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By John Hall

SOME months ago one of the authors was asked to write a paper on the heat treatment of steel castings that would be more comprehensive than other matter lie had published; this is an attempt to present in condensed form the results of experimental work and actual practice carried out for the most part in 1907, 1908; and 1909. It is not a laboratory research and the work done was carried out for the express purpose of putting in practice the processes discovered as fast as possible. Most of the tests are the result of commercial practice and not of bars annealed in our laboratory; most of the micrographs were made for our own records and not for publication. Certain of the micrographs, therefore, are not as clear as we may wish but they were good enough for our purpose at the time they were made. The tests and micrographs are selected from several thousands in our files. All of these micrographs are magnified 60 diameters. These experiments were conducted almost entirely on steel made in a 3-ton bottom-blown converter, and most of the steel illustrated analyzes about 0.05 in phosphorus and sulfur. Unfortunately, a certain number of specimens are shown of which we do not have the complete analysis, but in most cases in which the phosphorus and sulfur vary much from the figure just given that fact is noted. For purposes of comparison, we have included a certain number of micrographs and tests of steel made in the electric furnace, some of which were from specimens made in other shops and some in our own shop, and some tests of acid open-hearth and tropenas steel.

Jan 9, 1919

By Edward Wisser

The environment of an ore body is taken to mean not only its physical surroundings but every factor, passive or active, that conditioned the ore shoot, saving only the original composition of the solutions that formed the shoot. Collection of data related to the shape and characteristics of veins should be followed by correlation of the data and the formulation of generalizations regarding ore formation. The nature of the criteria used in arriving at these generalizations is discussed. Certain of the criteria, when used with an understanding of their underlying causes, become "ore- guides." In the course of his study the geologist must not confine himself to the immediate surroundings of the veins, but should attempt to work out the geologic history of events culminating in the crucial epoch for any mining district, the period of mineralization. The work may involve study of features far removed from the veins and related to them only indirectly. It commonly means thorough study of the geology of the entire region surrounding the ore deposits. THE following question has been suggested for discussion: [t] "Are there any criteria by which one can determine when mining operations have reached the roots of an ore deposit?" The question deals less with the roots of a particular ore body than with the lower limit of the vertical range within which ore bodies may be expected. It concerns a critical line upon the longitudinal vertical projection of a vein.[t] Below this line it is useless to look for ore shoots; above the line exploration below the bottoms of individual ore shoots may be justified. Can we learn to determine the position of such a line? In some multivein districts the line for any given vein may be approximated by simple interpolation or extrapolation. Where this is possible the bottoms of the lowest ore shoots in the various veins are points on an imaginary, gently rolling surface. The trace of this surface upon the vein is our critical line for that vein. The imaginary surface rolls so gently that it may be projected from two or more known veins to an unexplored vein near by. Here is a criterion for the bottoming of ore that has nothing to do with changes in the appearance of the vein in question; it rests solely on the depth at which the ore bottomed in adjacent veins. The method is not applicable in districts only slightly explored in depth, nor where the lower limit of the ore-bearing horizons

Jan 1, 1939

By Carl Pierce

The writer of this article has not attempted to write a technical paper; on the contrary, he has tried to express in "steel-plant English," for steel men, a viewpoint drawn from his practice and experience. It is not his thought that it shall be taken as the "last word" in the making of this class of steel, or that good rimming steel cannot be made by any other practice. The paper was written to serve as the basis of a discussion from which it is hoped much benefit may be derived. The paper deals only with rimming steel made by the open-hearth process, and with the lower carbon steels, such as sheet and tin bar, rivet stock, skelp, etc., although good rimming steel as high as 40 to 50 carbon can be made by having good open slags and giving proper attention to working and tapping temperatures. Definition of Rimmed Steel Rimmed steel is sometimes called open or effervescing steel to differentiate it from killed or partly killed steel. It is steel made by a process which permits the free escape of gases from the metal in the molds. In passing up through the metal, the gases keep it in a rolling action from the walls of the mold inward. The top of the ingot rims in as the metal chills and the walls of the ingot get thicker, the center part remaining fluid until it reaches a pasty condition when it is covered with a cast-iron cap. The Charge In making up a charge for rimming steel, clean stock of a known composition must be selected, special attention being given to the analysis of the materials. The silicon in good rimming steel must be eliminated to a trace, therefore it is advisable to have a uniform amount going into the charge. Too much silicon gives a raw melting heat from which rimming steel is unlikely to result. For this reason a charge containing not to exceed 40 per cent. pig iron and 60 per cent. scrap, with no iron ore, is preferred to the Monell type of charge in which as high as 65 to 70 per cent. pig iron is used with a proportional amount of iron ore for the removal of the excess silicon and carbon. The silicon content of pig iron varies so much from day to day in the average blast-furnace practice that there is danger of getting too much of this element in the latter mixture when

