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By S. A. Braley

DRAINAGE water from coal mines is probably the most serious water pollution problem today, varying in importance according to location of the mines and geological structure. Drainage may be either acid or alkaline in character. Acid discharge, the most severely detrimental to a stream, is caused by natural oxidation of the sulfuritic material (FeS2) in the strata associated with the coal seam. Since the acid is the result of a natural reaction the acid water differs because it does not cease with abandonment of the mining operations. There is no known economical method of neutralizing acid mine water or any practical method to prevent oxidation of exposed pyrite. Since production of acid from a mine does not stop when mining stops, the total quantity produced depends entirely upon the excavated areas. The increasing volume of acid water in manv mines has greatly increased operating costs. Pumping is expensive and acid mine waters are destructive of all equipment, especially metals in pumps and piping, and necessitate the use of corrosion-resistant materials. Discharge of the acid mine drainage into streams neutralizes their normal alkalinity, causes them to become acid, and produces an environment unfavorable for aquatic life and unsuited for industrial or domestic use without costly treatment. Mine drainages vary in percentage composition over wide limits, although the usual dissolved substances are ferrous and ferric iron, aluminum, calcium and magnesium sulfates, and lesser amounts of sodium, potassium and manganese sulfates and chlorides. Some alkaline discharges may contain heavy concentrations of iron as iron bicarbonate. These waters may produce iron hydroxide deposits in the receiving stream but do not cause it to become acid. In the extensive literature on acid mine water there appears to be a great deal of confusion about the importance of various components and the methods for their determination. In many instances faulty conclusions have been drawn from use of unsuitable methods of analysis. It is desirable that the factors and terms used in evaluation of analyses of mine waters should be so clearly defined that any interested person could properly appraise any analytical report. Some analysts report complete chemical analyses of mine waters and neglect to record drainage volumes. Others report only a minimum of analytical data after taking no precautions to preserve the original composition of the water during the time elapsing between collection and analysis. Some use methods of analysis intended for so-called pure waters of the potable and boiler water classes. These methods are not applicable to highly buffered waters such as mine water. Probably the most common criteria for evaluation are pH, free acidity, or acidity or alkalinity to methyl orange or methyl red, total acidity or acidity to hot phenol-phthalein, and the sulfate content. If these determinations are made on carelessly collected samples after a few days to weeks standing in warm rooms, they do not in any way represent the character of the water flowing from the mines. It is hoped that a brief discussion of the fundamental value of some of these factors may lead to a bettqr understanding of the need for more careful evaluation of mine water discharges. The term pH is one used by chemists to express relative acidity or alkalinity in terms of concentration of effective hydrogen ion in a solution. It is defined as the negative logarithm of the hydrogen ion concentration or activity in equivalents per liter. pH = logarithm A neutral solution, which is one containing the same number of hydrogen and hydroxyl ions, has a pH of 7. As the hydrogen ions increase and the solution becomes more acid, the pH decreases toward zero; and as the hydroxyl ions increase and the solution becomes more alkaline, the pH approaches 14. When dissolved in water to a dilute solution acids like sulfuric and hydrochloric, commonly known as strong acids, ionize completely, and the pH or hydrogen ion concentration varies with molar concentration of the dissolved acid. However, in high concentrations of such acids, the pH or hydrogen ion concentration is less than the acid concentration because the acid does not completely ionize. In only very dilute solutions does the pH represent the total amount of acid that can be neutralized by an alkali. All ionization reactions are equilibrium reactions. If other chemicals are added to the solution of an acid and the added chemical produces an ion that is the same as one of the ions of the acid, the degree of ionization of the acid is altered and the pH changes to some value that represents the active hydrogen ion of the new solution. Thus if iron sulfate is added to a solution of sulfuric acid the pH increases, since the common sulfate ion suppresses the degree of ionization of the sulfuric acid and decreases the effective hydrogen ion. However, the total acidity of the solution is increased. There are two salts composed of iron and sulfate— ferrous and ferric sulfate. In these salts there is no hydrogen ion that can ionize to give an acid solution, but when they are dissolved in water, the pH is less than 7 and the solution becomes acid. This is caused by a reaction known as hydrolysis and is represented by the equations FeSO4 + 2H2O ? Fe(OH)2 + 2H + SO,: or Fe,(SO,)3 + 6H2O ? 2Fe(OH)3 + 6H+ + 3504 A solution with a total acidity of 5000 ppm according to the first equation will have a pH of 4.40 but one with an acidity of 75 ppm, according to the sec-

Jan 1, 1958

By Jay A. Carpenter

IN six years the construction of Hoover Dam and the power plants probably will have reached the operating stage and this vast new source of power will then be continuously available for industry. The height of the darn will be 580 ft. and the average annual flow of the Colorado River is calculated at approximately 24,000 cu.ft. per sec. If this full head and average flow can be maintained the year round, about one million horsepower will be generated. The U. S. Reclamation engineers, taking into consideration the probable range in flow and head, have set the figure for power contracts at 660,000 firm horsepower, which can be assured at all times of the year. During the months of heavy run-off of water from the Colorado River Basin the power plants should be able to generate the full 1,000,000 hp. capacity of the power plants, so there will be 340,000 excess or secondary horsepower. The cost of firm power at the switchboard will be close to 2.15 mills per kw.-hr., equal to $14.05 per hp.-yr., and the cost of the secondary or off-peak power will be close to 0.8 mills per kw.-hr., or $5.22 per hp.-yr. In no place in the country can such cheap power be obtained today. The electrochemical industry is in a state of flux because the power costs from its old sources have increased so much (to $30 per hp.-yr. at Niagara) that the industries have had to seek new sources, so they have gone to Canada and Norway and built new plants.

