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By L. M. Barker, E. G. Lewis

THE literature of milling is replete with papers devoted to the subject of ball milling, all of which no doubt have contributed in one way or another to progress in that art. In this paper reference will be made to only a few of the many authors, but the value of the contributions of all who have written on that subject is hereby acknowledged. With a few exceptions the efforts of most students of the subject have been devoted to laboratory investigations of principles. Davis1 was one of the first to report on plant-scale investigations and to present to mill men certain criteria for the guidance of ball-milling practices. His conclusions with respect to the significance of sizing analyses of classifier sands in their relation to the factors of mill speed and ball sizing are particularly referred to. Those conclusions were: I. If the balls are large or the speed of the mill is high, crowding will appear at the finer sizes in the classifier sands. 2. If the balls are small or the speed is low, crowding will appear at the coarser sizes in the classifier sands. 3. The indications are that best efficiency is obtained when the screen analysis of the sands show a minimum crowding at any size. The last conclusion was qualified by the statement that it had not been proved conclusively. It is surprising, however, that the literature of later years reveals little evidence that millmen have made any serious attempt to apply those criteria. This may, perhaps, be explained by the fact that the average plant-scale ball-milling problem presents so many variables that an investigation of any one of them encounters the problem of control of the others-a problem that often is unsurmountable in a small plant. In a large plant, in which there is a duplication of grinding units in parallel operation, opportunity may be afforded for determining the value of such criteria. Nevertheless, a study of published data from grinding circuits of both large and small plants suggests that either Davis' criteria have been tried and found inapplicable or have never been given consideration. For the purpose of testing Davis' conclusions and with improvement of ballmilling efficiency as the final objective, a test program was initiated in the concentrator of the New Cornelia branch of the Phelps Dodge Corporation early in 1938. The practices of that plant have already been described.2 The fine-grinding units are arranged so that any one of eight units may be operated as a test unit and its results compared with those of the remaining units. All units receive similar feed from a common storage bin, in which the product of the fine-crushing plant is bedded by traveling tripper. The ball mills are all similar in size and operated at the same speeds, and are individually metered. The feeder conveyors of each unit are equipped with Merrick weightometers. The test program as initially outlined proposed investigation of the following:

Jan 1, 1941

By Gerhard Derge

THE extension of metallurgical control of steelmaking processes has always made it desirable to have some quick method for determining the oxygen content of molten steel. To meet the practical demands of the steel industry, this method must also be simple enough so that the sampling can be done by the regular furnace crew. This requirement was kept in mind continually in the work to be described. An oxygen value is required for proper control, whether it be of a rimming steel or of a fully killed steel. However, two factors have discouraged the development of such direct methods: I. The reaction C + FeO -+ Fe + CO is rapid enough to make the correct sampling of the bath a difficult problem. The most generally accepted means of circumventing this difficulty have been the aluminum-killed bomb test of McCutcheon and Rautio1 and the aluminum-killed spoon test of Herty.2 The theory of these tests is that all of the oxygen in the iron is converted to A1203, which can then be determined by chemical methods. Even when simplified so that the analysis is made by a measurement of turbidity, this method is too slow to be used in routine control. Furthermore, the results obtained have never been entirely satisfactory, and some of the reasons for this will be discussed later in this paper. 2. The ordinary chemical methods of analysis for oxygen in steel are not reliable. The only analysis that seems to possess sufficient dependability to be of value is vacuum fusion,3 and the apparatus and technique developed for this have been too complex and slow for routine work. INDIRECT METHODS Indirect methods of estimating the oxygen have been resorted to in mill practice. Generally these depend upon the chemical equilibria involved. The oxygen in the bath is related to other components in the slag or metal that can be analyzed more readily. These methods have many difficulties in common. Steelmaking processes involve a complicated set of interrelated reactions, all of which tend to approach equilibrium at varying rates. The degree to which any particular equilibrium is approached, is not , well established. Moreover, a proper use of theoretical equilibrium constants requires a more extensive knowledge of the activities of the reactants than we now have. This is well illustrated by the irregularities observed when trying to determine the oxygen in the metal by its ratio to the iron oxide in the slag. Equilibrium relations may be used more readily if they are evaluated by plant experience, but the extent of approach to equilibrium remains an undetermined variable and lack of an entirely satisfactory means of determining oxygen for the calibration is felt keenly. For example, the statistical analysis of the relation between carbon and oxygen4 5 illustrates that even in a single plant wide variations exist under normal conditions of operation. Nevertheless, this is one of the best methods now