Jan 1, 1926

By R. H. Landshof, Reynold Q. Shotts, James A. Redding

GENERAL INTRODUCTION by R. Q. Shotts The sizing of coal particles is one of the most important beneficiation operations performed from the time coal is broken at the face until it is delivered to the customer. Indeed, in many cases it may be virtually the only such operation performed in that interval, although it is most often accompanied by some comminution. Sizing, as the term implies, is a separation of the broken coal particles into arbitrary groups, each group containing particles within restricted mass, volume or size limits. Size itself is not a rigidly defined property, but is actually determined more by the operating principles of the sizing device than by unique properties of the particles themselves. Particle size cannot be uniquely defined independently of particle shape. Taggart l says, "In fact, it is not possible to state the size of a single irregular body by means of a single dimension; such possibility exists only for regular polyhedra and their limit, the sphere." In view of the scarcity of regular polyhedra among coal particles, even though most appear to approximate a cube, we actually define size in terms of a surface opening through which the particle barely will or will not pass, or of two openings, the smaller of which retains all particles of the group under study and the larger of which passes all of them. Alternatively, size is expressed in terms of rate of settling of particles in a fluid under closely controlled conditions. Historical Sketch2-Taken from Coal Preparation, 2nd Ed. "Sizing is the process of separating mixed particles into groups of particles all of the same size or into groups in which all particles range between certain definite maximum and minimum sizes. In coal preparation, sizing is generally accomplished by passing the coal over screens. Separation by differential settling in air or water currents has been adapted to the special field of separating very fine sizes, and these processes are in commercial use to a relatively small extent in the coal industry. Separation with screens is a very old industrial practice. Agricola described the use of screens for ore treatment in the middle ages. In the literature of the British coal industry references were made to the screening of coal as early as 1589, but the practice of extensive sizing for the market did not become general until the last half of the nineteenth century. "In America, the practice of sizing developed first in the Pennsylvania anthracite region, beginning with the use of rakes in the mine working places to recover coarse coal. The mining procedure left the fines in the mine.

Jan 1, 1968

By H. C. Goodrich

At the Utah Copper mine, steam shovels were first used, in 1906, for the removal of overburden, and in June, 1907, for the mining of ore. Prior to 1907, the ore came from underground development work and from stoping. The use of the steam shovel gradually increased until 1914; since then all material, both overburden and ore, has been loaded by power shovels operating on the various benches of the mine. During these 19 years, there has been a great improvement in open-pit mining, largely in the use of equipment having greater capacity and dependability. The history of the development of mechanical shovels and related equipment, as affecting the Utah Copper Co., would be interesting and would justify considerable space; but of greater interest at this time is the operation of the equipment now in use. The Utah Copper mine is in a great deposit of copper-bearing monzo-nite, which occurs as a prominent headland of the Oquirrh range of

