferromagnetic alloys of high manganese content. However the high temperature forms of pure manganese cannot be retained on quenching, although a face-centered-tetragonal form apparently can be produced by electrodeposition. Since the critical separation which must be exceeded in order to obtain ferromagnetism is dependent upon the number of electrons in the d-shell, no exact critical value of R/r can be assumed. Not only will this separation vary from one pure metal to another, but it will also vary with the valency of the given type of atom when it is in an alloy. Raynor8 has shown that metals of the iron group probably

Jan 1, 1955

By R. W. Smith, R. J. Charles

Fractions of glass particles in the size range 0.5 to 5.0 µ were prepared by an elutriator that operates in a centrifugal field. Although mean sizes of commercially graded abrasive powders were ten times greater than those of samples prepared by the elutriator, frequency size plots for elutriator samples compared favorably. Efficiency of separation was low and amounts of material recovered by the elutriator were small, but the method should find use in preparation of small amounts of sized materials for laboratory investigations.

Aug 1, 1956

By E. C. Tveter, R. B. Tippin

The separation of minerals by heavy liquids is a standard laboratory technique which goes back at least 50 years, but commercially economic application of this principal to ore concenfration has been prevented by problems of liquid recovery. Recent investigations which demonstrate the different types of liquid-solid systems on various ores using the two major heavy liquids - methylene bromide and tetra-bromoelhane — are discussed in this article. As a result of these studies, over-all liquid recovery, and especially desorption of heavy liquid from the mineral surface, was improved by the use of surfactants as well as both thermal and mechanical energy. The examples given in the article indicate that heavy liquids which should find major commercial use in ore concentration are now available. The sink-float process using a heavy liquid medium has long been recognized as a nearly perfect method of mineral separation based on the differences in density of the minerals. The use of heavy liquid separation (HLS) offers the potential of an ore dressing method that provides the maximum possible natural concentrate grade and high recovery. In addi- tion the process requires no water, causes no waste disposal problems, and is extremely simple — the crushed ore is merely introduced into a liquid whose density is between that of the minerals to be separated. The minerals lighter than the liquid float while those heavier than the liquid sink. As such, HLS has become a standard laboratory tool but major commercial application has been prevented by the requirement for complete recovery of the heavy liquid, imposed by economics and materials handling hazards. Investigations, both in the United States and abroad, have shown that improved liquid stability as well as improved chemical engineering techniques brighten the future for HLS in ore dressing. The available commercial heavy liquids suitable for mineral separation cost from $2.00 to $20.00 per gallon1 and on this basis alone their complete recovery is economically desirable. Possibly of even greater importance is the fact that any residual heavy liquid remaining on the products could create a serious toxicity hazard. Consequently, to make the process commercially attractive requires (1) essentially complete removal of the liquid from the ore products and (2) optimum recovery of the liquid for reuse. Until recently these two stipulations have remained as obstacles but laboratory and pilot plant research now in progress suggests that these may soon be overcome. Various heavy liquid recovery systems have been proposed to meet these requirements and might be classified into one or more of the following three general classes: Solvent Wash (Fig. 1): The dissolution of the heavy liquid from the ore with a less expensive solvent,

Jan 1, 1969

By W. R. Boehme

The laboratory test work compares various 6 x 16 mesh activated coconut carbons using a gold standard solution of 4 ppm dissolved in sodium cyanide. The carbon adsorption rate and maximum loading are measured for different feed solution ionic contaminates: silver, copper, thiocyanate, and ferrocyanide. Conclusions are drawn that among the activated coconut carbons tested, maximum loading and the effects of specific ions varied significantly. The work presented can be applied to 6 x 16 mesh carbon selection tests and specific ion monitoring for industrial carbon adsorption systems.

Jan 1, 1984

By John B. Pearse

The works heretofore used in carrying out the Bessemer process have been constructed substantially as follows: The whole works or plant has been divided into three parts. 1. The division in which the metals for conversion are melted. 2. The division in which the metals are converted into steel and cast into ingots, or the converting department. 3. The engine and boiler department which supplies the hydraulic power required to move the machinery, and which furnishes the blast used for converting the crude metal into steel.