Jan 1, 1943

By Harrison Souder

THE object of this paper is to bring to. the notice of engineers a safety detonating fuse by the use of which misfires in blasting may be eliminated and safety in blasting operations promoted. This new detonator is a French invention and is known as Cordeau detonant or detonating fuse and is sold under the name of Cordeau-Bickford. It consists of a lead tube 5 to 6 mm. in diameter, filled with trinitro-toluene. While applicable to all classes of mining it will appeal especially to those who have to do with deep-hole blasting in open-cut clines or quarries, or any operations where a large number of holes are to be shot at one time. Since the introduction of the deep-hole blasting in the Cornwall iron mines, much trouble has been experienced from misfires and many experiments were made to determine the best and safest method of blasting. Before the introduction of the detonating fuse it was found that to insure the best results, two or more high-power electric exploders, not less than No. 6, with specially insulated conductors, should be placed in each hole; the holes wired in parallel, and for a large number of holes the current to be supplied by a live wire to the bus or lead wires at about the middle of the bench. The voltage in general use is 110 to 220, but as our material is a magnetic iron ore and the holes are generally damp, we found it advisable to reduce this to about 30 volts to prevent short circuits. A No. 7 or No. 8 special insulated detonator was adopted. But with the best of care and attention to the minutest detail, we were not able to eliminate occasional misfires and the consequent ever present danger of accidents.

Jan 10, 1914

The discovery of nickel at the Wallace Mine in 1846, recorded in the Report of Progress of the Geological Survey of Canada, 1.848-49, was not the first find in the Sudbury mining district. As long ago as 1770 a company had been formed to work the rich deposits of, copper whose existence was then known. .But it was during the first mentioned year that, incited thereto by fresh finds and a favorable report by Mr. W. E. Logan, then Provincial Geologist, two [ ] groups of Canadians organized "The Montreal Mining Company and "The Upper Canada Mining Company." By the former company the Bruce Mines Location' was purchased and work was immediately commenced upon it. The Wallace Mine acquired by the Upper Canada Mining Company, was abandoned after a shaft had been sunk and a few test pits opened.

Jan 1, 1907

By C. M. Myrick

(Cleveland Meeting, October, 1912) THE following notes are submitted in the belief that they may interest some of the many owners of small water-power plants, so generally used in mining-work throughout the West. A small and somewhat primitive quartz-mill was taken over by a leasing company, and a 10-stamp battery was installed. Originally, power had been furnished by a Pelton wheel ; but, since the supply of water depended on the rate at which snow was melting on the mountains, a steam-plant had been added to help out during cold spells. This auxiliary plant was antiquated and expensive to operate, and, soon after starting up the remodeled mill, it broke down completely; so that it became necessary to get all the power possible out of the Pelton wheel,. with its ever-varying water-supply. The pipe-line, about a quarter of a mile long, was made up. of assorted sizes, from 7 to 11 in. in diameter; and, since it was buried deep under snow, there were no data from which to calculate the theoretical size of the nozzle. The stream being at this time at a low stage, it was important to make the most efficient use of the available water. This was done by using a nozzle large enough to pass the whole supply-taking care, however, to limit the size so as not to lower the water-level in the tank at the head of the pipe-line. As the weather became warmer, the nozzles were gradually enlarged to meet the increasing volume of water; but a point was finally reached where further enlargement only gave decreased power. Evidently the loss from friction had more than balanced the gain from the increased flow through the pipe. However, the area of nozzle that developed the, maximum power from the pipe-line had been found. This was used; and the surplus water was allowed to run to waste at the head of the line.

