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By Arthur F. Taggart

RECENT jig practice has shown such marked departures from the pronouncements of the textbooks, particularly as to particle size recovered and size range of feed, as to make it desirable to reexamine the mechanism of jig action. This paper presents the results of such a study. The general conclusions of the paper are: I. A jig bed is a bi-functioned as well as bi-parted body. The upper layer comprises a plastic stream moving slowly in a substantially horizontal direction from feed lip to tail board. It serves to rough out tailing and light middling and to pick up heavy middling and lift it out of the jig box. The lower layer acts as the ultimate separating means by absorbing concentrate and rejecting heavy middling. 2. The separating function of the lower layer is performed in three definitely different ways, depending upon the relative sizes of the grains presented to it and those composing it. EXPERIMENTAL The experimental work consisted of close and long-continued visual observations of operating beds in a plunger jig, eccentric driven, and in determinations of the settling rates of various metal spheres through a quartz bed under various conditions of operation. Falling velocities were measured by an ingenious method devised by A. C. Dorenfeld, a student in the Mining Department at Columbia. Spheres of different metallic alloys, Y8-in. diameter, were cast with a short piece of fine copper wire frozen in to serve for suspension. A short piece of black cotton thread was looped to the top of the wire to serve as an indicator. The set-up for a test is shown in Fig. I. The thread should be boiled, if necessary, to make it freely water absorbent. The piece of clean plate glass D is supported rigidly at the proper height in a vertical position. The white scale E is similarly supported behind the glass. At the beginning of a test, D is filmed with water by holding it (best with clean rubber finger cots) under running tap water. It is then clamped in approximate position and C, already wet, is laid across it and drawn straight and vertical with the upper end at or near the top of D, and the ball A slightly above the top of the expanded bed. The water film holds the string to the glass. For a test the string is released, its upper end position noted just after the top of the ball becomes submerged, when the corresponding time is read, and thereafter simultaneous readings of position and time are taken as desired. Results of the visual studies of the bed are given at the pertinent places in the following discussion. Results of the settling tests are summarized in Figs. 2 and 3. DEFINITIONS The name bed as used herein refers to the entire mixture of solids and water filling the jig box. A layer is a stratum of a bed composed of particles of substantially the same specific gravities. A granular product is long-range when it approximates a crushed or ground product limited as to upper size only; a shorlrange product is one closely sized. A bed is loaded when operating with continuous feed

Jan 1, 1943

By J. W. Spretnak, V. Kandarpa, G. W. Powell, R. A. Erickson

Neutron-diffraction pattens were obtained at room temperature from two austenitic manganese steels, oxc wth n carbon content of 1.23 zct PC/ and the olher 0.63 wt pct. Analysis of the data showed that the curOan atols (ions) occupy tile octaheadral site. any of the cavbon is in the tetahedral site, tile concentration is too strzall lo be detecled by tzeulr-on d$facton. AS part of a long-range investigation of eutectoid decomposition, a study of the structure of austenite has been undertaken in order to characterize as best as possible the parent phase in the Fe-C system. Austenite is a solid solution of carbon atoms in fcc iron, the carbon atoms being located in interstitial positions. Of the two kinds of interstitial sites, octahedral (0, +, +) and tetrahedral (4, a, $), in this lattice, the larger octahedral sites are usually considered as the probable position of carbon. The strain energy involved in pytting a neutral carbon atom (covalent di; ameter 1.54)' into a tetrahedral site (diameter 0.56A) is considerably greater than that associated with p2tting carbon into the octahedral site (diameter 1.08A). Dayal and arkeen' observed the migration of carbon in iron under the influence of an electric field indicating tha tye carbon is ionized. pe ionic radius of C'4 is 0.15A and that of C" is 0.29. Consequently, it is conceivable that carbon ions could occupy the tetrahedral sites if one considers only the strain energy. Size considerations alone do not provide a complete explanation for the formation of the interstitial solid solutions, however. because metals such as copper, zinc, and cadmium do not appear to form interstitial solid solutions in spite of their large interstitial sites. Evidence which suggests partial ioni-zation of the carbon in austenite has also been obtained by Speiser, Spretnak, and alor who derived formu- lae relating the observed change in lattice parameter to the effective diameter of the solute atom. Assuming the carbon occupies the octahedral site. they computed the effective diameter of carbon to be 1.33A whereas the covalent diameter is 1.54A. petch5 concluded from powder X-ray diffraction experiments that carbon atoms occupy the octahedral sites. His conclusion was based on the general trend observed in the intensity of the diffraction lines as the carbon content of the austenite was increased. However, the major difficulty associated with application of X-ray diffraction to this particular problem is the relatively poor scattering power of carbon. Williamson and smallman' utilized X-ray diffraction to study the line broadening caused by the strain field about carbon atoms inferrite: these investigators concluded that at least 85 pct of the carbon in ferrite occupies the octahedral (0: 3, B) site rather than the larger

Jan 1, 1967

AT THE dinner at the Cosmos Club in Washington, tendered to Herbert Hoover on Jan. 5, on the occasion of his retirement from the presidency of the Federated American Engineering Societies, he made an informal speech in which he pointed out not only the unique position which the engineer now occu-pies by virtue of the new Federated Societies but also the opportunity that is thus presented for national service. In this connection and in pointing out some of the reasons for the present economic maladjust-ment, Mr. Hoover said in part: This body has entered upon a path of public serv-ice that is unique in the whole United States. The engineer has devoted himself all these years to the upbuilding of the material values of the United States. He has done that effectively more effectively than-has been accomplished in any other of the great countries of the world. The birth of this society marks the evolution of the engineer into an interest in public affairs. With his intelligence, his experience and training and the unique knowledge that he possesses not only of the material but the intangible values among our people, he is now organized so that his united voice may become heard outside of his profession. We have probably 175,000 engineers in our country, representing an intellectual possibility for service possessed by no other group. It is therefore an augury of real social development when the engineers of the United States join together for purposes of public service.