Jan 1, 1925

By Chas Dobbel

THE Chief Consolidated properties are situated in the Tintic mining district of Utah, being included in Juab and Utah Counties, about 70 miles south of Salt Lake City. The drilling referred to in this paper has been carried on in the Chief Consolidated No. 1 mine, the Gemini mine, and the Eureka Hill mine at Eureka, the Grand Central mine, above Mammoth, and the Apex Standard mine just south of Dividend. The occurrence of ore through the district is typically in lenses on bedding or breaks, and in connecting pipes, without accurate geometrical trend so that the ratio between the total length of development in drifts and crosscuts to the square feet of ore exposed is very high. A section of one of the western channels (Fig. 1) through the Tintic district, shows, roughly, the irregularities calling for excessive development work, so typical of the Tintic mines. This irregular trend of the ore constantly calls for excessive cutting up of the whole area for offshoots and the connecting links always present.. In the planning of prospecting and development, no regular system may be laid out without overlooking some orebodies or segments of orebodies. It was with these facts in mind that Cecil Fitch, the general manager of the Chief Consolidated Mining Co., decided that some method of drilling was the more feasible way of successfully cutting up and bringing about the discovery of ore believed to exist; yet this drilling had to compete with a great. margin of cost over crosscutting, as a crosscut in the Tintic district, even though barren, brings out geologic facts that, in many cases, are worth the price of the working itself. Diamond drills met with the success that might be expected in a district that has a reputation for the wear and loss of diamonds; after 6 or 8 months, this method was discarded on account of the high costs. Simultaneously, a small churn drill outfit was set up underground but a few shifts proved its inapplicability. At the same time, the ordinary hammer drill was usd to drive prospect holes with jointed steel.

Jan 9, 1925

By C. H. G. Bushell, James E. Lawver, William Barbarowicz

Proper use of automatic controls in the process industries is a combination of art and science. Although many phases of automatic control theory can be expressed by precise mathematical relationships, the final control arrangement usually requires considerable effort by trial and error before the optimum control circuit is found. The failure of flotation operators to recognize that control circuits require study and plant experimentation before satisfactory operation occurs has led to a popular misconception that flotation plants simply do not lend themselves to automation. This widespread belief has been overcome to some degree in recent years by the numerous installations of automatic controls that have paid their way by increasing profits in terms of increased recoveries, lowered costs, or higher grade concentrates. WHY CONTROL ? Any form of automation must be justified in terms of ultimate return on in- vestment by improving quality, by increasing plant throughput, or by reducing the total operating costs. Generally speaking, automatic controls rarely re- place a mill operator. They assist the operator in holding important variables within predetermined limits. Large tonnage operations usually can justify a great deal of automation because an increase in one or two units of recovery results in enormous savings compared to the capital and operating costs of controls. DETERMINATION OF CONTROL The first questions in considering plant automation are: 1) What variables must be controlled? 2) What are the optimum levels of the variables? The optimum operating conditions for any process are those which maximize profits. Many of the variables active in flotation processing can be determined by experimentation in the laboratory or pilot plant 10. Although laboratory experiments can pinpoint variables, the true optimum conditions existing at plant level usually must be determined by plant scale experiments. In recent years, the concepts of detailed experimental designs in the laboratory, followed by plant experiments, have reached the stage of maturity where

Jan 1, 1962

Jan 1, 1890

By Thomas H. Leggett

A number of papers on the use of electricity in mining operations have been presented at the Institute meetings recently, but only a very few of them have given particulars and details of the

Jan 1, 1895

NEW MEMBERS The following list comprises the names of those persons who became members during the period Dec. 10, 1916 to Dec. 31, 1916. BOWMAN, Joseph V., Chem., Andes Copper Mining Co., Chanaral, Chile, South America. CALDER, NORMAN L., Min. Geol., Cia. Minera y de Fomento, Watkins, Fomento Prov. de Santa Clara., Cuba. CONRADS, RALPH A., Met. Engr., Magna Plant, Utah Copper Co., Garfield, Utah. FARR, T. H. POWERS, JR. Asst. Assayer, Argonaut Mining Coy Jackson, Cal HAURY, PETER S., Min. Engr ...Alta Loma, Cal. HOULE, ARTHUR, Supt., Shattuck-Arizona Copper Co., Box 2009, Bisbee, Ariz. LEIB, VICTOR E?, Min. Engr 4204 Ave. D, Austin, Tex. LINDSLEY, THAYER, Min. Engr 82 Beaver St.,. New York, N. Y. POTTER, OCHA Supt., Superior Copper Co., Houghton, Mich. ROCCA, BERNARD T.; Engr., Bluebell Mine, Cons. Arizona Smelting Co., Mayer, Ariz. SAFFOLD, RAY P P.'0. Box 436, Douglas, Ariz. THOMS, JAMES EZRA, Asst. Chief, Milling, -Roasting, Acid Depts., Mineral Point Zinc. Co., Depue, Ill. TOMITA, HARUYOSHI, Met. Engr., Mitsubishi Co., Osaruzawa Mine, Akita-ken, Japan.