Oct 1, 1912

By W. K. Hartley

INTRODUCTION AND HISTORY The Kimberley Division of De Beers Consolidated Mines Limited includes four underground mines in Kimberley viz. Dutoitspan, Bultfontein, De Beers and Wesselton Mine and two outside mines, viz. Finsch, an opencast operation and Koffie- fontein, an underground operation. Figure 1 is a map of the Kimberley district showing the location of these mines. Finsch is 150km west of Kimberley and Koffiefontein is 80km south of Kimberley. The head office of the De Beers Company is in Kimberley. The diamond sorting and trading companies for diamonds produced in South Africa are also in Kimberley. Diamonds were first discovered in the Kimberley area at Hopetown in 1866 on the banks of the Orange River. This, and later finds, triggered the start of the great "Diamond Rush". In 1870 the Jagersfontein (now closed) and Dutoitspan pipes were discovered, followed by those at Bultfontein, De Beers and Kimberley Mine (now closed) in 1871, Koffiefontein in 1880 and Wesselton in 1891. Finsch Mine was a more recent discovery in 1961. Finsch, due to go underground in 1986, and Koffiefontein, are fully self-contained with their own treatment plants, housing, services etc. The four Kimberley mines deliver ore by conveyor to a central treatment plant and use joint ser- vices. ,+lining is presently carried out mainly by Block Caving which has replaced the earlier Chambering method. De Beers Mine, on a limited scale, and Wesselton Mine have recently started to use Sub-Level Caving, and both Bultfontein and Du- toitspan are scheduled to use the same method at a later date. The combined ore tonnage produced by the Kimberley Mines is 17 400 tons per day, Koffiefontein 12 000 tons per day and Finsch Mine 17 000 tons per day. THE EVOLUTION OF MINING METHODS Initially, the open-cast mining was carried out to as great a depth as safety would permit. In the early stages of open-cast mining, the Kimberlite was removed in horizontal slices from levels 13m to 20m apart. The rock broken in the workings was raised to the surface by means of endless rope haulage inclines. As lower levels were attained these rope haulages were replaced by aerial hoists, inclined shaft haulages or inclined skip-ways. These methods were dependent upon the strength of the unsupported pipe walls, which became increasingly unstable with depth and a system was adopted whereby the ground mined in open excavations was tipped down passes to tunnels connecting with vertical hoisting shafts sunk in the country rock. As work in the open pit became more and more impracticable due to extensive falls of ground, this system was replaced by a combination of open-cast and underground mining, and eventually by wholly underground methods. Chambering With the advent of underground mining, a system known as "Chambering" (or "Inclined Chambering") was evolved. It was essentially a combination of shrinkage stoping and caving, whereby advantage was taken of the high pressure exerted by the loose rock that had accumulated in the open excavations due to the collapse of the pipe walls. This mining method was introduced around 1890 and was successfully employed, with little change, up to the early nineteen seventies. In this system each mine was served by a main hoisting shaft situated about 300m in the country rock. Main haulage cross cuts were driven from the shaft to the pipe at 200m (600ft) vertical intervals. A sub-vertical shaft, situated close to the pipe provided access to the working levels. Chamber levels were established from the sub-vertical shaft at 13m (40ft) intervals. See figure 2. On each of these levels successive 6,8m wide “cuts" were mined out across the pipe, from one rock contact to the other, and extended vertically to the footwall of the level above. Every 6,8m cut was made up of a chamber and a "back-pillart1, each approximately 3,401 wide. The chamber was mined by overhand shrinkage stoping. The levels were carried "en echelon", with each chamber on a level approximately three cuts ahead of the chamber on the level

Jan 1, 1981

By J. M. Dahl, R. J. Warrick, O. K. Riegger, H. Van Vlack

A liquid which is rich in oxygen (and silicon) develops at steel rolling temperatures in resulfurized and plain-carbon steels. This liquid fluxes solid manganese sulfide. The composition of the liquid and the resulting microstructures vary with the manganese/silicon/oxygen ratios in the steel. The observations and conclusions zuhich are presented are used 1) to suggest a mechanism for MnS deformation in resulfurized steels, and 2) to rationulize hot-shortness anomalies between plain-carbon and resulfurized steels. 1 HIS is a summary of the dependence of sulfide inclusion microstructures upon steel composition, and an interpretation of the behavior of sulfide inclusions at steel-rolling temperatures. Sulfide inclusions are known to have several metallurgical effects upon steel. Their effects upon hot-shortness are best appreciated. Although hot-shortness is well known, it is questionable whether its mechanism is fully understood because there are several anomalies in our interpretation. For example, why should not high sulfur contents produce extensive hot-shortness in free-machining steels ? A second metallurgical consequence of sulfide inclusions appears because sulfide inclusions improve machinability. Furthermore, globular inclu- sions are to be desired over elongated inclusions. To date, we know that lower silicon contents (< 0.010 pct) favor the globular sulfides, but we do not know why this interrelationship between silicon content and inclusion shape exists. Finally, the evidence is strong that the surface quality of a hot-rolled steel is particularly sensitive to the sulfur content. SUMMARY OF PREVIOUS WORK The role of sulfur in hot-shortness was first suggested during the past century. It is uncertain who was the first to conclude that sulfur was less deleterious in the presence of manganese, and that manganese altered the sulfide phase from a liquid FeS to a solid MnS at steel-rolling temperatures. However, Urban and chipman&apos; emphasized that liquid FeS may accumulate as an interdendritic film during solidification. More recently, Keh and Van lack&apos; showed that the intergranular liquid does not completely surround the metal grains at all rolling temperatures. Rather, a small finite dihedral angle exists when geometric equilibration is attained at

Jan 1, 1962

By A. W. Fahrenwald

The physics and chemistry of the flotation process are not well understood. Many papers dealing with the theory of flotation have been published but most have been narrow in their viewpoint. No theory advanced has done more than explain, more or less satisfactorily, various isolated phenomena. Scientists have done little with the problem, doubtless because they have not seen sufficient of the practical side of the subject to have their interest aroused. Any attempt to explain the mechanism of flotation that considers only contact angles, surface tension, adsorption, the floating of needles, flocculation, or any other single and narrow phase of the question is sure to be disappointing. Flotation is a problem of colloid chemistry and must be studied as such. In a broad general way, flotation is a process involving and operating by virtue of surface reactions. The term "surface reactions" may be open to criticisms. Many chemists give the term "reactions" a definite and restricted meaning and to such the term may connote change in atomic grouping in strict stoichiometric ratios. While Langmuir looks upon adsorption as a reaction, and there is much evidence in support of this view, we may look on adsorption as a physcial process as far as this paper is concerned. The term seems well suited to the work in general. Presumably these reactions are reactions between the same kinds of forces (electrical) and differ only in degree and in the surface manifestations produced. This premise seems logical if we accept the theory that the atoms of dl elements are made up of grains of electricity, having mass and charge, called electrons; and that atoms differ from one another only in the number and arrangement of the electrons about a positive nuclear