Jan 1, 1932

By Fred C. Bond

MOST investigators are aware of the present unsatisfactory state of information concerning the fundamentals of crushing and grinding. Considerable scattered empirical data exist, which are useful for predicting machine performance and give, acceptable accuracy when the installations and materials compared are quite similar. However, there is no widely accepted unifying principle or theory that can explain satisfactorily the actual energy input necessary in commercial installations, or can greatly extend the range of empirical comparisons. Two mutually contradictory theories have long existed' in the literature, the Rittinger and Kick. They were derived from different viewpoints and logically lead to different results. The Rittinger theory is the older and more widely accepted. In its first form, as stated by P. R. Rittinger, it postulates that the useful work done in crushing and grinding is directly proportional to the new surface area produced and hence inversely proportional to the product diameter. In its second form it has been amplified and enlarged to include .the concept of surface energy; in this form it was precisely stated by A. M. Gaudin2 as follows: "The efficiency of a comminution operation is the ratio of the surface energy produced to the kinetic energy expended. According to the theory in its second form, measurements of the surface areas of the feed and product and determinations of the surface energy per unit of new surface area produced give the useful work accomplished. Computations using the best values of surface energy obtainable indicate that perhaps, 99 pct of the work input in crushing and grinding is wasted. However, no method of comminution has yet been devised which results in a reasonably high mechanical efficiency under this definition. Laboratory tests have been reported' that support the theory in its first form by indicating that the new surface produced in. different grinds is proportional to the work input. However, most of these tests employ an unnatural feed consisting either of screened particles of one sieve size or a scalped feed which has had the fines removed. In these cases the proportion of work" done on. the finer product particles is greatly increased and distorted beyond that to be expected with a normal feed containing the natural fines. Tests on pure crystallized quartz are likely to be misleading since it does not follow the regular breakage pattern of most materials but is relatively harder to grind at the finer sizes, as will be shown later. This theory appears to be indefensible mathematically, since work is the product of force multiplied by distance, and the distance factor (particle deformation before breakage) is ignored. The Kick theory' is based primarily upon the stress-strain diagram of cubes under compression, or the deformation factor. It states that the work required is proportional to the reduction in volume of the particles concerned. Where F represents the diameter of the feed particles and P is the diameter of the product particles, the reduction ratio Rr is F/P, and according to Kick the work input required for reduction to different sizes is proportional to log Rr/log 2.5 The Kick theory is mathematically more tenable than the Rittinger when cubes under compression are considered, but it obviously fails to assign a sufficient proportion of the total work in. reduction to the production of fine particles. According to the Rittinger theory as demonstrated by the theoretical breakage of cubes the new surface produced, and consequently the useful work input, is proportional to Rr-1.5 If a given reduction takes place in two or more stages, the overall reduction ratio is the product of the Rr values for each stage, and the sum of the work accomplished in all stages is proportional to the sum of each Rr-1 value multiplied by the relative surface area before each reduction stage. It appears that neither the Rittinger theory, which is concerned only with surface, nor the Kick theory, which is concerned only with volume, can be completely correct. Crushing and grinding are concerned both with surface and volume; the absorption of evenly applied stresses is proportional to the volume concerned, but breakage starts with a crack tip, usually on the surface, and the concentration of stresses on the surface motivates the formation of the crack tips. The evaluation of grinding results in terms of surface tons per kw-hr, based upon screen analysis, involves an assumption of the surface area of the subsieve product, which may cause important errors. The'evaluation in terms of kw-hr per net ton of 200 mesh produced often leads to erroneous results when grinds of appreciably different fineness are compared, since the amount of -200 mesh material produced varies with the size distribution characteristics of the feed. This paper is concerned primarily with the development, proof, and application of a new Third Theory, which should eliminate the objections to the two old theories and serve as a practical unifying principle for comminution in all size ranges. Both of the old theories have been remarkably barren of practical results when applied to actual crushing and grinding installations. The need for a new satisfactory theory is more acute than those not directly concerned, with crushing and grinding calculations can realize. In developing a new theory it is first necessary to re-examine critically the assumptions underlying