Jan 2, 1917

By Robert Glass Cleland

THE THIRD and most important item on the Phelps Dodge agenda for the development of new ore supplies involved the site of the company's initial venture into Arizona, the old mining camp of Morenci. By the purchase of the Arizona Copper Company's holdings in 1921, Phelps Dodge had obtained control of virtually the entire Morenci district, including town, public utilities, and, above all, the huge mineralized area that had once been divided between numerous competing companies. Because of this rigid concentration of ownership in the hands of one organization, Phelps Dodge was able to plan and successfully carry out a program of greater magnitude than any Arizona mining company had previously under- taken. The immense deposit of low-grade ore, usually referred to as the Clay Orebody, had been known to the mining companies of Morenci for many years. Pioneer exploratory work, both by the Arizona Copper Company and later by Phelps Dodge, revealed that the deposit, mostly of light-gray porphyry rock, lay under a hard capping from two hundred to five hundred feet thick. An intensive drilling

Jan 1, 1952

By J. J. Marcus

Mineral industry decision-makers are frequently required to select between competing equipment. Various criteria and methods are currently being used, and this writer would like to suggest a method recently tested on two large power shovels in an open pit mine. An initial examination of both shovel A and shovel B indicated that the latter should give the best overall results. since: 1) It swings 13% faster; 2) it carries a dipper with 23%: greater capacity; 3) purchase price is 8.6% lower; 4) shovel weight is 12.5% greater; 5) it has considerably greater bail pull under most operating conditions; 6) the electric controls are of a more modern design. However, shovel B has half the service life. With these points in mind, shovel A was selected by The Buckingham Method, an operations research technique.

Jan 9, 1965

By William H. Case

The Bertha zinc-mines of the Bertha Zinc and Mineral Company are in that part of the State of Virginia popularly known as Southwest Virginia, and are included in that division of the " Great Valley " named in recent years the " New River-Cripple Creek Mineral-Region of Virginia." The mines lie wholly in the southeast part of Wythe county, two miles from Pulaski county on the east, and three miles from Carroll county on the south. The only other zinc-mines that are opened in the region are those of Manning & Squier, which adjoin the Bertha zinc-mines on the northeast, and the mines of the Wythe Lend and Zinc Company, at Austinville, eight miles to the southwest. A little farther southwest of Austinville, near Ivanhoe, a new company is preparing to mine zinc-ores. The mining properties named are included within a distance of ten miles along a direct line, bearing north 55' east, on the south side of New river, which here flows northeastward in a tortuous course. The greatest distance of any of the mine-openings from the river is a little less than two miles. It is not the magnitude of the operations in zinc-mining in this particular region, as compared with that of other well-known districts, that merits especial notice. Aside from the exceptional purity of the spelter made directly from the Bertha ores, some peculiar features of the zinc-ore deposih, as shown at the Bertha mines, and

Jan 1, 1894

IN CHAPTER I it was shown that the approximate market value of the copper produced at the twelve big Porphyry mines from 1905 to 1931, inclusive, was $2,821,300,000. This represented the refined metal ready for shipment to the consumer. In addition, at least $50,000,000 worth of gold and silver had been produced, more than $30,000,000 coming from the Utah mine alone. Consequently, the revenues totaled about $2,871,300,000, as shown in Table 58. Society paid this sum for a given quantity of metal that it wanted for industrial use. Where did the money go? Who were the direct beneficiaries of the creation of this new wealth? These questions this chapter will answer in a general way. Because some necessary data are not available, an exact tabulation in detail can not be made, but Table 58 is believed to reflect with reasonable accuracy the disposition made of the aggregate income from the twelve operations. The word "operations" is used advisedly because the complexity of corporate relations precludes the possibility of picking out companies, as could be done if each enterprise were an independent entity such as Miami, for example. Some explanation of the table and some elaboration of the items under "Outgo" is desirable. In the first place, it is assumed that the sum of surplus current assets existing in various treasuries at the end of 1931 was substantially equivalent to the “new” money that actually was put into the enterprises as a group. Less than $20,000,000 in new capital was realized from securities sold by the Utah Copper Co., the remaining funds coming from earnings. In respect to some of the others, such as Chuquicamata, Inspiration, and Andes, the amount of new money was much greater. However, of the two items-total new capital invested and total current surpluses-the former