Jan 3, 1922

By H. A. Doerner, F. D. DeVaney, J.B. Clemmer

Production of magnesium metal in the United States during the past decade has increased from less than 600,000 lb. in 1928 to more than 4,800,000 lb. in 1938.1 The growing industry has stimulated interest in methods of production and in sources of raw material. All of the present domestic production is from brines. However, the availability of large tonnages of magnesite in the state of Washington, together with the possibility of cheap power, have created interest in the utilization of magnesite for the production of magnesium metal as well as for other purposes. Magnesite, the carbonate of magnesium (MgCO3), usually is associated with limestone (CaC03) and dolomite (MgCO3. CaC03). It occurs in crystalline and microcrystalline (sometimes termed amorphous) forms.2 The more common crystalline mineral is found in various shades of white, gray, yellow and brown, the colors being due to impurities, chiefly iron. Magnesite has a hardness of 3.5 to 4.5 and a specific gravity of 3.0 to 3.1. It forms rhombohedra1 crystals in the hexagonal system, having a vitreous, pearly luster. The quarries and mines of the Northwest Magnesite Co. near Chewelah, Wash., are large producers of magnesite for refractories. The deposits are extensive, some veins being hundreds of feet wide. The boundaries of the veins are indefinite, and the magnesite is interfingered with dolomite. Calcite, talc, serpentine, quartz and shale are disseminated through the deposits. The quality of the ore cannot be determined by appearance or by simple tests; chemical analysis is required. These features make it difficult to determine the boundaries or total quantity of workable ore.3 Only the higher-grade ores are mined, and a considerable proportion of the magnesite content is discarded by hand-sorting and log-washing methods. After calcining, the sorted product meets the specifications for usual refractory material, as it contains approximately 81 per cent MgO, 7 per cent SiO2, 6 per cent CaO and 6 per cent Fe203 + A1203. A sample of the reject material had the following analysis after calcining: 65.4 per cent MgO, 6.3 CaO, 21.1 SiO2 and 6.5 R2O3. The percentage reported as R2O3 includes compounds whose chlorides are acid-soluble but give precipitates in ammoniacal solution that are ignited and weighed. The more common constituents are Fe2O3, A1203 and TiO2. The sorted magnesite product is not the best grade. even for refractory purposes, and leaves much to be desired for a reduction process. The rejected material contains a large proportion of magnesite. It is apparent that an effective concentration method would yield a higher recovery and a better product. In general, the refractory grade is not suitably pure for other uses. Recent developments in the production of

Jan 1, 1943

By J. B. Clemmer, F. D. DeVaney, H. A. Doerner

Production of magnesium metal in the United States during the past decade has increased from less than 600,000 lb. in 1928 to more than 4,800,000 lb. in 1938.1 The growing industry has stimulated interest in methods of production and in sources of raw material. All of the present domestic production is from brines. However, the availability of large tonnages of magnesite in the state of Washington, together with the possibility of cheap power, have created interest in the utilization of magnesite for the production of magnesium metal as well as for other purposes. Magnesite, the carbonate of magnesium (MgCO3), usually is associated with limestone (CaC03) and dolomite (MgCO3. CaC03). It occurs in crystalline and microcrystalline (sometimes termed amorphous) forms.2 The more common crystalline mineral is found in various shades of white, gray, yellow and brown, the colors being due to impurities, chiefly iron. Magnesite has a hardness of 3.5 to 4.5 and a specific gravity of 3.0 to 3.1. It forms rhombohedra1 crystals in the hexagonal system, having a vitreous, pearly luster. The quarries and mines of the Northwest Magnesite Co. near Chewelah, Wash., are large producers of magnesite for refractories. The deposits are extensive, some veins being hundreds of feet wide. The boundaries of the veins are indefinite, and the magnesite is interfingered with dolomite. Calcite, talc, serpentine, quartz and shale are disseminated through the deposits. The quality of the ore cannot be determined by appearance or by simple tests; chemical analysis is required. These features make it difficult to determine the boundaries or total quantity of workable ore.3 Only the higher-grade ores are mined, and a considerable proportion of the magnesite content is discarded by hand-sorting and log-washing methods. After calcining, the sorted product meets the specifications for usual refractory material, as it contains approximately 81 per cent MgO, 7 per cent SiO2, 6 per cent CaO and 6 per cent Fe203 + A1203. A sample of the reject material had the following analysis after calcining: 65.4 per cent MgO, 6.3 CaO, 21.1 SiO2 and 6.5 R2O3. The percentage reported as R2O3 includes compounds whose chlorides are acid-soluble but give precipitates in ammoniacal solution that are ignited and weighed. The more common constituents are Fe2O3, A1203 and TiO2. The sorted magnesite product is not the best grade. even for refractory purposes, and leaves much to be desired for a reduction process. The rejected material contains a large proportion of magnesite. It is apparent that an effective concentration method would yield a higher recovery and a better product. In general, the refractory grade is not suitably pure for other uses. Recent developments in the production of