Jan 1, 1952

By Robert B. Fulton

It is a pleasure to talk to you about the world economics of industrial minerals on this occasion of AIME's 100th anniversary. In order to fit this topic into the time and tenor of such a session as this where you are hearing subjects equally broad in scope one after the other, I will deal with the subject through the medium of eight minerals selected to illustrate import- ant economic factors. I must admit to some bias in the selection, since I picked personally familiar ones to elucidate facts of life in the industrial minerals business. It is a very interesting field, different in many respects from metals and certainly far different from petroleum. The minerals chosen are celestite, diamond, fluorspar, salt, sulfur, ilmenite, rutile and zircon. Celestite The market for celestite is mainly in the U. S. where it is consumed in the manufacture of strontium compounds for red-colored flares and other pyrotechnics, for the production of electrolytic zinc, for glass manufacture, and for the manufacture of ceramic ferrites. Smaller markets exist in Europe and Japan. Production comes from England, Mexico and Spain, and a new mining operation is being started in Nova Scotia. There is no significant production in the U. S. nor in the other principal consuming countries, except England; so it is mainly an international commodity. Celestite illustrates many of the factors important in the economics of industrial minerals. To have a marketable product, grade and freedom from deleterious elements are important for materials used by the chemical industry. In the celestite used to make strontium nitrate, calcium is the most objectionable of the commonly associated cations. Selling price is adversely affected because nitrate manufacturing costs are greatly increased if calcium is pre sent, hence less can be paid for the raw material. Meeting specifications can add significantly to the vendor's cost; hence a balance must be struck between the producer's capability and the consumer's requirements, permitting a reasonable profit for both. Sizing of industrial mineral products is often important to the consum-

Jan 1, 1971

By J. A. Rowland, C. W. Funk

THIS investigation is a part of the United States Bureau of Mines work in conserving the Nation's resources. The isothermal sections presented were developed as a guide to a comprehensive investigation of the properties and fabricating characteristics of copper-base alloys containing manganese and tin. These alloys are being investigated as possible replacements for the commercial bronzes containing substantial quantities of tin. Development of the isothermal sections has, therefore, been limited to the area between 0 to 20 pct Mn and 0 to 25 pct Sn. Since Heusler1 reported the ferromagnetic properties of the Cu-Mn-Sn alloys in 1903, the system has been the subject of several investigations. These studies were largely concerned with correlation of magnetic properties and crystal structure.2-6 Vero,' however, established the solidus surface of the system, between 20 pct Mn and 40 pct Sn, by thermal and microscopic methods. The a, ß, and y solid solutions were also observed by Vero and reported to extend into the interior of the ternary system. For the present study, the annotated diagram of the Cu-Sn system prepared by Raynor' and the Cu-Mn diagram by Dean and coworkers9 were used as the binary borders of the ternary system. Procedure The metals used in this investigation were electrolytic manganese, washed cathode copper, and three-star tin. The manganese, produced at the Boulder City pilot plant of the Bureau of Mines, had a purity of 99.9 pct; the copper contained less than 0.004 pct Bi and 0.001 pct S; the tin contained 0.070 pct Pb, 0.070 pct Sb, 0.050 pct Fe, and 0.065 pct Si. Heats of 1 lb were melted in alundum crucibles with a 3-kw induction furnace. The heats were chill-cast in iron or copper molds to form 6-in. ingots 3/4 in. in diameter. Molds were washed with graphite or zirconia, depending on the manganese content, and preheated to 150°C. The maximum impurity content of any alloy was 0.03 pct Fe, 0.02 pct Si, and 0.02 pct Al; in most cases, they contained less than half this quantity of any of these impurities. Composition of the alloys and their treatment prior to homogenizing are shown in Fig. 1. Designations adjacent to each composition refer to the heat numbers. With a few exceptions, ingots containing 20 pct Sn or less were hot-swaged. These ingots were swaged, after a 24-hr preheat at 650°C, to obtain a total reduction of 70 pct in cross section. Intermittent reheating was necessary, and reductions were limited to either 0.025 or 0.050 in. in diameter per pass. Selection of a 200-hr homogenizing treatment was based on extensive experiments which indicated that virtual equilibrium was reached after 150 hr. Structures were retained by quenching in water. Grain sizes in the specimens homogenized at 350" and 450°C were generally small. These homogenizing treatments were preceded by a 50-hr anneal at 650°C to facilitate phase identification by increasing the grain size in these specimens. The specimens were furnace-cooled from this temperature to the homogenizing temperature, and the heat treatment was continued for the standard 200-hr period. The homogenized samples were bisected to provide an internal section for metallographic examination. Optimum definition of the microstructure was obtained by successive polishing and etching. Quarter-strength ASTM copper reagent No. 13 gave the most satisfactory results.10 Filings from interior sections of the samples were used for diffraction studies. These filings were annealed in evacuated glass tubes at homogenizing temperatures for periods exceeding 50 hr and quenched in water. During this treatment the manganese content was reduced by significant amounts, probably by vaporization. This change in composition was considered in applying the diffraction data, which were used only as a means of identifying phases where the data were consistent with the metallographic evidence. Phase Identification The a solid solution is readily distinguished by the copper-colored, twinned, polyhedral grains, as shown in Fig. 2, and a face-centered cubic X-ray pattern. Tukon indentations, using a 25-g load, indicate a Knoop hardness of 155 for a quenched from 650°C. Using white light developed by a Wratten 78 A filter, the ß grains in etched specimens have a dark brown tint in contrast to the lighter copper-colored grains of the a phase. Knoop hardness values of 284 were obtained for the ß structure in specimens quenched from 650°C. The ß structure was also distinguished from the a phase by its body-centered cubic lattice. The acicular structure appearing in Fig. 3 was evident in several alloys of higher tin content after quenching from either 750' or 650°C. Since a trans-