Jan 1, 1933

By W. G. Whildin

This paper will be confined to a discussion of the methods in use in the property of the Lehigh Coal & Navigation Co. in the Panther Creek valley. Only the methods used in mining the Mammoth vein will be .considered, as the methods in use in the smaller veins are adaptations of the same. Typical cross-sections of the coal basin are shown in Figs. 1, 2, and 3. The Mammoth vein varies in thickness. It is 21 ft. at one colliery, 50 to 125 ft. at another, and over 200 ft. at still another. Its normal thickness is 35 to 40 ft. At some points in the Panther Creek basin the Mammoth vein is made up of three splits-the Bottom, Middle, and Top, while in other portions the dividing strata between the benches amount to only 1 ½ to 2 ½ ft., and the vein is practically in one. The section of the Mammoth vein shown in Fig. 4 was taken at the Greenwood colliery, and shows a thickness of 60 ft. The miners know it by benches, as follows:. The Three-Foot, the Four-Foot, the Eighteen-Inch Slate, the Bony, the Grey Slate, the Grain Clear, the Five-Foot, the Seven-Foot, the Slaty or Dirty benches, the Blue Slate, and the Top bench. The present-day methods of mining are an outgrowth of the experiences of the last 100 years in this territory. The first attempt at mining, according to the company's maps, was made at Summit Hill in 1792, when open cuts and pits were made. Later the pillar-and-breast method was introduced. This and other methods will be discussed hereinafter. Drifts, slopes, and shafts, according to the requirements and physical conditions, are the main openings, from which the gangways are driven. The gangways (haulageways) and airways are driven along the strike of the vein; the gangways in the various veins on the same level being connected by tunnels through the intervening rock. The gangways and airways in thick veins are heavily timbered ;throughout. The sets of timber, consisting of a collar and two legs, are spaced 5 ft. apart, and later, after the ground :has settled, additional sets, called "liners" or "relief sets," are placed between, to relieve the strain on the original timbers, so that the result is a set of timber every 2 ½ ft.

Jan 12, 1914

By M. Marciano, K. Kreider

Silicon carbide coated boron fiber (Borsic) reinforced aluminum composites were made which exhibit strength and modulus values predicted by the rule of mixtures. A successful technique for fabricating these composites consists of plasma spraying of monolayer filament reinforced tapes and subsequent diffusion bonding of these tapes into composites. The elastic properties of the composites were given by the rule of mixtures in reinforced directions. (The modulus parallel to the fibers was 32 x 106 psi for a composite with 50 pct by volume fibers. The other elastic properties for a composite with 50 pct fiber were transverse modulus = 14 to 16 X 10 psi, G = 9.5 x 106psi, and v,, = 0.23. Tensile strengths of over 190,000 psi were measured for the composites with 50 pct fiber and the average strength for these composites was 162,000 psi. These same composites had interlaminar shear strengths of 12 to 15,000 psi. Multidirectionally reinforced composites demonstrated the same reinforcement efficiency in strength and modulus as the uni-directimlly reinforced composites. Comrpessive strengths of composites with 50 pct by volume fiber were found to be greater than 250,000 psi. STRUCTURAL filamentary composite materials are being considered for use where high strength, high modulus, and lightweight materials are required. Metal matrix composites have been considered for use at temperatures ranging from cryogenic (aluminum stainless steel) to 6000°F (plasma sprayed tungsten plus tungsten fiber) and other requirements have been similarly diverse. Metal matrix composites offer high modulus and high strength in unreinforced directions, the ability to be joined by brazing or welding, and greater fracture toughness compared with the more easily fabricated resin matrix composites. This paper deals with the mechanical properties of aluminum boron composites, a system which is being considered for applications where lightweight, high modulus, and high strength are required at temperatures including those above the normal operating temperatures of resin composites. The boron fiber used in this investigation had an average room temperature strength of greater than 400,000 psi, a modulus of 58 x 106 psi, and a density of 0.09 lb per cu in. (2.6 g per cu cm).' The fiber used in the composites had a coating of silicon carbide to retard degrading reactions with the matrix at elevated temperatures. The coated fiber, which is sold commercially as Borsic: has been reported to have excellent resistance to degradation at temperatures up to 600°C in air and in aluminum2 and the same mechanical properties as the standard bare boron fiber. Fabrication of Composites. The composites were diffusion bonded from plasma sprayed monolayer tapes. These tapes were fabricated on a substrate of aluminum alloy foil which is incorporated in the composite matrix. The technique consists of winding a layer of fiber on the foil, and plasma spraying the balance of the matrix alloy over the windings which bonds all of the components together. Advantages of this process include the ability to achieve good fiber spacing, Fig. 1, the formation of a good bond between fiber and matrix (achieved during the plasma spraying), and the adaptability of the process for forming large and complex parts. This is particularly true when used with brazing foil backing which alloys low pressure bonding of the structural shape. PROPERTY EVALUATION TECHNIQUES Tensile test specimens, 5 to 7 in. long with a reduced gage section and ,030 to ,040 in. thick, were cut from composite sheets with a diamond saw. The gage section was .200 in. wide, while the gripped section was .240 in. wide and over 2 in. long in each grip. The standard gage length used was 1 in., however, transverse specimens (i.e., specimens with the fibers oriented 90 deg to the loading axis) were tested without a reduced section. Twelve specimens were also made with a 3-in. gage length but no significant effect of gage length was measured. The tensile test specimens were mounted in a fixture and aligned in friction grips using a 10X microscope. Unbonded foil doublers (0.010-in.-thick aluminum) were used to insure firm gripping with a minimum of damage to the fibers. The test specimens were transferred to the testing machine in a rigid fixture and mounted in the self aligning loading train. Testing was performed using a Tinius Olsen four screw testing machine with a torsion bar LVDT load