Jan 1, 1943

By Tracy L. Atherton

THE mapping of slant-hole oil fields is complicated by the fact that relationships between wells are subject to variations in three dimensions and are not readily adaptable to representation on a plane surface. Formalized perspectives usually expand into an undecipherable maze of lines and peg models with the necessary supporting dividers serve only to awe the beholder with their perplexity. The usual compromise arrived at is a plotting of the projection of the courses of all of the wells in the field onto a horizontal plane (Fig. I). These complicated plots, when used in conjunction with the individual vertical projections of the wells, heretofore have been the most practical foundation for the study of general field problems. This method has many shortcomings, however, which are not found in the recently developed straight-hole equivalent projection method. The new method supplies a series of maps similar to the type now in common use in straight-hole fields. It is remarkably simple to use and supplies a ready check on clearances between wells. Perhaps the method can best be understood if the steps that led to its development are described. Common practice of oil-well surveyors is to supply the owner with a tabling of the survey and a plan view plotted against the true meridian. If a vertical projection is plotted, it is usually plotted separately and shows vertical depths against the projection of the stations along a section line joining the top and bottom of the hole on the plan view. A more practical method would be to plot the section line of the plan parallel to the abscissa of the vertical projection (Fig. 2). If this is done, any line parallel to the "base line" will pass through the points on each projection that represent a single point in the hole. Another very useful graph, which the well surveyor could provide, would be plotting of vertical depths against measured depths. Fig. 3 illustrates the next step facilitated by this form of plotting. The lines designated "well A" represent the proposed course of a hole to be drilled. Superimposed on the drawing, in its proper relative position, is the plan view and a portion of the vertical projection of a present producer, "well B." From an inspection of the plan view we arrive at the correct conclusion that the only indicated possibility of collision is at the point of crossing of the wells on the plan view. Referring to the vertical projection, note that the vertical depths of stations on B, in the vicinity of the crossing, are plotted on the same plane of projection as A. The clearance is the indicated distance between the two vertical sections measured along the parallel to the "base line," which passes through the point of intersection of the plan view. Repeat this process for the other wells crossing the proposed course of A on the plan view and you have a preliminary check of clearance for the proposed course. A similar comparison of crossings on the vertical projections of all

Jan 1, 1941

By J. L. Carne

ORE control is a matter of planning and supervision based on a foreknowledge of the content and distribution of ore. The Inspiration orebody is predominately a copper-sulphide blanket, overlain by an oxidized zone of copper silicates, copper carbonates, and barren capping. Most of the ore is treated by leaching, and the optimum requirement of this method is the goal toward which the mining schedule is aimed. Both open-pit operations and underground mining are used. The problems of ore control at Inspiration are: I—maintenance of metallurgical balance, 2—conservation of oxide ore, and 3—mining out of the orebody at an average grade which parallels that of the reserve figure. Prior to the fall of 1926 all of the Inspiration division ores were treated by flotation, but the mixed ores of the Live Oak—Keystone section carried too large a proportion of copper in the form of chrysocolla for satisfactory flotation. For this type of mixed ore the leaching process now in use had been worked out, and the plant was put into operation in the fall of 1926. The process is known as the ferric sulphate leach, and the leaching solvents contain both ferric sulphate and free sulphuric acid. Sulphide copper in the form of chalcocite is leached by the ferric sulphate, and the oxidized copper minerals are dissolved by the sulphuric acid. From the beginning of operations to the present time the plant has treated some 67 thousand tons of ore having an average copper content of 1.142 pct, of which 0.605 pct has been in the form of oxidized copper minerals. The plant is currently being operated at the rate of 11,500 tons per day, 7 days per week. However, this tonnage is produced in 6 days of mine operation. The average current grade is approximately 1 pct copper, of which about 50 pct is present as oxidized or acid-soluble copper. Close control of metallurgical procedures is a vital factor in operation of the plant and is in turn dependent upon careful regulation of feed. An excess of sulphide copper requires the presence of more ferric sulphate in the leaching solvents. On the other hand, an excess of ferric sulphate in the presence of a low sulphide feed seriously reduces the efficiency of electrolytic precipitation in tank house operations. This condition must be avoided at all costs. An additional factor must be considered. It is known that the sulphide content is predominant in the ore which remains in the reserve. To maintain the necessary sulphide-oxide proportion, so that the entire reserve will mine out at both a grade and a sulphide-oxide ratio that will be permissible throughout the remaining life of the property, it is obvious that the oxide reserve must be currently preserved to as great an extent as possible. Fortunately much of the oxide reserve is found in the open-pit operation. This happy circumstance adds flexibility to the procedure and makes it possible to maintain the close regulation necessary to insure best overall results, both mining and metallurgical. The two ore streams from the pit and underground operations join at the coarse-crushing plant, and proper control must be exercised before the ore reaches this point. The open pit is currently providing about 55 pct of the ore and underground operations the balance of production. Both operations are under a single superintendent so that the closely interrelated problems may be more readily controlled. The least flexible of the two producing units is the underground mine, where the mining method is block caving. Although a block of ground may be undercut and developed, characteristics of the material determine its caving action. The ore in the chutes must be drawn. The area currently mined from underground is in a zone predominately sulphide and is generally