Jan 1, 1954

By K. C. Ten Brink, Allen D. Garrison

A rrevious paper1 reviewed some of the properties of clays and shales and presented some data on the nature of the gelling phenomenon. It included a brief discussion of origin of clays and shales, their composition, structure, retention of water, ionization, base exchange, and the nature of the gelling of water suspensions. Other publications of the Institute2-8 have discussed the properties of clays and their relations to drilling-fluid behavior and control. It is the purpose of this article to present some new data obtained and arranged in a different manner and making possible a better understanding of the chemical reactions involved in the control of drilling-fluid characteristics during the normal drilling operations. Experimental Procedure The difficulties involved in the measurements of the flow characteristics of plastic fluids are too familiar to recite. The possible applications of some of the well-known instruments to this problem have been adequately discussed by Hobson9 and he has given an excellent bibliography. Although there are many viscosimeters of varied design, none of them appear to fulfill the most obvious demands of accuracy in the measurement of clay or shale suspensions. Since the fluid characteristics are dependent on the rate of flow of the fluid, and the viscosity usually falls off very rapidly as the rate is increased, it is clear that the first requirement is that the instrument shall be capable of a wide variety of speeds. The viscosity of the fluid at any fixed and constant rate of flow is also a function of time, so that the second demand is that the instrument shall be capable of a constant and continuous motion for a prolonged period of time.

Jan 1, 1940

By K. C. Ten Brink, Allen D. Garrison

A rrevious paper1 reviewed some of the properties of clays and shales and presented some data on the nature of the gelling phenomenon. It included a brief discussion of origin of clays and shales, their composition, structure, retention of water, ionization, base exchange, and the nature of the gelling of water suspensions. Other publications of the Institute2-8 have discussed the properties of clays and their relations to drilling-fluid behavior and control. It is the purpose of this article to present some new data obtained and arranged in a different manner and making possible a better understanding of the chemical reactions involved in the control of drilling-fluid characteristics during the normal drilling operations. Experimental Procedure The difficulties involved in the measurements of the flow characteristics of plastic fluids are too familiar to recite. The possible applications of some of the well-known instruments to this problem have been adequately discussed by Hobson9 and he has given an excellent bibliography. Although there are many viscosimeters of varied design, none of them appear to fulfill the most obvious demands of accuracy in the measurement of clay or shale suspensions. Since the fluid characteristics are dependent on the rate of flow of the fluid, and the viscosity usually falls off very rapidly as the rate is increased, it is clear that the first requirement is that the instrument shall be capable of a wide variety of speeds. The viscosity of the fluid at any fixed and constant rate of flow is also a function of time, so that the second demand is that the instrument shall be capable of a constant and continuous motion for a prolonged period of time.

Jan 1, 1940

By J. H. Pratt

In a recent paper* on the origin of corundum associated with the peridotites of North Carolina, attention was called to the constant occurrence of the mineral, chromite, in these rocks. The field-data obtained during the preparation of that paper have been used largely in the present paper. During the past summer the general occurrences of chromite have been carefully studied, and those in North Carolina have been especially examined in the field. It is purposed to give, in this paper, the reasons why the chromite should be regarded as having been formed at the same time with the peridotites, having been held in solution by the molten mass of the peridotite, and crystallizing out among the first minerals as the mass began to cool. This theory is essentially the same as that advanced by me for the origin of the corundum associated with the peridotite rocks, and a similar line of reasoning has been used to substantiate the theory proposed. Investigations concerning the igneous origin of some of the ores have been materially aided, during recent years, by the skillful experiments of Morozewicz and Lagorio,§ and by researches that show us, more clearly, why we should regard a fused mass of rock as a liquid, having similar properties to an ordinary solution. These experiments have shown that a molten basic glass, similar in composition to the basic magnesian rocks, dissolves alumina readily, and that upon cooling, the first minerals to separate out are corundum and spinel.