Jan 1, 1924

By W. R. Crane

MINING of the red iron ores of the Birmingham district has been carried on energetically during the past 50, years, and their development has created a large iron and steel manufacturing, center, the only important one in the South. The district produces approximately 10 per cent. of all the iron ore of the United States (80 per cent. of the Alabama ore mined is red ore) ; also, 43 of the 419 blast furnaces of the country are in the district tributary to Birmingham. The rapid growth of the district has been made possible through investigations that resulted in radical changes in furnace practice, and a still greater impetus will come from the study of the low-grade, high-silica ores, as a result of which they will be made amenable to treatment by concentration. Mining practice in this district has been comparatively simple because of the occurrence of the ore; but with the-rapid extension of the work-ings and the disturbed condition of the ore bed at some distance from the outcrop, more difficult conditions are encountered and the tendency is toward worse rather than better conditions. Should the high-silica ores of the lower bench of the Big Seam become available through beneficiation, mining practice will have to be modified to meet the new conditions, which will be rendered more difficult by the increased weight of cover that will exist at considerable distances from the outcrop. Support of workings will require greater attention and the efficient and economical operation of the mines will depend largely on the successful solution of the problems of working-and handling the ore.

Jan 9, 1924

By H. M. St. John

THE melting department is the heart of the foundry. From it clean metal flows to the molding floor, in the right quantity, at the right time and at a sufficiently low cost-or, if not, the foundry fails to thrive. Since it is our purpose here to discuss present-day melting practice in the foundry, it may be well at the start to emphasize the wide differences in foundries and the conditions under which they must operate in order to reach the goal stated in the first paragraph. In every foundry, large or small, jobbing or captive, the same factors must be dealt with but they are of varying consequence and must be dealt with in varying ways in different foundries. Particularly in the nonferrous industry, there is so wide a variation in the composition of alloys used and in the character of castings made that no one type of furnace, nor any one method of operation, can be expected to produce the best results at all times. The present discussion will not attempt to include all nonferrous metals. It is intended to cover only the copper-base alloys in which copper is the dominant element, most of which are commonly known as brass or bronze. The discussion is further limited to the melting of these copper-base alloys for use in sand-casting foundries. Die-casting plants, the brass-rolling mills and certain other metal industries have melting problems peculiar to those industries, which will not be considered here. The following considerations are elemental in all foundries and are listed in what is believed to be the relative order of their present-day importance: I. The melting department must meet the foundry&apos;s requirement for dean, fluid metal, of specified composition and purity, at a temperature suitable for the work the foundry has in hand. This is the prime consideration in any foundry, at any time, anywhere. 2. Working conditions surrounding the melting furnaces and attending their use must be good. 3. The cost of molten metal at the mold must be as low as possible. 4. The operating characteristics of the furnaces must be sufficiently flexible, to meet normal variations in foundry requirements. 5. The operation and maintenance of the melting equipment must not require more skill and experience than is available in the particular foundry in question. METAL QUALITY It is not a function of the melting furnace to refine or improve the quality of the metal but simply to melt it without impairing its quality. Good practice requires, therefore, that care must be taken to supply the melting department with metal of inherently good quality, assured by specification or otherwise, having all the properties necessary to produce the castings desired. An alloy contains a number of primary elements, which must be present in definite proportions if the castings made from it are to have the desired properties. If the castings are made to a customer&apos;s specification, this will state the range within which these principal elements (for example, copper, tin, lead and zinc) are permitted to vary.

Jan 1, 1946

By AIME AIME

ELECTRICAL methods of geophysical exploration attracted major attention at the three sessions devoted to geophysics on Monday and Tuesday. At the opening&apos; meeting Monday morning Hans Lundberg gave the paper by himself and Folke Kihlstedt; he described the use of the Racom, discussed last year, and the method whereby it is possible to eliminate all but subsurface, effects in electrical exploration. In the discussion Theodor Zuschlag pointed out some of the theoretical considerations behind the field procedure. Karl Sundberg&apos;s paper on the effect of impregnating waters was presented by title only, and S. F. Kelly then gave an abstract of a paper on the electrical properties of ores, by J. J. Jakosky, R. N. Anderson and Clyde H. Wilson. , This dealt with an exhaustive series of laboratory tests on the variation of -conductivity of ore specimens with the frequency of the energizing current, and of the effect of moisture content in typical clays and gouges on their conductivities at various frequencies. Considerable discussion was provoked during which the authors came in for some-criticism on their measuring circuits and for their assumption that a solid conductor could be compared with linear ones. They were praised, however, for making a start toward a compilation of resistivities for ores and rocks. Dr. Slichter disagreed with them that the conductivity of an orebody was primarily capacitive. He also called attention to the fact that it was high time geophysicists adopted the standard physical expression of ohm-,centimeters or ohm-meters to express resistivities. This same point was the burden of the next paper,. "Uniform Expression for Resistivity," by Sherwin F. Kelly, presented by title only.