Jan 1, 1953

By J. L. Carne

ORE control is a matter of planning and supervision based on a foreknowledge of the content and distribution of ore. The Inspiration orebody is predominately a copper-sulphide blanket, overlain by an oxidized zone of copper silicates, copper carbonates, and barren capping. Most of the ore is treated by leaching, and the optimum requirement of this method is the goal toward which the mining schedule is aimed. Both open-pit operations and underground mining are used. The problems of ore control at Inspiration are: 1-maintenance of metallurgical balance, 2-conservation of oxide ore, and 3-mining out of the orebody at an average grade which parallels that of the reserve figure. Prior to the fall of 1926 all of the Inspiration division ores were treated by flotation, but the mixed ores of the Live Oak-Keystone section carried too large a proportion of copper in the form of chrysocolla for satisfactory flotation. For this type of mixed ore the leaching process now in use had been worked out, and the plant was put into operation in the fall of 1926. The process is known as the ferric sulphate leach, and the leaching solvents contain both ferric sulphate and free sulphuric acid. Sulphide copper in the form of chalcocite is leached by the ferric sulphate, and the oxidized copper minerals are dissolved by the sulphuric acid. From the beginning of operations to the present time the plant has treated some 67 thousand tons of ore having an average copper content of 1.142 pet, of which 0.605 pct has been in the form of oxidized copper minerals. The plant is currently being operated at the rate of 11,500 tons per day, 7 days per week. However, this tonnage is produced in 6 days of mine operation. The average current grade is approximately 1 pct copper, of which about 50 pct is present as oxidized or acid-soluble copper. Close control of metallurgical procedures is a vital factor in operation of the plant and is in turn dependent upon careful regulation of feed. An excess of sulphide copper requires the presence of more ferric sulphate in the leaching solvents. On the other hand, an excess of ferric sulphate in the presence of a low sulphide feed seriously reduces the efficiency of electrolytic precipitation in tank house operations. This condition must be avoided at all costs. An additional factor must be considered. It is known that the sulphide content is predominant in the ore which remains in the reserve. To maintain the necessary sulphide-oxide proportion, so that the entire reserve will mine out at both a grade and a sulphide-oxide ratio that will be permissible throughout the remaining life of the property, it is obvious that the oxide reserve must be currently preserved to as great an extent as possible. Fortunately much of the oxide reserve is found in the open-pit operation. This happy circumstance adds flexibility to the procedure and makes it possible to maintain the close regulation necessary to insure best overall results, both mining and metallurgical. The two ore streams from the pit and underground operations join at the coarse-crushing plant, and proper control must be exercised before the ore reaches this point. The open pit is currently providing about 55 pet of the ore and underground operations the balance of production. Both operations are under a single superintendent so that the closely interrelated problems may be more readily controlled. The least flexible of the two producing units is the underground mine, where the mining method is block caving. Although a block of ground may be undercut and developed, characteristics of the material determine its caving action. The ore in the chutes must be drawn. The area currently mined from underground is in a zone predominately sulphide and is generally

Jan 1, 1952

By J. L. Carne

ORE control is a matter of planning and supervision based on a foreknowledge of the content and distribution of ore. The Inspiration orebody is predominately a copper-sulphide blanket, overlain by an oxidized zone of copper silicates, copper carbonates, and barren capping. Most of the ore is treated by leaching, and the optimum requirement of this method is the goal toward which the mining schedule is aimed. Both open-pit operations and underground mining are used. The problems of ore control at Inspiration are: I—maintenance of metallurgical balance, 2—conservation of oxide ore, and 3—mining out of the orebody at an average grade which parallels that of the reserve figure. Prior to the fall of 1926 all of the Inspiration division ores were treated by flotation, but the mixed ores of the Live Oak—Keystone section carried too large a proportion of copper in the form of chrysocolla for satisfactory flotation. For this type of mixed ore the leaching process now in use had been worked out, and the plant was put into operation in the fall of 1926. The process is known as the ferric sulphate leach, and the leaching solvents contain both ferric sulphate and free sulphuric acid. Sulphide copper in the form of chalcocite is leached by the ferric sulphate, and the oxidized copper minerals are dissolved by the sulphuric acid. From the beginning of operations to the present time the plant has treated some 67 thousand tons of ore having an average copper content of 1.142 pct, of which 0.605 pct has been in the form of oxidized copper minerals. The plant is currently being operated at the rate of 11,500 tons per day, 7 days per week. However, this tonnage is produced in 6 days of mine operation. The average current grade is approximately 1 pct copper, of which about 50 pct is present as oxidized or acid-soluble copper. Close control of metallurgical procedures is a vital factor in operation of the plant and is in turn dependent upon careful regulation of feed. An excess of sulphide copper requires the presence of more ferric sulphate in the leaching solvents. On the other hand, an excess of ferric sulphate in the presence of a low sulphide feed seriously reduces the efficiency of electrolytic precipitation in tank house operations. This condition must be avoided at all costs. An additional factor must be considered. It is known that the sulphide content is predominant in the ore which remains in the reserve. To maintain the necessary sulphide-oxide proportion, so that the entire reserve will mine out at both a grade and a sulphide-oxide ratio that will be permissible throughout the remaining life of the property, it is obvious that the oxide reserve must be currently preserved to as great an extent as possible. Fortunately much of the oxide reserve is found in the open-pit operation. This happy circumstance adds flexibility to the procedure and makes it possible to maintain the close regulation necessary to insure best overall results, both mining and metallurgical. The two ore streams from the pit and underground operations join at the coarse-crushing plant, and proper control must be exercised before the ore reaches this point. The open pit is currently providing about 55 pct of the ore and underground operations the balance of production. Both operations are under a single superintendent so that the closely interrelated problems may be more readily controlled. The least flexible of the two producing units is the underground mine, where the mining method is block caving. Although a block of ground may be undercut and developed, characteristics of the material determine its caving action. The ore in the chutes must be drawn. The area currently mined from underground is in a zone predominately sulphide and is generally

Jan 1, 1953

By Clyde E. Williams

THROUGHOUT its history the American iron and steel industry has constantly striven to improve the quality and reduce the cost of its products. No one needs to be told how well it has succeeded. Its success in making steel products at such a low cost has been achieved mainly through large-scale production. While costs have gone down as the rate of production and the degree of mechanization increased, quality of product has been actually improved. Thus high-tonnage records are important only in the light of high quality of the steel. Quality is the watchword! Continual improvement in quality has come as a result of research, not only in the laboratory but in the large commercial units in the plant. In fact, so many operations of the industry are on such a huge scale that experimental work must be done under commercial operating conditions so every progressive steel plant is a research laboratory.