Jan 1, 1970

By Robert M. Grogan, Gill Montgomery

Fluorspar, the commercial name for fluorite, is a mineral composed of calcium fluoride, CaF,. Its valuable properties are due to its content of fluorine, and it is the principal commercial source of that element. The name is derived from the Latin word "fluere," to flow, because the mineral has a low melting point and is used in metallurgy as a flux. Cryolite is a sodium aluminum fluoride, Na,AlF,. It is a rare mineral which has been found in commercial quantities in only one place in the world, Greenland. Small amounts of it are shipped from stock in that country, and its place in industry has been taken over almost completely by synthetic cryolite. Because of the relative unimportance of cryolite, discussion of it has been relegated to a short section at the end of this chapter, which is principally concerned with fluorspar. History Fluorspar was used for ornamental purposes by the Greeks and Romans, who fashioned such things as vases, drinking cups, and table tops from it. Various peoples including the Chinese and the American Indians carved ornaments and figurines from large crystals. Its usefulness as a flux was known to Agricola (1564), who wrote about it in 1546. Mining began in England about 1775 and at various places in the United States during the period 1820 to 1840. Production became substantial following the development and widespread adoption of the basic open hearth method of making steel, in which fluorspar is used as a flux. After that, production and usage were stimulated by growth in the steel, aluminum, chemical, and ceramic industries and were accelerated by World Wars I and II. Fluorocarbons entered the picture in 193 1 and the use of anhydrous hydrogen fluoride (HF) as a catalyst in the manufacture of alkylate for high octane fuel blends began in 1942. The development of the froth flotation process in the late 1920s, a differential flotation scheme for separating fluorspar from galena, sphalerite, and common gangue minerals in the 1930s, and the application of heavy-media concentrating methods to the treatment of low grade ores in the 1940s were outstanding technological contributions that enabled increased production of fluorspar concentrates. In recent years the development of pelletizing and briquetting processes for making flotation concentrates into gravel-size particles for use in steel furnaces and the development of more efficient and selective flotation schemes for treating crude ores containing abundant dolomite and barite have been major improvements in the industry.