Jan 1, 1932

By R. W. Fountain, J. Chipman

THE solubility of nitrogen in iron and steel has been of considerable interest due to the role of nitrogen in strain-aging and in quench-aging. It is generally considered that quench-aging is the result of precipitation of carbide and nitride particles from supersaturated solid solution, while strain-aging results from the segregation of carbon and nitrogen atoms to dislocations.1-9 The rate of strain-aging is controlled primarily by the amount of carbon and nitrogen in solution and the temperature. In experiments on strain-aging of pure iron with only carbon or nitrogen as impurities, Fast" has shown that even after 40 hr at 50°C carbon did not give any strain-aging while

Jan 1, 1959

By Anthony William Deller

MAJOR battles in the present war have been fought in American research laboratories. Without the outstanding contributions made by our scientists, engineers, and technologists in mining and metallurgy, many of our achievements in the present war would have been impossible. All recognize the fact that modern research has found the answer to problems which seemed insuperable. For the duration, many of these inventions must be maintained as war secrets. But aside from these, the record of accomplishment in peace and war, especially in the past few decades, known to all, demonstrates what scientific and industrial research can accomplish. It may be regarded as axiomatic that the development of modern industry is based principally upon research devoted to making inventions and discoveries. We have come a long way from the sixteenth century, when Agricola published his classical "De Re Metallica:" With our vastly increased knowledge we are now commercially producing dozens of metals entirely unknown at that time. Further, with our modern methods of research and development, a large number of invaluable alloy compositions have been produced. These include numerous age-hardenable alloys, cupronickels, "Everbrite", "Everdur", "Duralumin", "Elinvar", "Hastelloy", "Inconel", "Invar", "Monel", "Ni-chrome", "Ni-Hard", "Ni-Resist", "Permaloy", "Silcrome", stainless steels and irons and other alloy steels, "Stellite", and zinc die-casting alloys.

Jan 1, 1945

By Helmut Thielsch

METALLOGRAPHIC specimens whose surfaces are to be investigated are too thick to allow either light or electrons to pass through them for microexamination by transmission. This difficulty is overcome with optical microscopes by illuminating the surface through the objective lens. For electron microscopes best results are obtained by preparing thin replicas of the surface of the specimen, placing them in the microscope and passing the electron beam through them. Since electron microscopes represent a rather recent invention, they have found little application as compared with optical microscopes. This is true because, aside from the cost of new equipment and development of new techniques, some of the older replica methods necessitated the destruction of the surface from which a thin transparent replica was obtained. In other processes in which the surface is preserved, either too tedious a procedure is required, unsuited to "mass-production " requirements of industrial laboratories, or replicas of insufficient contrast and sharpness are produced. In general, investigations with the electron microscope involve five steps: (I) polishing the surface of the metal specimen, (2) proper etching of the surface, (3) preparing the suitable replica from the surface, (4) examination and photography in electron microscope and (5) interpretation. PREPARATION OF METAL SURFACES Usual metallographic polishing is generally satisfactory. It is important though that the final polish be applied very care- fully, since otherwise evidence of a small amount of deformation might show up even on relatively deeply etched surfaces. A high-power microscope should be used to determine the suitability of the polish and etch. This, of course, is not different from optical examinations, but in these, if a deep etch is used, the problem is less serious. It is frequently difficult to detect with optical microscopes poor polishing aside from superficial scratches before etching. Figs. I and 2 show micrographs of "well" and "poorly" polished nickel etched by immersing the same for 20 min. in a solution containing 8 grams cupricsulphate in 40 C.C. of concentrated HCl and 40 C.C. of H20 (Marble&apos;s reagent). The sample shown in Fig. I was polished in its final stages for 5 min. on a carborundum wheel followed by 10 min. on a rouge wheel, whereas the sample shown in Fig. 2 was polished only I min. on the rouge wheel. Generally, this procedure does not guarantee that the surface of the specimen will be either undistorted or deformed. It was applied here merely to show the effects of good and poor polishing. To the experienced investigator, plastic deformation is