Jan 1, 1938

In the period considered all coal produced was used locally, and the totals in Table 43 are estimated from the few data available. [ ]

Jan 1, 1942

By R. T. DeHoff

Saltykov's meihod for the determination of the size distribution of spherical particles from measurements made on random plane sections is extended to particle shapes which are not equiaxed. The analysis permits the deterrrzination of the size distribution of aggregates of ellipsoidol particles provided all particles have tlze same shape. V ARIOUS solutions to the problem of the determination of the size distribution of spherical particles from measurements made on random plane sections have appeared in the literature since the original solution was first published by Scheill in 1934. Solutions have been offered utilizing measurements of the diameter1y2 or area3'4 of circular sections that result when a test plane is passed through an aggregate of spherical particles, and the length of intercept5y6 that results when a test line is passed through the aggregate. To the best of the knowledge of the author, no method has been proposed which permits the determination of the size distribution of particles that are not equiaxed. The purpose of this paper is to develop a method which permits the determination of the size distribution for a specific class of particle shapes which cover awide range of axial ratios; namely, ellipsoids of revolution. In the interests of simplicity, it is necessary to limit consideration to aggregates of particles which are all similar in shape; that is, the ellipses which, by rotation about one of their axes, generate the particles, all have the same axial ratio. The size distribution, which is to be determined by the analysis, is arbitrary. It is also assumed that the section, or sections, upon which the measurements are made represent the average of all possible sections through the structure. This ideal condition can be approached as nearly as may be desired by observing a sufficient area of sections taken at random positions in a randomly oriented aggregate, and in random positions and orientations in an aggregate that has one or more axes of orientation. SIZE DISTRIBUTION FOR PROLATE ELLIPSOIDS Consider an aggregate of cigar shaped, geometrically similar particles (prolate ellipsoids), dispersed in an opaque matrix. Let the size of each particle be specified by the length of its minor axis, By Fig. 1, and the shape, q, be defined by the ratio of its minor to its major axis, B/A. Let the index, j, refer to dimensions of ellipsoids in a particular size class, and the index, i, to dimensions of an ellipse resulting from the intersection of a test plane and a particle. The size of such a section is specified by the length of its minor axis, b. Let the particles be arbitrarily divided into k size classes, and let the increment of size between each class be A, where A = &,$, B,,,being the minor axis of the ellipsoids in the largest size class. Since all sections are represented on the test plane, the minor axis of the largest of the sections observed, results from the intersection of the test plane with an ellipsoid in the largest size class at its center, and is therefore equal to Bmax Thus the increment, A = The final result will consist of the k numbers which represent the number of particles in each size class per unit volume of aggregate:

Jan 1, 1962

By W. L. Falke, A. A. Cochran

A basic problem in connection with manganese is to find economical ways to utilize domestic resources. As a part of its program to conserve domestic mineral resources and to reduce dependence of foreign sources, the Bureau of Mines had conducted extensive research to broaden manganese technology and to find economical, technically-feasible processes for recovering manganese from these low-grade resources. This paper analyzes some of the major difficulties which have been encountered in previously proposed processes for treating low-grade manganese ores and discusses a process which may circumvent many of these difficulties. Vast deposits of low-grade manganese ore exist in the U.S., yet domestic production is insignificant compared to the more than 2 million tons of ore imported yearly into this country.' A basic problem in connection with manganese is to find economical ways to utilize domestic resources.2 As a part of its program to conserve domestic mineral resources and to reduce dependence on foreign sources, the Bureau of Mines has conducted extensive research to broaden manganese technology and to find economical, technically feasible processes for recovering manganese from these low-grade resources. Although a great many processes have been proposed for treating domestic manganese resources, a failing of many of them is that too many steps are involved for economical operation.' Some of the major processes that have been proposed were reviewed by Norman and Kirby.3,4 Chlorination processes have become increasingly useful as a means of recovering valuable metals from raw materials, because the relative cost of chlorine (Cl2) and hydrogen chloride (HC1) has decreased through the years, and because the corrosion resistance of materials of construction has been improved. The objective of these processes is to form metal chlorides that are more volatile or water-soluble than are the corresponding oxides or sulfides; these compounds then can be separated from the gangue or other minerals by volatilization or leaching. Several chlorination processes for the recovery of manganese have been investigated by the Bureau of Mines5-8 and others; the major chlorination processes were reviewed in 1963, and their main features were compared.4 In all these processes, the simultaneous leaching or volatilization of iron chlorides with manganese chloride (MnCl2) is a problem, and the more iron that is present, generally the greater the problem. The leaching step is not highly selective, and the product of this step is, at best, suitable only for the production of ferromanganese or for further processing to remove iron. The product of the volatilization step is, in all cases, a mixture of iron and manganese chlorides that contains too much iron even for ferromanganese production, and must undergo additional processing steps for iron removal. THERMODYNAMIC CONSIDERATIONS The literature indicates that the desired chlorination reactions will proceed at practical rates at temperatures near 1200°K (927°C). Free energy changes at 1200°K for the critical steps of the reduction and chlorination reactions9-11 follow: Most of the research on chlorination processes for manganese recovery has been done with HC1 rather than with C12, for the following three main reasons: 1) Recycling of the chlorinating agent, which is economically vital, appeared to be easier for HC1