By John G. Hall

The pumping operations at the Chief mine have been unique in the respect that for many years the entire flow of water into the mine has been disposed of by pumping into natural underground "caverns" or "sinks." Volumes up to 4000 gpm have been pumped into these caverns with no apparent tendency for the caverns to fill with water, and it has never been proven that any of the water returns to the mine workings. Major pumping operations have been carried on during two periods, the first from 1918 to 1927, and the second from 1942 to the present time. Early Pumping Operations CHIEF MINE The Chief Consolidated Mining Co. was organized in 1909, and for the following nine years ore was developed and mined from the 600 level down to the 1800 level, all above the permanent water table, the 1820 level. The ore occurs as extremely-irregular pipes of replacement ore in limestone, and to date a total of approximately 2 1/2 million tons of ore has been produced. The ore was highly oxidized down to the permanent water level, and some highly-oxidized ore runs occur beneath the permanent water table. In 1916, with extensive mining being done on the 1800 level, 20 ft above the water table, there were indications that a zone of secondary enrichment might exist below the water, and in order to follow the ore runs to greater depth, it was realized that arrangements would have to be Made to pump water to the surface. The original pumping installation was designed to pump 350 gpm in five stages against a 1900-ft head to the surface. By January 1919, the water had been lowered to the 1900 level and an excellent grade of silver-lead ore was being produced from the zone of secondary enrichment. At this point. the flow of water equalled the capacity of the pumps, and pumps of an additional 1000 gpln were installed to supplement the original system. By January 1920, the water had been lowered approximately 100 ft, and the total flow of water was 1000 gpm. The water was being pumped from many different places in the mine where ore was being followed downward, with 27 different pumps operating below the 1800 level, mainly Cameron air-driven sinker pumps. These first sinking pumps were cumbersome and inefficient, and each of the larger pumps weighed approximately three tons and required tremeridous volumes of compressed air. During this period, an exploration drift being extended north and west from the Chief shaft on the 1600 level encountered a large natural cavern some 2000 ft from the shaft. This cavern in places measured 60 ft in diameter. Sometime later the disposing of water into this cavern was suggested, and as an experiment to determine the results of such an operation, an attempt was Made to fill the cavern by pumping in a small volume of water, but the water disappeared as fast as it was discharged into the cavern. It was decided then to try the cavern as a drainage outlet for the entire volume of water being pumped to the surface, and the water apparently did not return to the mine workings. Subsequently water up to 3000 gpm was pumped into this cavern. An attempt was made to determine whether any water was being returned from the cavern to the lower workings, and 100 lb of sodium fluorescein salt was added to the water discharged into the cavern, but no trace of it was found in the drainage water in the mine. Results of another such test will be described later.

Jan 1, 1950

By Klaus Detert, Joseph Ziebs

The progress of primary recrystallization has been studied in alloys of nickel with 0.01, 0.1, and 1.0 at. pct Ta. It can be shown that in one particular alloy containing 0.1 Ta and 0.02 02 a large grain structure (0.5 mm diameter) could he developed by primary recrystallization in thin sheetswhich were cold-worked to 97.5 pct reduction in thickness. Electron transmission microscopy demmstrates that this phenonmena was linked to the occurrence of fine precipitates at subboundaries. A discussion of how fine precipitates can exert under certain circumstances a rather distinct retarding effect on nucleation rate is included. IN a former investigation, the recrystallization and texture formation in a series of binary nickel alloys including pure nickel were studied.' Nickel alloys with some tantalum added displayed peculiarities during primary recrystallization which have been studied in more detail by the present investigation. The alloys were prepared by powder-metallurgy techniques similar to the previous investigations.' Pure metal powders were thoroughly mixed to obtain alloys with 0.01, 0.1, and 1.0 at. pct Ta, referred to as alloy A, B, or C, respectively. The metallic impurities were kept below 0.005 at. pct except iron and cobalt with slightly higher values, by using highly purified nickel and tantalum powders.' The mixtures were pressed into bars and sintered at a temperature of 1300°C. The sintering atmosphere was either high vacuum, dry hydrogen, or wet hydrogen. The sintered rectangular bars were rolled 60 pct to a thickness of 3 mm, then annealed under dry hydrogen at 1000°C and again cold-rolled to about 97.5 pct reduction of thickness. After this heavy deformation, the sheet thickness was 0.07 5 mm. In order to study primary recrystallization which occurred in measurable times between 400" and 600°C, samples of the rolled sheets were annealed in a salt bath at various temperatures for various times in order to study the progress of recrystallization by hardness measurements and microscopic examination. For optical observation, the samples were polished and etched in a solution of 65 cu cm HNO3, and 18 cu cm acetic acid in 20 to 100 cu cm H2O, depending of the particular sample and annealing treatment. For transmission electron microscopic observations, thin foils were prepared by the combined window and Bollmann techniques.273 The electrolyte consisted of 23 pct perchloric acid and 77 pct acetic acid held at 15°C. Polishing was performed at a current density of 0.4 amp per sq cm. The examination of the thin films was accomplished by a Siemens Elmiscope-type electron microscope operating with an accelerating voltage of 100 kv. RESULTS The progress of primary recrystallization was very accurately followed in these alloys in order to establish the time when recrystallization started and ended during an isothermal anneal. Table I gives some of the data obtained by these studies. The time for beginning is a mean value, where somewhat less than 5 pct of the matrix has recrys-tallized. The end of recrystallization is the time when at least 95 pct of the matrix has recrystal-lized. Listed are three different B alloys, with 0.1 Ta distinguished by different sintering atmospheres which influence the rate of primary recrystallization. By the above method, the state of the grain structure when primary recrystallization was completed, but grain coarsening had not started, could be clearly defined. The average grain diameter of the recrystallized matrix is also given in Table I. The grain size did not depend on the temperature of the isothermal anneal. With one exception the grain size of the primary recrystallized matrix was found to be rather fine