Jan 1, 1946

By J. B. Cohen, B. W. Howlett, M. B. Bever

The partial molar heats of solution at infinite dilution in tin of aluminum at 300° and 350°C and of gallium, indium, and thallium at 240°, 300°, and 350°C have been measured by tin solution calori-metry. Aluminum, gallium, and thallium are endo-thermic on solution; indium is exothermic. Any temperature dependence of the heats of solution lies within the experimental scatter. Over the dilute ranges investigated, only aluminum has a measurable change in its heat of solution with composition. HEATS of solution of one element in another reflect the interaction between them. The investigation of partial molar heats of solution in dilute alloys is of particular interest as the properties of the solvent are altered to only a limited extent by the presence of the solute and also as the interaction between solute atoms is small. When the heats of solution of a related series of elements in a solvent are known, a systematic comparison may be made. In the investigation reported here, the partial molar heats of solution of the Group III elements aluminum, gallium, indium, and thallium in dilute solution in tin were measured. This work follows an investigation of the heats of solution in tin of the Group IB elements.&apos; EXPERIMENTAL PROCEDURES Materials. Samples of gallium, indium, and thallium were obtained from Johnson, Mathey and Co., Ltd. Indium and thallium were supplied as wire, 1.6 mm in diam; gallium was in the form of irregular pieces. The supplier reported the following minimum purities: gallium—99.95 pct; indium—99.99 pct; thallium—99.99 pct. The aluminum, obtained from Alcoa Research Laboratories, was reported to be 99.995 pct pure. The tin was supplied by Baker and Co., Inc.; the reported analysis indicated a tin content of at least 99.96 pct with lead as principal impurity. Calorimeter. this description will cover only the essential features of the calorimeter with special attention to modifications made since an earlier description was published.&apos; A Dewar flask containing the tin bath was held in a constant-temperature bath of a near-eutectic mixture of lithium, sodium, and potassium nitrates. This bath, which was stirred vigorously, was heated by a primary resistance winding in the container wall and by a secondary winding immersed in the salt. The voltage supplied to both windings was stabilized. The temperature of the salt bath was controlled by means of a platinum resistance thermometer in one arm of a Wheatstone bridge. The light from a mirror galvanometer in the bridge circuit fell on a photocell which controlled the current in a saturable reactor in series with the secondary winding. In this manner, the temperature of the salt bath was controlled to ±0.003°C and that of the tin bath to at least ±0.002°C. Each of these temperatures was measured by two iron-constantan thermocouples in series, coiled in a helix to minimize heat loss and immersed in the salt and tin baths in protective sheaths. The temperature of the laboratory was kept constant to ± 1°C during a run. Specimens were dropped into the tin bath from an addition arm held at O°C which was part of the cal-orimetric system. The system was evacuated to about 0.02 1 to minimize oxidation and to reduce transfer of heat. The bath was stirred by a glass stirrer introduced through a double Wilson seal. The samples were scraped clean before weighing, which was carried out as rapidly as possible. Each sample was immediately placed in the evacuated addition arm to minimize contamination. These precautions were especially necessary with aluminum. After the runs with gallium and thallium at all temperatures and with indium at 240° and 350°C (Series I) were completed and before the runs with indium at 300°C and aluminum at 300° and 350°C (Series 11) were begun the following changes were made. The shape of the Dewar flask was changed so as to result in a lower surface to volume ratio of the bath and at the same time the amount of tin was reduced from 500 to 400 g. The paddle type stirrer was replaced by a helical screw and the rate of stirring was increased to about 150 to 200 rpm.

Jan 1, 1962

By C. B. Manula, H. Gezik

In planning modem-day mining operations, management needs to pass from the area of subjective decision-making to an area of objective decision-makirlg. Planning procedures currently being practiced by many mine managers depend primarily on common sense and trial-and-error methods which oftentimes never reach optimum solutions. For this reason, an attempt is made to provide an understanding of linear programming to solve production planning problems. More specifically, a linear programming model is proposed to solve the production scheduling problem as it was found to exist in the crushed stone industry. Optimal production scheduling is of prime concern here, since the industry has to cope with seasonal demands for its products. The solution, as determined through the application of the model, provides management not only with an optimal production schedule but also with the following information regarding the improvement of existing facilities: (I) the location of bottlenecks which limit production and profit; (2) the directions in which the company i fixed equipment should be expanded or modified; and (3) the evaluation of unused resources. Future trends in stone production will quite likely parallel those in the construction industry. Continued record expenditures for structures such as dams, highways, and buildings are the reasons for this accelerated growth. Production forecasts for the next decade are predicted to be around one billion tons with an approximate 7% annual increase. The problem of increasing production from 570 million tons in 1966 to 1 billion tons by 1976 offers no small challenge. This situation is further compounded by other problems and economic factors. Population expansion, government regulation, labor agreement, and changes in technology wd cause managers and operators to be confronted by many more complexities than characterize present-day operations. Some of these. problems and economic factors are: (1) In some metropolitan and other urban areas, easily accessible sources of stone are almost depleted. Transportation costs will be a prime concern of the industry. (2) Population growth has forced residential expansion to a point where many quarry operators must conform to urban and suburban regulations. (3) Government regulation in regard to air and water pollution, noise, and reclamation as well as zoning developments necessitate that crushed stone producers keep abreast of changing conditions and requirements. (4) The crushed stone industry operates under pressure of a highly competitive situation. Every company&apos;s objective is to retain a favorable marketing position by exercising great attention to production and marketing costs. This is rather difficult, since the crushed stone industry is one segment of modern mining which has to cope with a seasonal demand for its products. Among the problems posed here, the one of exercising great attention to production and marketing costs is of current interest to management. Under seasonal demand conditions, this problem creates further difficulties for managers who need to plan and forecast. STATEMENT OF THE PROBLEM Various methods of Operations Research have been developed to handle a problem whose genesis is explained above. It is very seldom, however, that these quantitative means are used in practice by the mining industry. This paper describes the use of Operations Research methods to solve a production planning problem when demand is seasonal. The seasonal production planning problem is an important one because of the substantial costs associated with changing production, employment, and inventory levels to adjust to seasonal sales patterns. Most companies try to operate within reasonable levels of employment changes and hold inventories within storage capacities. In general, little or no attempts were made to develop minimum cosl schedules or to provide objective means to compare various production policies. When demand varies within a fixed time period, 11:-oduction requirements may be met by following one