Jan 1, 1968

By H. Case Willcox

AERIAL photographic surveying and mapping is not new or unknown to geologists. However, it has been utilized but little before, principally because it is only within the last few months that practical means of adapting this method to the requirements of petroleum companies have been discovered. No American aerial survey company has taken up aerial surveying and mapping as an engineering service to be offered to petroleum companies and their geologists for use in undeveloped and jungle-covered petroleum areas. This method of surveying and mapping, however, is used by the War Department, the U. S. Reclamation Service, and Geological Survey, and the corresponding Departments of the Canadian Government. ADVANTAGES OF AERIAL PHOTOGRAPHIC SURVEYING AND MAPPING The advantages of aerial photographic surveying and mapping to petroleum companies and their geologists are many. One of the most valuable is presented in the very first phase, or the exploration period. It is assumed that, when convinced that petroleum possibilities are present, a petroleum company has secured a concession of 8000 sq. mi. of unexplored area in Eastern Venezuela. The company may either have the entire area surveyed and mapped by the aerial photographic method, or send its chief field geologist up in an airplane so that he, from-an aerial reconnaissance, may reject certain portions that, for geological reasons, the company would not be interested in. He will then point out the regions that are to be surveyed and mapped. After the air-work is completed, the results are brought down, developed, and printed; then the resulting "contact prints" are given to the field geologist. They may be roughly assembled to give a general idea of the property in its entirety, or they are in shape to be examined in adjoining pairs under the stereo-scope. With this instrument, the geologist is enabled to make a minute examination of surface indications. The work of assembling and constructing a complete general or detailed map, according to the scales desired by the petroleum company and for the later office use, is continued by the aerial photographic company.

Jan 3, 1925

By R. E. Zimmerman

WITH coal-preparation plants becoming more complex and containing widely diversified types of equipment and processes, it is necessary that considerable control be exercised in their operation. Rising mine costs make it imperative that close check be made of plant efficiencies to assure a minimum loss of coal to the refuse bank. Further, much of the coal sold today is sold on a competitive specification basis, and this makes it necessary to provide frequent analyses and inspections. Many ' plants are required to operate on a two or three-shift basis, and as a con- sequence an efficient organization is required to eliminate down time and provide for smooth low-cost operation. ORGANIZATION Probably the most important item for successful operation of a preparation plant is a capable, properly organized staff, with the individual workers properly informed and well trained for their particular jobs, clear-cut lines of authority among the supervisors and in general a standard operating procedure for most conceivable conditions that may occur. The idea of placing the older, partially disabled and generally incompetent coal miner in the cleaning plant, as was the practice on the earlier tipples, is entirely inadequate in a modern complex cleaning plant. The new plants require intelligent, alert, and reasonably educated men. With perhaps millions of dollars invested in an installation, it is poor business to use other than a well organized, well trained, and properly supervised force. Organizations naturally must vary with the size of the plant and local conditions; whether it is a central cleaning plant at a distance from the mine organization or one of several cleaning plants operated by the company. Smaller operations can be operated by a washery foreman, under the supervision of the mine superintendent or perhaps the "outside" superintendent. Large installations should have a plant superintendent either responsible to the mine superintendent or preferably, where possible, responsible to a preparation manager or engineer. A typical organization chart for companies operating in the anthracite region of Pennsylvania is shown in Fig. 1. Breaker (preparation plant)

Jan 1, 1950

By C. W. Price, G. A. Wheeler

BERYLLIUM has always been a difficult material to prepare for metallographic examination. Severe surface deformation occurs during mechanical grinding and polishing, and a suitable bright-field etchant has not been developed. Certain electrolytic techniques produce satisfactory polishes, but these techniques have not yielded entirely consistent results. Furthermore, they have not provided satisfactory bright-field etches. The purpose of this paper is to describe an electrolytic polishing technique that can also provide an excellent bright-field etch. In addition to the bright-field etch, other advantages of this technique are excellent reproducibility, rapidity, and ease of polishing and etching. Experimental Technique. Specimen Preparation. Beryllium specimens may be either mounted or unmounted for the use of this technique, and the size of the specimens is not important. Specimens with surface areas up to 4 sq cm have been electrolytically polished and etched successfully with this technique. The most important requirement is to provide a good electrical contact to the specimens, preferably on the back of the specimens. When specimens are mounted, a hole is drilled through the back of the mount so that the specimen can be contacted with a probe. Conventional metallographic grinding techniques are used to prepare the specimens for electrolytic polishing. The specimens are ground on silicon carbide papers. Grinding through the usual series of papers down to the 600-grit paper is recommended to minimize the electropolishing time, although successful electrolytic polishes and etches have been obtained by grinding only through the coarser papers. A continuous flow of water is used during grinding to provide lubrication and to minimize contamination. Mechanical polishing is not required because high rates of material removal are achieved with the electrolytic polishing. Polishing. Most of the results reported in this paper were obtained with a "Disa-Electropol", manufactured by Struers Scientific Instruments of Copenhagen, Denmark. Satisfactory results have also been obtained with the conventional beaker technique of electrolytic polishing, however, the conditions applicable to both the Disa unit and the beaker technique will be described. Stainless-steel cathodes are used in both techniques. In the beaker technique, a flat cathode, with an area at least twice that of the specimen, is used, and the specimen is held parallel to the cathode. The distance between the specimen and the cathode affects the quality of the electropolish. This distance is set at about 7 mm in the Disa unit; the same distance produces the best results in the beaker technique. Specimen-cathode distances greater than about 20 mm produce considerable relief, while distances less than 7 mm cause high current densities that are difficult to control. An electrolyte of 2 pct HC1, 2 pct HNO3, and 2 pct HCIOl in ethylene glycol gives the most satisfactory results for both electrolytic polishing and etching. This electrolyte was developed by altering the composition of a similar electrolyte used for electrolytic lapping1 and adding HClOl to improve the etching characteristics of the electrolyte. Variations of from 0.25 pct to as much as 5 pct of each of the acids do not significantly affect the results, but the greatest amount of control and reproducibility can be obtained with the composition of 2 pct of each acid. Generally, lesser amounts of the acids produce lower polishing and etching rates, while greater amounts of acids cause high current densities that are difficult to control.