Jan 1, 1975

By M. R. Geer, H. F. Yancey

WITH the exception of pneumatic processes and ii few special beneficiation methods of comparatively limited application, all mechanical coal-cleaning and mineral- dressing processes involve the admixture of solid particles with a liquid, generally water. Upon completion of the beneficiation process the solids must be reclaimed from the water. With coarse particles separation of solids from water is accomplished readily by the simple expedient of screening, but recovery of fine solids presents a much more difficult problem. Thickening and dewatering are practiced so generally, both within the mineral industry and else- where, that they have been the subject of study through the years and a number of processes for making the solid-liquid separation have been developed. Dorr-type thickeners have been standard equipment for many years, and more recently driers employing centrifugal force have gained in popularity. 1 A cyclone device employed to thicken the loess suspension is a novel feature of the loess heavy-medium coal-cleaning process developed by the Netherlands State Mines in the thirties. The cyclone thickener was first mentioned by Driessen2 in 1939 in a paper describing the loess process and has been the subject of two more recent reports by the same author.3 The cyclone does not differ greatly from cyclone dust collectors used in this country for years, except that it is operated wet rather than dry. Advantages of the cyclone over other thickening devices are its simplicity of construction, high capacity, and the fact that, having no moving parts, it is simple and inexpensive to operate. The Bureau of Mines became interested in the cyclone after one of the present authors observed it in operation in the Netherlands in 1945. The first experimental work with the cyclone conducted by the Bureau of Mines was an investigation of the unit functioning as a heavy-medium cleaner for fine coal-a use for the cyclone that was conceived long after it had been perfected as a thickener. The results of this investigation were published last year.4 The present investigation deals exclusively with operation of the cyclone as a thickener. OBJECT AND SCOPE OF INVESTIGATION Although the cyclone thickener has been dealt with in three reports by Driessen, its use as a thickener was not the principal subject of any of these reports and therefore neither the cyclone itself nor its operation was described completely. The design of the cyclone was described only in a very general way, only limited operating results were given, and virtually nothing was disclosed about the influence of the various possible operating variables. Moreover, the Netherlanders apparently had employed the cyclone exclusively for thickening loess and had not employed it on coal slurries. One English colliery has used a cyclone to thicken coal slurry,5 and as it appeared that

Jan 1, 1948

By W. S. Reid

In March, 1938, E. P. Fleming, metallurgist for the American Smelting and Refining Co. inaugurated an investigation into the possibilities of recirculating the gases from Dwight-Lloyd sintering machines operating on lead charge, with the twofold object of concentration of the SO2 content and reduction in volume of total gas produced. The possibility of recovering a commercial grade of SO2 gas from D & L machines operating on lead charge had previously been considered by several investigators. Early History of Recirculation The Selby Smelter Commission Report, published by the Bureau of Mines in 1914, contains a chapter by A. E. Wells regarding results obtained at Selby, wherein some of the richer gas was recirculated through a hood over the pallets. Oldright and Miller of the U. S. Bureau of Mines had also made tests at Trail, B.C., and at Kellogg, Idaho. R. C. Rutherford, while at the Chihuahua, Mexico, Smelter of the American Smelting and Refining Co., in May, 1937, proposed recirculation of D & L gases to decrease the volume of gas handled by the baghouse. At none of these plants, however, was the operation commercialized. In July, 1938, Mr. Fleming, in correspondence with the Selby Plant, inquired regarding the possibility of obtaining 6 pct SO2 gas from the Selby D & L machines. At that time, the writer advised that there was slight possibility of obtaining 6 pct SO2 gas without re- circulation, but believed that it was possible with recirculation, and that experimental work toward that end should be tried at some plant where spare D & L machines were available. The foregoing statement was based on the following information then available— 1. Tests on Selby first-over machines showed 2.28 pct SO2 from first windbox and 1.03 pct SO2 from second windbox, and corresponding figures for second-over charge of 0.81 pct SO2 for first windbox and 0.08 pct SO2 for second windbox. 2. Oldright and Miller (US. Bureau of Mines) in 1932 at Bunker Hill, on 42 in. X 22 ft machines found: a. First-over charge—Maximum SO2 concentration (leaving cake) of 9.5 pct. b. First-over charge—SO2 concentration of over 8 pct from the 4 ft to the 12 ft points beyond the front dead-plate and that the concentration then dropped rapidly. c. That approximately 80 pct of the total sulphur eliminated on the second-over machines occurred during the travel of the pallets from the 1 ft to the 6 ft distances from the dead-plate. d. That approximately 94 pct of the total sulphur eliminated on the second-over machines occurred over the first windox. 3. Oldright and Miller (US. Bureau of Mines) in 1932, at Trail, on 42 in. X 50 ft machines found: a. First-over charge—SO2 varied from 2.0 pct to 5.5 pct (leaving cake) from the 12 ft to 28 ft points from the deadplate. b. That the average SO2 increased from 1.0 pct at 7 ft from dead-plate to maximum of 3.3 pct at 20 ft, then dropped to 1.5 pct at 40 ft. c. That on a 22 ft, second-over machine with an 11 in. bed the SO2 varied from 4.5 pct to 6.5 pct from the 2 ft to the 8 ft points from the dead-plate. d. That on the final roast, the SO2 concentration also varied across the pallets; that is, 2 1/2 pct at the side, increasing to 6.0 pct, 5 in. in, and to 7.0 pct at 10 in. to 20 in. in, then decreased vice versa at the opposite side. e. Concluded that most of the sulphur on a 22 ft machine was removed over the first 7 ft of the first windbox; therefore, they partitioned the first windbox so that the exit gas from the second 4 ft section was returned to the surface of the pallets over the fist 7 ft section, and during a seven-day trial the gas from the 4 ft section averaged 2.4 pct SO2, while the recirculated gas from the first 7 ft section only increased to 3.8 pct SO2. (Excess suction over the 7 ft section to prevent escape of SO2 laden gas from the 4 ft section caused dilution by air.)