Jan 1, 1970

By Kenelm C. Appleyard

In introducing to an American audience a description of the work done in dry coal separation in England and in Europe generally, it is perhaps desirable to give a short history of the development outside the United States of the particular processes with which the author is connected, and perhaps to stress a little the differences in the nature of the conditions to be faced. England actually presents one set of problems and the Continent of Europe another, since the character of the coals to be treated is entirely dissimilar, and the attitude of the coal industries towards coal preparation is also dissimilar. England for generations has been working by hand good seams of clean coal of the highest possible quality and has not felt any pressing need to clean its coal. Most Continental countries, however, especially in Central and Western Europe, have been working friable and often very dirty coals for so long that a washery is accepted as a part of the normal equipment of a colliery, with the result that the art of wet washing has reached a high state of perfection, particularly in Belgium and Germany. The advent of machine mining has intensified the problems to be faced in all countries, and in England has been coupled with the working out of the best seams in the older coal fields, thus bringing about the necessity of working thinner and dirtier seams with resultant difficulties as regards marketing the products, while Poland, with clean and well prepared coal, has in recent years been attacking established British markets in Scandinavia. In recent years, therefore, interest in coal preparation has been steadily gaining ground in England, although it has not yet reached Continental dimensions. Broadly speaking, the differences in preparation problems between England and the Continent boil themselves down to one of size. In England, for example, with the exception of the Kent coal field, the proportion of coal of a size less than ? in. to the coal less than 2 in. is very much less than on the Continent, and since the principal difficulties in dry cleaning lie in the fines it is easier to produce good figures on

Jan 1, 1931

By J. C. McDonald

A DDITIONAL experimental data are presented in this note on a phenomenon which has been touched on but lightly in the literature. The common magnesium base rolling alloy (AZ31B) contains about 3 pet A1 and 1 pet Zn. When cold rolled, it hardens until it starts to break up in the rolls at 30 to 50 pet reduction. When it is heated after rolling to full hardness, it softens. However, other magnesium alloys have shown themselves capable of being cold rolled indefinitely, provided the ratio of roll diameter to starting gage is sufficiently high. As measured by the tensile test, these alloys actually soften when the reduction becomes large (60 to 90 pct). A subsequent heating at 300&apos; to 500°F can produce hardening of this metal. Heating at higher temperatures (600&apos; to 800°F) causes recrystalliza-tion and softening. Menzen&apos; has shown that a magnesium base alloy can be cold rolled without intermediate anneals as much as 90 pet. The alloy studied, Mg-1.5 pet Mn, softened considerably between 20 and 90 pet reduction, as measured by tensile properties; a decrease in yield strength of 30 pet, and an increase in elongation from 3.2 to 16.7 pet (longitudinal direction of the sheet). However, Menzen gives no data on the reaction of this metal. to heating. Ansel and Betterton2 and Hurst and Hatch3 report a similar effect in Mg-1.5 pet Mn-0.1 pet Ca. Their data show an initial hardening, followed by softening. The former authors also show that the alloy softened by cold rolling will harden when heated at low temperatures, and that these effects are absent in Mg-3 pet Al-1 pet Zn. In the course of work4, 5 some years ago, the author observed similar effects with alloys containing cerium, thorium, or zirconium; but not with elements like aluminum, silver, or zinc. Table IA contains the results of an experiment which shows these effects rather well. A binary alloy with 0.3 pet total rare earths (Mischmetal* added) was rolled hot to 1/16 in. and then annealed to full softness by heating 30 min at 800°F. Rolling was continued at 1.5 pet per pass to 80 pet total reduction on 8 in. diam rolls. The metal was harder after 20 pet work than initially but, at 80 pet reduction, properties were back to their original values. Heating for 1 hr at 450°F produced about the same hardness whether the reduction was 20 or 80 pet. At intermediate reductions, the metal hardened, but not to as high values. Extensive cold rolling may also be carried out on metal hot rolled at temperatures sufficiently low to produce a semihard temper. Mg-0.6 pet Zr was rolled to 0.1 in. at 700° to 800°F, 30 pet per pass. As shown in Table IB, a further cold reduction of 85 pet, at 1.5 pet per pass, definitely softened the metal.