Jan 1, 1968

By H. L. Talbot, A. J. Hall, R. S. Young

A METHOD for the chemical differentiation of oxidized cobalt from sulphide cobalt in complex mine and concentrator products is of considerable interest to a number of mining companies throughout the world. At Rhokana Corporation, in Northern Rhodesia, for instance, cobalt is found chiefly in the sulphide form, as carrollite, but a cobalt oxide mineral also occurs irregularly underground, and relatively small quantities in the mill feed from time to time are difficult to detect it visually at the concentrator. The oxide cobalt is not amenable to the sulphide flotation process used at Rhokana and nearly all of it passes into the tailing. The occasional. period when high tailing losses occur with no appreciable change in the cobalt content of the mill feed has naturally been disturbing to the concentrator staff, who had no means of determining whether an influx of cobalt oxide ore or laxity on the part of the mill flotation operators were responsible for these results. A procedure for determining oxidized copper minerals, based on their solubility in dilute sulphuric acid saturated with SO2, has been used for many years in the copper industry. In the presence of cuprite the procedure fails, but otherwise it gives results quite satisfactory for control purposes. Empirical methods for determining oxidized forms of zinc and lead have been proposed, but they are generally unsatisfactory except in isolated cases of a very simple ore. No published procedure for determining oxidized forms of cobalt or similar elements has come to the authors' notice. EXPERIMENTAL PROCEDURE There are obviously two alternatives in attacking a problem of this nature. One is to employ a solvent for oxide cobalt that will not dissolve cobalt sulphide. The other is, to find a substance capable of oxidizing the cobalt sulphide to sulphate, which of course is water-soluble and can then be leached from the water-insoluble oxide cobalt. A representative sample of Rhokana cobalt oxide ore was secured, ground to minus 200-mesh, and carefully rolled to furnish a standard sample for this work. The analysis of this material gave the following percentages: Co, 0.55; Cu, 1.64; Fe, 2.98; S, 0.08; SiO2, 41.42; A1203, 11.20; CaO, 13.82; MgO, 3.80. As a source of cobalt sulphide a similar large sample of carrollite was prepared, analyzing Co, 23.36 per cent; Cu, 27.32; and S, 31.13. This sample of course was not pure carrollite, but was the best obtainable in quantities sufficient for a standard intended for considerable work. It was a pure sulphide sample, microscopic examination showing only carrollite, a little bornite, and gangue, with no visible sign of oxidized ore present. A wide variety of reagents were used in an effort to selectively dissolve oxide cobalt or to oxidize sulphide cobalt to

Jan 1, 1946

By R. M. Hardgrove

PULVERIZED coal was first used for firing cement kilns about 45 years ago, with such success that it has continued in general use. Based on cost, pulverized coal is usually the most economical fuel. In some localities natural gas or oil is cheaper, but 85.5 per cent of all cement kilns at present are fired with pulverized coal. The pulverized-coal flame is well suited to kiln conditions. Its soft, luminous nature gives it a high radiating property and the length of the flame can be readily varied to suit conditions in a kiln of any length. Ash and slag cause most of the difficulties incident to the burning of coal. In a cement kiln, however, the ash, being similar in composition to the raw mix and being deposited uniformly in the mix and clinker, occasions no difficulty as long as it remains constant in quality and quantity. Until 6 years ago almost all cement kilns using pulverized coal employed the bin system. The coal was dried and pulverized in a separate plant and delivered into a bin at the kiln. Thus the bin system requires driers, storage-type pulverizers, conveying systems, individual kiln bins, pulverized-coal feeders and primary air fans. The direct-fired system requires only a pulverizer and a blower. Fig. I shows typical plants. The first installations of direct-fired pulverizers were of the impact type but fineness of pulverization was both inadequate and variable. About 1927 a few direct-fired installations employing tube-type pulverizers were made, which were much more satisfactory. As more efficient pulverizers were developed from 1934 on, equipment for direct firing was installed at a large number of plants, until now 196 of about 900 kilns in use in the United States are direct-fired. Most of these replaced antiquated bin systems. PROCESS OF BURNING CEMENT CLINKERS The process of burning cement clinkers is fundamentally one of heat transfer. In boilers and other heat-exchange apparatus the area of the surface receiving heat from the gases is the best unit for comparing output. This seems to be equally true for kilns. The inside kiln surface is the heat-receiving surface and its area is used as the unit of kiln size in the following comparisons. Heat input is based on British thermal units in the fuel. E. G. Bailey,1 in his work on boiler-furnace heat absorption, has used the unit "B.t.u. available." This is the practical calorific power of the fuel, plus the heat in the air used for combustion. In other words, it is the heat available for useful work in the furnace or kiln. The unit of kiln output is barrels per hour. The composition of the raw mix, the percentage of water in the slurry of wet plants and hardness of burning all affect the output. These conditions, however, are hard to evaluate and probably have less effect on the output than irregularities of raw feed and clinker discharge and inexpert control of combustion. WET VS. DRY PLANTS The study presented here covers 88 plants, of which 41 are wet and 47 dry. Of