Jan 1, 1900

By H. H. Hausner, H. S. Kalish

IN recent years zirconium and beryllium have become of great interest because of their special properties. Zirconium is known for its remarkable ability to absorb the gases oxygen, nitrogen and hydrogen, has extensive use as a getter in electron tubes and has indeed become the primary metal in the manufacture of photoflash bulbs. Beryllium is of interest because it is light metal and as such could supplement aluminum and magnesium in light metal applications. The late development of zirconium ductile at room temperature produced by both the iodide and calcium reduction methods has stimulated frequent articles on this metal in the recent literature. Beryllium is so closely akin to zirconium crystall-ographically that some investigators are of the opinion that beryllium should be ductile at room temperature if it were pure enough. The crystallographic relationship of beryllium and zirconium, the interest created in Be-Zr alloys where the zirconium is believed to function as a remover of solid solution gas in beryllium and thus increase the ductility of the beryllium rich alloy, and the amenability of these metals to powder metallurgy methods stimulated this approach to an investigation of the system Be-Zr. Preliminary work showed that a zirconium compact melted when pressed in contact with a beryllium compact and sintered at 1250°C, without any apparent final bond between the two. A search of the literature revealed nothing pertinent about the Zr-Be system or about any zirconium - beryllium alloys. In 1936 Misch³ vaguely identified the compound ZrBe2 and reported that it has a deformed cubic structure. Two patents," , were issued to Mach-lett Laboratories in 1943 and in 1946 regarding the addition of small amounts of zirconium to beryllium in order to improve the ductility of beryllium so that it may better be used for X ray windows. In considering the alloy formation of beryllium with other metals Raynor thought it unlikely that zirconium would have any solid solubility in beryllium. From this indefinite information nothing could be ascertained about the possible use of Zr-Be alloys or the use of these metals intimately. A brief preliminary study of the system Zr-Be by powder metallurgy methods was undertaken to provide regions for more intense studies of Zr-Be alloys and as a guide to a formal phase diagram determination. The investigation described in the following pages was completed in less than three weeks. The usefulness of this rapid method to facilitate phase diagram studies and new alloy investigations will be shown. Materials and Procedure Zirconium powder was produced from crystal bar stock of high purity supplied by Foote Mineral Co. The crystal bar was converted to hydride at 730°C, crushed, and the hydride decomposed in purified argon at 850°C. 60 pct of the powder passed through a 325 mesh screen, 40 pct of the powder was —140 mesh, +325 mesh and the average particle size of the total was 8 to 12 microns. The beryllium powder was used as re-

Jan 1, 1951

By M. A. Childers

Recent work with controlled laboratory tests.' field experience and a new analytical approach indicate that casing centralization, pipe movement and relative rheological properties between the mud and cement (yield point and density) are the three controlling factors in successfu1 primary cementing of dual casing. Turbulent flow has been demonstrated to be unnecessary under the severe displace ment conditions of parallel dual strings of casing. Simultaneous reciprocation of dual casing in directionally drilled boreholes is highly dependent on centralization and on the rapid initiation of pipe movement once the casing has been run to total depth. Successful cementing of 30 out of 32 dual, 3%-in casing wells substantiated these conclusions. A review of 15 dual, 3 1/2-in. casing wells completed prior to the adoption of these principles disclosed only one success. Mathematical and graphical techniques permit the engineer to analyze and anticipate the success or failure of a cementing operation. Standoff, mud displacement and reciprocation can be analyzed mathematically and planned in deep, directional or straight boreholes by review of annular geometry, relative rheological properties between the mud and cement, and forces caused by caring weight, functional drag and differential sticking. Introduction The objective in primary cementing is to displace the drilling fluid with cement that will bond to the formation and casing to prevent communication through the annulus. The predominant cause of failure appears to be channels of gelled mud that remain in the annulus after cementing.1 '* Past investigations have noted many procedures that could improve the displacement process.2-' These procedures include centralization, chemical preflushes, mechanical mud cake removers, controlled flow rates in certain regimes, pipe movement an1 control of mud and cement properties. Not until recently have the effects of many of these procedures been isolated so that their individual importance could be determined. In a recent paper,' a comprehensive and analytical laboratory study of various factors controlling the displacement process concluded that there are three major factors in obtaining complete mud displacement by the cement slurry. 1. Good relative rheological properties between the mud and cement. The cement should be heavier and have a yield point and plastic viscosity higher than the mud being displaced. Buoyancy forces often can dominate the displacement process. 2. High degree of standoff or centralization. The more centered the casing, the greater the likelihood of complete mud displacement and removal. 3. Pipe movement obtained by rotation or reciprocation. Often unfavorable rheological properties between the mud and cement, and poor standoff can be compensated by pipe movement. This work demonstrates the compatibility of the latter concepts with field results extended to the more severe dual casing displacement process. Early Experience With Dual 3'/t-In. Primary Cementing Early slim casing completions in the West Delta Block 73 field, offshore Louisiana, included 17 dual 3%-in., 3 single 3% -in. and 1 dual 2% -in. casing installations. Table I provides data on these earlier wells. Usually, common Class A cement was placed across the pay zone and 12 percent gel cement was placed to a height of 500 ft above the highest productive sand. Centralizers (6%-in. OD maximum expansion) were placed two or three per joint on both casing strings from total depth through the completion intervals and at least one per joint for the remainder of the cement sheath. A number of techniques and procedures were attempted to improve cementing success. Casing was run simultaneously or separately with a number of mud conditioning and circulating techniques. Mechanical aids such as cement bonders, slmiding sleeves, turbulizers and various sizes and strengths of centralizers were used. Common and gel cements were mixed with fresh or sea water with a number of different additives to control retardation and dispersion. Cements were thinned and mixed at high rates to try to induce turbulent flow in the annulus. On one well the "slow flow" method was used." Perforating was accomplished with a jet orienting gun. Unsuccessful cement jobs were indicated by production of extraneous fluids or by pressure communication between zones. On test, five wells flowed water, four flowed sand and six had pressure communication with the other string or the 10%-in. casing. A total of 25 remedial squeeze ce-