Jan 1, 1965

By A. Gittens

In two recent papers stevens13,14 has compared the methods used by various investigators to calculate the contribution of grain boundary sliding to the total creep strain. In previous work Gittins15 has shown that the strain due to grain boundary sliding Egb may be estimated from the lateral, d, and vertical, v, displacements of a longitudinal marker line as it crosses a grain boundary. He obtained the formula, where nL is the number of grains per unit length measured on the specimens before creep, and \p and ?2 are, respectively, the angle between the grain boundary and the stress axis in the plane of the surface and normal to the surface (see Ref. 13, Fig. 1). The subscript L indicates that observations are made in a longitudinal direction parallel to the stress axis. It was generally found more convenient to measure the grain size of a specimen before creep; this was necessary to characterize the specimen. Stevens13,14 however prefers to define Egb in terms of NL, the number of grains per unit length after creep. Since nL - NL(l + €1) where cr is the total strain, Eq. 1 11 becomes The quantities d and ?, can be determined directly from the marker lines, while v and ?2 are found by sectioning normal to the surface. In the interior when symmetry considerations suggest that v = d and ?1 - ?2 the strain due to grain boundary sliding may be estimated from a suitable internal longitudinal marker line by the equation To determine cgb at the surface using either Eqs. |1| or [2] requires a large number of observations, particularly if many estimates of cgb are required, and also it is impossible to determine the variation of cgh with creep strain on the same specimen because the specimen must be sectioned to determine ?2. Gittins15 found that for a number of materials the sliding vectors d and v were related to each other so that provided one was measured the other could be estimated. Also during creep, grain boundaries tend to migrate to meet the surface at angles close to 90 deg,16 so that the term nL(v/tan?2)L was comparatively small when 4,- 90 deg. By determining the frequency distributions of ?2 on control specimens this term can be estimated without sectioning the creep specimen. In this way an empirical equation based on Eqs. |1| and |2| enables cgb to be determined with reasonable accuracy with fewer observations of the angles and sliding components. For B brass, the relationship between the sliding vectors indicated that ceb = 0.85nLVL . In Table I of Stevens paper this equation is stated to be incorrectly derived. However it should be pointed out that this is an empirical equation derived from Eq. |1|, based on a knowledge of the relationship between d and v and on the frequency distributions of ?1 and ?2 for a particular material under given experimental conditions. It was used only when these relationships were known, to reduce the number of observations required to estimate cb. Therefore the comment "derivation incorrect does not apply. Gittins did in fact make use of Eqs. |1| and 131. Author's Reply I must apologize to Dr. Gittins for the inadvertent misrepresentation of his work. Other slight errors in Table I have been pointed out to me. Langdon (No. 12 Table I), Gifkins (No. 7, Table I), and possibly other workers did not measure the number of boundaries per unit length after creep (A') but the corresponding quantity bejove creep (n). As Dr. Gittins points out, a completely analogous set of equations to those given in the paper can be derived by the same method: using the number of boundaries per unit length before instead of after creep. The basic equations corresponding to Eqs. [2] and 171 of the paper, from which all the others can be derived, are