Jan 1, 1941

By R. L. Manegold

This paper outlines the reasons for and the application of suspended magnets above belt conveyors for the purpose of removing tramp iron to protect equipment. The design of a magnet strong enough to extract iron through a distance of 24 in. is described along with the proper method of application in the field. Advantages of the suspended magnet in contrast to the conventional magnetic pulley are also discussed. IN RECENT years there has been a decided trend toward bringing ores and coal out of open-pit and underground mines by long, sloping single-stage belt conveyors. Because the high investment cost of haulage conveyors dictates maximum speed and full loading for economy, conveyor belts are running faster than ever before, with a greater burden on the belts. In addition, larger blocks of ore are being mined and conveyed. Rougher use of large mining equipment generates more tramp iron; mechanical loading rejects nothing. Therefore the extraction of tramp iron and other deleterious materials for the protection of haulage conveyors at transfer points, chutes, and idlers is becoming an increasingly difficult job. In our experience, the best method of protecting long haulage conveyor belts is to extract the tramp iron, at least the large tramp iron, ahead of the main haulage conveyor or series of conveyors. It is most practical to collect the tramp iron at the earliest source possible, usually directly after the primary car unloading dump or after the primary crushers, which are ordinarily not affected by tramp iron, see Fig. 1. The coal or ore from its primary source is fed by chutes or feeders onto a buffer or shock belt conveyor, a short, ruggedly built conveyor system that can be considered somewhat expendable. Its primary function is to bring the ore up to speed for transmission to the haulage conveyor. There is no assured method of protecting this conveyor from tramp iron damage at its feed end. Likewise, there is a possibility, although remote if the magnet applied to this belt is installed correctly, that the tramp iron extraction process will damage the shock belt, which is, however, short and comparatively inexpensive. Magnetic head pulleys with their maximum effective range of about 8 or 10 in. cannot be considered for large installations, even though they are desirable because of the continuous and automatic tramp iron removal feature. Further, the most dangerous form of tramp iron with respect to potential damage to a haulage conveyor is the long rod, roof bolt, or rail segment, which is not extracted satisfactorily because of the limited tangential contact of such an iron shape with the magnet. Magnetic head pulleys are also undesirable because the dribble or fines from the head pulley discharge at the same point as the tramp iron. Screens must be used to separate the fines from the magnetics, resulting in additional expense and loss of head room. Magnetic detectors have been used with success, but they only indicate the presence of iron, and the conveyor belt must be stopped for its manual removal. The stoppage of any conveyor means that all preceding equipment must be stopped or a cascade of material on the idle conveyor will result. Occasionally combinations of detectors and magnetic head pulleys are used. The detector precedes the magnetic head pulley and detects only large pieces of tramp iron, while the magnetic pulley subsequently takes out the smaller ferrous metals. Suspended magnets overcome the disadvantages of magnetic head pulleys, detectors or a combination of both. There are several magnetic extraction problems involved in applying a suspended magnet at this

Jan 1, 1953

By R. L. Manegold

This paper outlines the reasons for and the application of suspended magnets above belt conveyors for the purpose of removing tramp iron to protect equipment. The design of a magnet strong enough to extract iron through a distance of 24 in. is described along with the proper method of application in the field. Advantages of the suspended magnet in contrast to the conventional magnetic pulley are also discussed. IN RECENT years there has been a decided trend toward bringing ores and coal out of open-pit and underground mines by long, sloping single-stage belt conveyors. Because the high investment cost of haulage conveyors dictates maximum speed and full loading for economy, conveyor belts are running faster than ever before, with a greater burden on the belts. In addition, larger blocks of ore are being mined and conveyed. Rougher use of large mining equipment generates more tramp iron; mechanical loading rejects nothing. Therefore the extraction of tramp iron and other deleterious materials for the protection of haulage conveyors at transfer points, chutes, and idlers is becoming an increasingly difficult job. In our experience, the best method of protecting long haulage conveyor belts is to extract the tramp iron, at least the large tramp iron, ahead of the main haulage conveyor or series of conveyors. It is most practical to collect the tramp iron at the earliest source possible, usually directly after the primary car unloading dump or after the primary crushers, which are ordinarily not affected by tramp iron, see Fig. 1. The coal or ore from its primary source is fed by chutes or feeders onto a buffer or shock belt conveyor, a short, ruggedly built conveyor system that can be considered somewhat expendable. Its primary function is to bring the ore up to speed for transmission to the haulage conveyor. There is no assured method of protecting this conveyor from tramp iron damage at its feed end. Likewise, there is a possibility, although remote if the magnet applied to this belt is installed correctly, that the tramp iron extraction process will damage the shock belt, which is, however, short and comparatively inexpensive. Magnetic head pulleys with their maximum effective range of about 8 or 10 in. cannot be considered for large installations, even though they are desirable because of the continuous and automatic tramp iron removal feature. Further, the most dangerous form of tramp iron with respect to potential damage to a haulage conveyor is the long rod, roof bolt, or rail segment, which is not extracted satisfactorily because of the limited tangential contact of such an iron shape with the magnet. Magnetic head pulleys are also undesirable because the dribble or fines from the head pulley discharge at the same point as the tramp iron. Screens must be used to separate the fines from the magnetics, resulting in additional expense and loss of head room. Magnetic detectors have been used with success, but they only indicate the presence of iron, and the conveyor belt must be stopped for its manual removal. The stoppage of any conveyor means that all preceding equipment must be stopped or a cascade of material on the idle conveyor will result. Occasionally combinations of detectors and magnetic head pulleys are used. The detector precedes the magnetic head pulley and detects only large pieces of tramp iron, while the magnetic pulley subsequently takes out the smaller ferrous metals. Suspended magnets overcome the disadvantages of magnetic head pulleys, detectors or a combination of both. There are several magnetic extraction problems involved in applying a suspended magnet at this

Jan 1, 1953