Jan 1, 1969

B. S. and E. M. from the School of Mines at Socorro, New Mexico, in May, 1912. Immediately after graduating, he began as a practical miner in the U. S. mines at Bingham, Utah, but in November, 1912, he entered the operating department of the Chino Copper Co.'s concentrator at Hurley, New Mexico, where he stayed until July, 1913. During the latter half of 1913, he was employed as chemist in the laboratory of the El Paso smelter, at El Paso, Texas. In December, 1913, he became engineer and geologist with the Empire Zinc Co., at Socorro, New Mexico, which position he held at the time of his admission to the Institute in 1914, and until he entered the army in June, 1918. Before beginning the study of mining, Mr. Ringlund had completed a course in pharmacy, which accounts for his enlisting in the Medical Corps. Lieutenant George Roper, Jr. George Roper, Jr., a Junior Member of the Institute, was killed in an aeroplane accident near Shotwich, England, on Play 25, 1918. He was born at Steubenville, Ohio, in 1893, and at the time of his admi~sion to the Institute in 1916 was a student in mining and metallurgy at the Massachusetts Institute of Technology, Boston, Mass., from which he received the degree of S. B. in June, 1917. In August, 1917, he enlisted in the Royal Flying Corps, and after preliminary training on this side of the water, was sent to England. He was making the final cross-country flight of his course of training when the accident occurred.

Jan 1, 1920

By H. M. Howe

(New York: Meeting, October, 1913.) 1. INTRODUCTION.-This paper gives the results of our micro-graphic determinations of the position of Ae3 in a series of 14 hypo-eutectoid steels of varying carbon content, one of them very rich in phosphorus, and of SEe for one hyper-eutectoid steel. We have pro-ceeded chiefly by noting the temperature -at which the reabsorption of the pro-eutectoid element is found to be complete after a 30-min. holding, but in part also by noting the temperature at which its precipitation begins. Observations on these same steels by Messrs. Burgess, Crowe, and Rawdon, of the United States Bureau of Standards, are given in a paper' issued simultaneously with this, while in a third paper s one of us compares these data with those of earlier observers. These three papers are treated as one whole by numbering their sections, tables, and illustrations continuously, and by setting the folding plates of micrographs at the end of the last paper. THE COMPOSITION OF THE STEELS EXPERIMENTED ON is given in Table I. Notation. and Definitions.-Following Osmond's development of Tschernoff's notation; in which " Ac3 " is the upper limit of the hypo-eutectoid part of the transformation range as observed in heating up, and " Ar3 " is that same limit as observed in cooling down, and following the usual custom of calling the upper limit of the hyper-eutectoid part of that range "SE," we may go farther and call this latter limit " SEc " when observed in heating up .and " SEr " when observed in cooling down.

Jan 6, 1913

By D. C. Shelton

This paper discusses the metallurgical testing procedures and their applications at the laboratories and mills of Pima Mining Co. The author shows, using discussions and case histories, that routine sampling and evaluation of daily and monthly data can be as valuable as thorough laboratory investigations and mill test work. A great deal of valuable information can be made available this way. This paper discusses laboratory and mill metal-lurgical testing procedures and their applications at Pima Mining Co., located about 25 miles southwest of Tuscon, Ariz. The Company, incorporated in 1951 (the year the orebody was found) in California to operate the Pima mine, is jointly owned by Cyprus Mines Corp., Utah Contruction and Mining Co., and Union Oil Co. Milling started in December 1956. The main Pima orebody is a contact metamorphic vein of copper-bearing hornfels with an average dip of 45° to the south with an east-west strike. The hanging and footwalls are quartzitic pyroclastics (arkosite) containing some sulfides. Barren igneous intrusions are also present. Conglomerates and alluvial wash cover the deposit to a depth averaging about 200 ft. The principal copper mineral is chal-copyrite. Native copper, chalcocite, and minor amounts of chrysocolla and other copper minerals are also present, together with sphalerite and some molybdenite. Generally, it has been noted that the ratio of chalcocite to chalcopyrite is higher in the lower grade ores, while as the grade increases, there are only minor amounts of chalcocite. Open pit mining is employed at Pima, using both trucks and an inclined skiphoist system. A detailed description of Pima Mining Co. operations was presented by Spaulding et all. More recently another orebody to the north and east of the present pit has been discovered and will be placed in operation within a few years. This ore is of a porphyritic rather than a lode nature since values are disseminated in both igneous and metamor- phosed sedimentary rocks. The new orebody is expected to be harder and cleaner than the present deposit. The views expressed in this article are those of the author and do not necessarily reflect those of Pima Mining Co. nor of the profession in general. PlMA MILL The following outline of milling practice is presented to furnish a background for later chapters of this report wherein specific problems will be detailed and analyzed. Run of mine ore is crushed in three stages, using conventional equipment to 85% minus 1/2-in. for feed to two 10 x l6-ft rod mills operating in open circuit. Each rod mill discharge is split between two 10% x 13-ft ball mills in closed circuit with three 20-in. cyclones. In addition, a 7 x 12-ft ball mill with one 20-in. cyclone receives part of the discharge from one of the rod mills. This equipment was manufactured by Allis Chalmers and Equipment Engineers. The cyclone overflows are combined, sampled and, after joining the regrind cyclone overflow, are distributed into 12 10-cell Fagergren 66-in. flotation machines. The primary rougher concentrate produced