Jan 1, 1968

By L. Burris, C. H. Stockman, I. G. Dillion

Zone melting is a purification process in which separation of impurities is effected by slowly moving a narrow melted zone through a bar of solid material. Equations are presented which 1—predict the concentration profiles after successive passes of the zone and 2—give the limiting distribution of solute along the bar after an infinite number of passes of the molten zone. These equations take into account the effect of normal freezing in the last zone length. These equations were solved numerically using an electronic computer, and the concentration profiles are presented as a function of the solute distribution coefficient and the ratio of molten zone to total bar length. ZONE MELTING is a purification process in which separation of impurities is effected by slowly moving a narrow melted zone through a bar of solid material. Separation occurs if the solid freezing from a solution at any moment has a composition different from that of the liquid from which it is freezing. Solutes which lower the freezing point will tend to remain in solution and will be moved in the direction of zone travel. Solutes which raise the freezing point will preferentially freeze out, and they will be moved in a direction opposite to that of zone travel. After repeated unidirectional passage of a molten zone through a bar of material, impurities will be concentrated in the ends of the bar. A limiting concentration gradient will be reached in the bar after a large number of passes of the molten zone. Both the magnitude of this limiting concentration gradient and the rate of movement of solute in any pass depend on the zone length and on the equilibrium concentrations of solute in the liquid and solid phases. The basic theory of zone melting was thoroughly covered by Pfann.1 The purpose of this paper is to extend the theory of zone melting by presenting 1—equations which relate the concentration profiles in successive passes of the zone and 2—equations which give the limiting distribution of solute along the bar after an infinite number of passes of the molten zone. These equations take into account the effect of normal freezing* in the last zone length concentration of a given solute in the solid to the concentration of the same solute in the liquid. Analytical methods for determining the concentration profiles after any pass have been developed by Lord2 and Reiss. The method developed by Lord does not take into account normal freezing in the last zone and also becomes very complicated when the number of passes becomes large. The method developed by Reiss enables the direct calculation of the complete concentration profiles (including that in the last zone) after any pass for values of k close to unity (0.9 < k < 1.1). Both methods are useful and provide valuable insight into the zone melting process. The equations given in this article are based on the same assumptions used by Pfann. These are: 1—uniform composition in the liquid, 2—negligible diffusion in the solid, 3—constant distribution coefficient, 4—constant length of the liquid zone (except at the ends), and 5—solubility of the solute in the melt is not exceeded at any point. Equations for Concentrations in Successive Passes For a melted zone of length 1 at the left-hand end of a bar of uniform unit cross-section, the amount of solute in the liquid is J0Cn-1(x) dx [1] where the subscript is the number of the pass. Cn-1(x) represents the concentration (expressed as weight per unit volume) obtained in the previous

Jan 1, 1956

By Rodney P. Smith

The solubility of niobium carbide in y iron has been determined for the temperature range 1000" to 1300°C from the carbon content of a series of Fe-Nb alloys equilibrated with hydrogen-methane mixtures. The method is similar to that used by Darken, Smith, and Filer 1 to determine the solubility of aluminum nitride in y iron and by smith2 to determine the solubility of niobium nitride in y iron. The samples* were strips 1.2 by 3 by 0.035 cm. The niobium oxide.* For samples of composition slightly

Jan 1, 1967

By T. K. Matthews

Minerals are mined as a one-time crop, and the processing of one ton of ore results in there being one ton less to be recovered from the deposit. Although some minerals occur in greater abundance than others, they are to a greater or lesser degree exhaustible. Depletion Allowances One of the distinctive features of the mineral industries is the allowance for depletion. The depletion allowance is a much maligned and misunderstood principal of our tax structure. Although imperfect in design, it has nevertheless served as a constructive incentive for the development and expansion of the minerals industry.

Jan 1, 1969

By J. H. Johns, G. R. Stephan

The Climax mine has a history of dealing with ground control problems caused by major stress concentrations in the orebody. These stress concentrations, called weight, can arise in a variety of ways. The general effects of depth, rock type, and mining method give rise to weight problems, but local geologic features and daily cave line management can have even greater effects. Over the years a number of techniques have been developed for identifying weight problems in advance and for cow with them. These techniques all help us to amid interrupting the pace of cave advance and minimize the loss of production.