Jan 1, 1959

By N. E. Maccallum

(Cleveland Meeting, October, 1912.) ALMOST half a century has passed since the Siemens brothers, after tedious and costly experiments, finally began the manufacture of open-hearth steel. The furnace of that time was very small, having a hearth-capacity of from 3 to 4 tons, and we can appreciate the wonderful development that has taken place in those 50 years by the knowledge that to-day furnaces of 50 times the original capacity are in operation. As the new steel process assumed importance, and furnaces began to grow in number and in size, a question arose, to which no satisfactory answer has ever been given, and that was: ?What is the safe and economical limit of the open-hearth furnace?" Various stopping-places have been suggested, and 20, 30, 40, and 50 tons have each been advocated, but all such arbitrary restrictions have been passed, and 60 tons would probably represent the average furnace, although in a few instances from 70 to 80 tons are regularly tapped into one ladle. The furnaces which are the subject of the present paper have gone far beyond even these limits, and since, in my opinion, they mark another step in the progress of open-hearth steel, they will be considered fairly and frankly, in order that those to whom the matter may be of interest may be able to judge and form their own Conclusions. Short accounts of the work of these furnaces have appeared from time to time in various technical journals, of which the following is a brief résumé. These furnaces are the property of the Phoenix Iron Co., of Phoenixville, Pa., and their distinctive and important feature is, that they charge to twice the capacity of one ladle, and tap simultaneously into two ladles, with a device for controlling the flow into either. In this way the capacity of the furnace is not limited, as formerly, but heats of any size within the capacity of two ladles may be charged.

Oct 1, 1912

By P. A. Flinn

ALTHOUGH many physical properties of superlat-tices have been studied intensively, relatively little attention has been paid to their mechanical properties until recently. Even for the well-known transformation in 0 brass, the effect of the transformation on plastic behavior has only lately been extensively investigated.&apos;&apos;&apos; A partial analysis of dislocation structure in superlattices has been given by cottrel13 and his theory agrees well with the experimental results on CU3AU.4,5 His analysis, however, is implicitly limited to low-temperature effects, since it does not include the behavior of dislocations at temperatures high enough for diffusion, and hence climb, to occur. The work of Herman and Brown2 on the creep of 0 brass indicates that important effects occur under such conditions, and consequently, the theory should be extended to include them. Such high-temperature effects cannot be observed in Cu3Au, since it disorders below the temperature at which diffusion occurs at a reasonable rate. There are materials of the same structure, however, such as Ni3A1, which retain their order up to, or close to, the melting point. GEOMETRY OF THE COMMON SUPERLATTICES The two best-known ordered structures are the 0 brass type, B2, shown in Fig. 1, and the Cu3Au type, Ll2, shown in Fig. 2. In the B2 structure the (000) and (1/2, 1/2, 1/2) sites are different in that they are occupied by different types of atoms. The symmetry of the structure is then lowered from a body-centered cubic, A2, in which the sites are equivalent, with atoms located randomly in the sites, to simple cubic. This lowering of symmetry has the consequence that two types of regions may occur within a crystal: regions where A atoms are in the (0, 0, 0) sites, and regions where they are in (1/2, 1/2, 1/2) sites. The boundary between two such regions will contain bonds between like atoms, and be a surface of higher energy. These regions are known as antiphase domains and the boundary as antiphase boundary. In the B2 structure, boundary formed during the ordering process should disappear as a result of domain growth, since any domain growing to the boundary of two others will join one and engulf the other.&apos; In any well annealed crystal, only a single domain should exist, with antiphase boundary present only in connection with dislocations, as discussed below. The L1, structure is related to the A1 (face-centered cubic) in much the same way as the B2 is to the A2. The face-center sites are occupied by A atoms, and the corner sites by B atoms. The symmetry is again lowered to simple cubic, and antiphase domains can exist. In this case, however, there are four initially equivalent sites in the unit cell: (0,0,0), (0, 1/2, 1/2), (1/2,0, 1/2) and (1/2, 1/2, 0). In the ordered structure any one may be occupied by a B atom and the other three by A atoms. Thus four types of antiphase domains can occur in this structure. This is sufficient to permit a metastable "foam structure" to exist within a crystal.&apos;&apos;7 Antiphase boundaries along three different planes are shown in Figs. 3, 4, and 5.* *The presence of a layer of disordered material between antiphase domains, as suggested by Jones and Sykes8 does not seem likely. Such a layer would be a region of unnecessarily high energy which could readily be reduced by the formation of the type of boundary discussed here.__________________________________________________________ The notation (111) [110] means a boundary lying along a (111) plane between antiphase regions related by a 1/2 [I10] translation. It may be seen in Figs. 3 and 4 that an antiphase boundary usually results in incorrect nearest neighbors (shown connected by braces). However, as pointed out by Wilson, a boundary of the (001) [I101 type can exist

Jan 1, 1961