Jan 1, 1965

By William B. Phillips

The possible sources of potash salts in the United States have been considered from many points of view during the last several years, but it is only within the last two or three months that the situation has become acute. We import from Germany, the only sourcc of large supply, more than $10,000,000 worth of potash salts annually, and this business has been badly crippled by the European war. If there was ever a time in our industrial existence when the necessity of providing for our own needs was more acute than it is now, we are not aware of it. One has only to study the price lists of chemicals sent out by the large supply houses to see the change wrought within a very short time. Whether or not these prices are to maintain for a considerable time, no one can say; but it appears to be probable that we shall not see prices materially lower for a period of uncomfortable length, to say the least. It is not my purpose at this time to discuss the various sources of supply of potash salts that have been suggested or recommended. Some of them might afford a pleasing relief but all of them combined would not affect the situation seriously. If we are to have our own sources of potash salts, they will have to compare, in some efficient manner, with the sources upon which we have drawn during the last 20 or 30 years; i.e., they must exist in minable quantities. It seems to me to be quite idle to speculate upon our domestic sources of potash if those sources are not of a character similar to those upon which we have so long depended. A few thousand or, perhaps, a few hundred thousand dollars' worth of chemically derived potash may be produced under the stress of present conditions, but when we consider that each year demands more than $10,000,000 worth, we must look to underground sources; we must look to mines similar to the famous Stassfurt and Halle deposits. If we cannot mine potash salts, we shall have to do without them for a while.

Jan 1, 1915

By Douglas R. Cook

The urgent need for new base metal ore discoveries was discussed by various authoritative exploration personnel at a CIM symposium on the future of the Canadian mineral industry held at Ottawa in 1967. Data presented at this meeting showed that approximately 150 new mines have started or restarted production in Canada since 1955. Over half of these mines were discovered prior to 1950. The distribution of Canadian ore discoveries during the last 15 years (shown in Fig. 1) illustrates a striking reduction in the discovery rate in recent years in spite of increasing exploration expenditures. These data indicate that $20 million is now spent in exploration for every significant metal deposit found, and evidently the cost of discovery is still rising. Even fewer discoveries are being made in some other important continental areas of the world such as Africa, Australia and South America. The discovery record in the U.S. is thought to be more rewarding than in Canada because the new deposits appear to be substantially larger. Fortunately, most good base metal ore deposits have attractive investment re- turns to justify large exploration expenditures. The magnitude of the exploration expenditures and the increasing need to develop additional metal resources justifies a critical review of the procedures used in the search for ore deposits. Of particular concern is the need to improve the use of exploration techniques in the attempt to obtain better discrimination and efficiency. There is a tendency by inexperienced exploration groups to think of any geological, geochemical or geophysical anomaly as an exploration target and ignore the fact that anomalies can result from many factors other than economic concentration of metals. The desire for a unique and favorable explanation of exploration data often results in too rapid a movement from the reconnaissance to the drilling stage without the use of all the relevant information.

Jan 7, 1968

By D. A. Vaughan

COMPOUNDS in the U-N system have received little attention since the work of Rundle, Baen-ziger, Wilson, and McDonald in 1948.' They described the three nitrides UN2, U2,N2, and UN as the only phases in this system. The region between UN and U,N, was reported to consist of the two phases while the region between U,N, and UN, is a homogeneous solid solution in which the U,N3, structure transforms with increasing nitrogen content into the UN, structure. This report presents the results of an X-ray diffraction study of the phases present in the region of the diffusion gradient on uranium reacted with nitrogen at several temperatures between 500" and 915°C. Several samples of uranium used in U-N gas-metal reactions by Mallett and Gerds were available for X-ray diffraction study of phase composition through the diffusion gradient. X-ray diffraction data were obtained using filtered iron radiation at intervals through the diffusion layer of the several samples on filings or turnings at approximately 0.001 in. increments down to the base metal. Representative X-ray diffraction patterns from various layers through the case on uranium after nitriding at 850°C are shown in Fig. 1. These data show that three different phases are present in the