Jan 1, 1984

By O. B. Harden Harden

IN working out the geological structure of a complicated district, where the problems are difficult to solve by the ordinary methods in use, a model, upon which all the geological and topographical data may be put, will be found of great value, bringing, as it does, all the facts in their proper relationship at once under the eye. In the teaching of structural geology, also, a model is of great value, since the student can thus have before him, in miniature, a representation of the rocks, as he would see them in place. Topographical models-may often be useful in suits at law for damages in rail-road accidents, or for trespass against mining companies, etc., in order

Jan 1, 1882

By Booker Morey, Ashok Gupta

Amid controversy, the resource recovery industry is rapidly approaching its first commercial operations. Some of the problems are reviewed along with some of the important systems that are being developed for resource recovery. Market specifications for the inorganic fraction recovered from municipal solid waste are reviewed along with the details of the unit operations that have been tested.

Jan 1, 1976

By Walter Crafts

STRUCTURAL steels containing chromium have become widely used in the last 20 years. In the earlier part of this period the major applications were in chromium-molybdenum aircraft tubing and similar special high-strength steels. During the past 10 years there has been a notable expansion in the use of chromium-bearing steel for structural purposes when welding was utilized for fabrication. The chromium-bearing steels introduced by Saklatwalla1 (chromium-copper), Kinzel2 (chromium-manganese-silicon), and Schultz3 (chromium-copper) are being extensively used at the present time.

Jan 1, 1939

By Ying-Chu Hoh, Renato G. Bautista

The chemically based models previously developed to describe and predict the distribution coefficients and separation factors for several different single and multicomponent metal extraction systems are presented in a more general form in this paper. The degree of formation of the aqueous phase inorganic metal complexes, the extraction mechanism of the complexes, and the equilibrium constant for the extraction reaction are used in the development of the models. The concept has been successfully applied to extraction systems where the organic solvent extracts an ionic species and where the organic solvent extracts only the neutral species. In addition, a model to describe a system where the organic solvent is extracting several extractable species has been derived on the same basis

Jan 1, 1981

By Kenneth L. Ancell

This paper traces the history and performance of a coal degasification project in the Black Warrior Basin of Central Alabama. The primary coal beds being degasified are the Mary Lee and Blue Creek. The area being developed for gas production is also scheduled for underground mining in two to seven years. The paper discusses the early data collection and evaluation. The coals are extremely gassy with some measurements in excess of 500 SCF/ton. Five experimental wells were drilled and produced. Geologic heterogeneities were encountered which affect the performance of the project. Drilling and completion practices together with well design considerations compatible with under- ground mining are discussed. Finally, the gas gathering and sales system together with the early production history is illustrated.

Jan 1, 1982

By B. S. Taylor, J. D. Fraser, W. L. Chen

Steam coal quality is less stringent than metallurgical coal quality. Steam coal is prepared either by cleaning a full-size range of raw coal or cleaning coarse coal only, leaving the fine coal raw. The latter scheme is more cost competitive. It is particularly applicable to US eastern soft coal where low ash coal is concentrated in fine sizes. For raw coal size separation the Rotating Probability Screen, developed by the British Coal Board, may be used for easy change of raw coal separating size, and to avoid screen deck plugging when handling wet and clayey raw coals.

Jan 1, 1985

By Clark A. Sumner

An apparatus for agglomerating fine particles was developed in 1969 as a joint effort of Stauffer Chemical and O’Brien Industrial Equipment. Commercial sized units were built for agglomerating home dishwashing detergents. Recently, as a result of further pilot development, variation of bed depth and cage configuration has resulted in a means of controlling the amount of internal classification from a minimum, producing small agglomerates, to a maximum, producing compact rounded agglomerates. Little or no external classification of discharged agglomerates is needed. At maximum internal classification the agglomerates roll on a bed of fine particles rather than the shell resulting in increased agglomerate size.

Jan 1, 1977

By F. N. Kissell, W. E. Schroeder, N. I. Jayaraman

This article discusses methods of improving airborne capture by conventional dust suppression systems. The work was performed in a full-scale plywood model of a mine entry with a wooden mockup of a Joy 12 CM continuous miner. Dust was introduced at the front of the entry, and water sprays knocked down part of the dust cloud. Sampling in the return air determined the percentage dust reduction. Based on the results, guidelines are offered for continuous miner spray systems to control dust. Empirical relations have been developed to determine the effect of changes in water spray pressure on air movement, dust knockdown, and pollutant concentrations in the face return. Nozzle selection and arrangement are discussed for high dust knockdown efficiencies.

Jan 1, 1986