Jan 1, 1950

By C. A. Stout, S. W. Nicksic

The materials that stubilize emulsions of some Southern California crude oils were isolated from the produced crude. These substances were separated into acidic and nonionic fractions by ion exchange chromatography. Each fraction was further broken into groups of homogeneous molecular weight composition by gel permeation chromatography. Both ionic character and molecular weight of the individual fractions are reflected in the properties of the original crude oil emulsion. Specifically, the more acidic and the higher molecular weight fractions appear to be the most effective stabilizers. INTRODUCTION The treatment of oilfield emulsions is a continuing problem for the oil industry. In almost all cases, these are water-in-oil emulsions. Much more is known about the formation and stabilization of oil-in-water (O/W) emulsions than of the water-in-oil (W/O) type. To understand W/O emulsions, more information is needed on the materials responsible for their formation and stabilization. This is especially true in the case of oilfield emulsions where the stabilizing materials are complex mixtures of large asphaltene-like molecules. To study the chemical and physical properties of these emulsion stabilizers, it is necessary first to separate them from crude oil. This separation was achieved in the case of two Southern California crude oils. The materials that were isolated appear at first to be little different from an asphaltene mixture. However, if a mixture of distilled water and mineral oil containing about 1 percent of the natural emulsion stabilizers is passed through a colloid mill, the W/O emulsion formed may be stored in the laboratory for many months without any sign of breaking. This could not be done using undifferentiated asphaltenes from the same wells, indicating that there is a real difference between these materials and ordinary asphaltenes. The colloid mill used to generate the emulsion was required because tile natural emulsion stabilizers had very little effect on oil-water interfacial tension. The one characteristic — complexity of composition — that emulsion stabilizers share with asphaltenes makes virtually impossible the exact structure determination of the components of the mixture. For this reason, it was decided to investigate first the influence of molecular weight on the ability of these materials to stabilize emulsions. The emulsion stabilizers were first separated into acidic and neutral fractions according to whether or not the material was retained on a weakly basic ion exchange column. The extraction procedure was not designed to remove basic components from tile oils, and no attempt to isolate such a fraction was made. Following this separation, selected fractions were analyzed by gel permeation chromatography. Gel permeation separates according to molecular size, largely independently of chemical or polar characteristics. A heterogeneous mixture is separated into a number of fractions having much more homogeneous molecular weight distribution. A number average molecular weight of each fraction is therefore close to the true molecular weight of the principal component of that fraction. The fraction weights, expressed as weight percent of the sample, and the number average molecular weight of a representative selection of the fractions, will be referred to in this paper as the molecular weight profile of the original sample. DESCRIPTION OF GEL PERMEATION ANALYSIS The first reported gel permeation analyses were for separating partially hydrolyzed protein molecules in an aqueous system using expanded starch granules. Most of the theoretical work on gel permeation has been done by biochemists.1 This theoretical treatment is still incomplete, but the

Jan 1, 1969