Jan 1, 1957

By George F. Weaton, Eugene P. Pfleider

13.4-1. History. During the past ten years the mining and production of iron ores has been through a revolution. From 1939 through 1948, which included World War II, 561,000,000 tons of iron ore was mined and shipped from Minnesota's Mesabi Range to the blast furnaces of the United States for the production of iron and steel products (20, 28, 29). The iron mining industry, anticipating the exhaustion of this highly productive source of raw material, instigated a worldwide search for new sources of iron ores. This search was highly successful, both as to quantity and quality. Venezuela (38), Canada (17, 28, 42), Peru (19, 23, 50, 52, 54, 56) Chile, (1, 66, 75), Brazil, and Liberia were among the countries where enormous deposits of high quality iron ores were discovered. The exploitation of these deposits took away the monopoly held by the Lake Superior region for so many years and created a highly competitive market for high quality natural iron ores. 13.4-2. Exploration and Development Practices. SPECIFIC TO IRON MIN- ING. American industries are motivated to continue exploration because new ore discoveries often lead to an improved competitive position with respect to raw material supply. Exploration techniques have advanced rapidly in recent years, to a degree that potentially important iron ore reserves, stratigically located, may now be detectable by application of a combination of refined and diversified geophysical exploration methods. Since the days of the dip needle, iron ore people have placed greatest reliance on the magnetometer. Its successive stages of development progressed from the balance and the torsion type to the fluxgate, and then to the "sophisticated" magnetometers-conceived and developed in th field of atomic physics. By far the greatest advancement in magnetometer prospecting was made when the self-orienting fluxgate magnetometer took to the air. This was not a new development, but it was a breakthrough in reconnaissance investi-

Jan 1, 1968

By D’Arcy A. Shock

In this year's review, the concept of solution mining has been expanded to include total solution as well as the technology of leaching, or partial solution methods. Leaching methods continue to make strides in the processing of metallic minerals, particularly in copper and uranium production. No quantitative data are available on the increase of leach operations; however, it is known that there were more systems installed in 1969 and that the estimated 15% of U.S. domestic production of copper produced in 1968 was by leach methods.1 While the number of publications appearing during the year were few,2,3 there were special sessions in this field at the AIME Annual Meeting and SME Fall Meeting in 1969 and more scheduled for these meetings in 1970. Significant areas that should be mentioned are: the continuing interest in bacteria to convert sulfide sulfur to acid thus reducing reagent costs in both copper and uranium sulfide ores, the exploration of leaching agents other than common mineral acids or alkaline solutions,4,5 kinetic studies of leach mechanisms and the development of microfractures for increased leach solution penetration by use of chemical or atomic explosives.6,7

Jan 1, 1970

By Rudolf Gahl

Some time ago the Detroit Copper Mining Co. had to decide the question whether it would pay to re-treat slime-tailings, and several machines were tested in order to ascertain the type of construction which would give the greatest saving. In previous tests on the ore of this company, several tables proved far inferior to the Frue vanner, and for this reason they did not receive any serious consideration. In the late tests, however, one table was found to surpass the vanner in the percentage of saving made. The Frue vanner used in these tests had been adjusted approximately to the slope which the slime-vanners in the concentrator of the Detroit Copper Co. had at that time, about 2 in. between the inside of the wooden posts, or 0.272 in. per ft. It had not been determined, however, if this slope was the most economical one for treating the slime-feed in the mill, or for re-treating the tailings from the slime-vanners, and I was instructed to ascertain by test-runs, under different conditions, the best adjustments for the work to be done. In order to determine if a change in the adjustments of the machines would improve the results, it seemed sufficient to continue running the tests between the vanner and the table before mentioned, keeping the feed and all other conditions the same, but varying the adjustments of the vanner from run to run. I employed a vanner-man of long experience to operate these machines, who expressed the opinion that a slope between 28 and 2.5 in. between the posts (0.323 and 0.340 in. per ft.) would give the best results. Some runs made this way gave a considerable improvement. While for a slope of 2 in. (0.272 in. per ft.) the saving had been only 40.5 per cent. of that of the table, it was 49.2 per cent. for 2.5 in. (0.340 in. per ft.) slope. A decrease of

Jan 1, 1910

By Effingham Humphrey

THE Lehigh Valley. Coal Co. finished the rebuilding of its Drifton No. 2 breaker at Drifton, Pa., in the summer of 1917. The new construction comprises an addition and the complete remodeling of the old breaker. It is of interest to note that the old structure was the first iron breaker erected in the anthracite region. THE OLD BREAKER This old breaker was built by the Cross Creek Coal Co., under the direction of the late Eckley B. Cole, in 1888-89, and was fully described by him in the Transactions.1 The columns were cast iron, 8,'10 and 12 in. square. The floors and partitions in pockets were cast-iron ribbed plates. The struts were cast iron of "H" section; the other members were of structural iron either rolled I-beams or built-up sections. The sizes of the rolled beams varied from 6 to 20 in. deep, while the built-up sections, made of plates and angles, varied from 6 to 63 in. in depth, so that there apparently was no rule to determine the type of section used. In the remodeling of the old structure, practically everything was taken clown excepting the columns and roof, and such beams as were necessary for temporary construction purposes, and, therefore, a close inspection of the old structural members could be made and the effect of the acid water oil them could be noted. The cast-iron members were but slightly attacked by the water, while the wrought-iron beams, in the wet places, were corroded almost to failure.

Jan 2, 1918