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By Dever C. Ashmead

COAL companies in the anthracite region are studying various methods of mining that will permit a considerably shorter life of gangway and therefore a decrease in the maintenance charges. Maintenance charges in the upkeep of gangways where the measures are steeply pitching are high due to the excessive cost of the replacement of timbers and the cleaning up of falls. Therefore, if the length of gangway in the coal beds can be decreased and the mining more concentrated, the life of the coal gangways will be decreased and consequently the maintenance charges will be lower per ton of coal produced. Customary practice in the anthracite region is to drive long coal gangways and to work the breasts as the gangway from which they were driven is advanced.

Jan 1, 1925

By J. F. Myers, R. J. Tower

"THERE is nothing new under the sun." All over the world, mineral-dressing engineers are working at their problems, no two of which are alike. Each encounters equipment and process problems. Many devise simple ways and means of overcoming their difficulties. Many of these problems are quite common in the industry, yet, since no one human being seems capable of thinking of everything, these problems at some plants often are looked upon as necessary evils. The purpose of this symposium is to gather together, from competent engineers, devices and practices that have been useful to them. Most of the ideas contributed are not new, yet there is no plant operator, old timer or youngster, who will not find some idea herein that will give him a suggestion for improving his operation. Space does not permit a complete elaboration on all of the contributions offered. Each item is summarized, but this will be ample for those skilled in the art of mineral dressing. A list of contributors appears at the end of the paper, and the superior number at the end of each item identifies its source. GENERAL EQUIPMENT A signal light to show that a steel bin, containing relatively fine, moist ore, is full, can be made by suspending a copper rod in the bin to a sufficient depth to ensure ample contact with the ore. One wire of a 32-volt circuit should be attached to the rod and one to the side of the bin. Sufficient current will flow to make the signal light burn. A 110-volt circuit could be used, but 32 volts ensures safety.10 Fig I shows a circular bin for crushed ore. It is a concrete structure footed on the ground, using no supporting columns. Tailing from the sink-float was used for aggregate. Cross conveyors for discharging ore are not required, nor is a tripper for filling. Segregation is minimized by drawing from different gates. An exhaust fan on top of the bin eliminates condensation due to temperature of the ore.21 A beam with an adjustable point of suspension near the center permits rigging rolls for removal in a very few minutes. This makes unnecessary the lacing of the cable through the hub bolts on one side, around the shaft on the other, and the trouble of balancing the load.8 The ratchet handle on a Symons crusher may be replaced by a drum with a cable wound on it. When the cable is pulled by a crane or hoist it causes the drum to revolve and thus screws or unscrews the bowl. This reduces handwork and materially shortens the time required for the job.3 To spread ore feed from an apron feeder over a belt conveyor, place a heavy steel bar about midway between the top

Jan 1, 1947

By P. Duwez, H. Martens

Dimensional changes generally occur during the sintering of metal powder compacts. These changes may have several causes and their magnitude depends upon the numerous variables involved in the process, such as the nature of the powder, size and shape of the grains, compacting pressure, rate of heating, maximum temperature, length of time at temperature, and furnace atmosphere. Among all the variables, time, temperature, and compacting pressure are probably the most important ones, and the effect of each one of the other variables can be studied separately once the effect of these is well understood. The most precise method of studying the phenomenon of dimensional changes in a compact consists of heat treating the compacts at various temperatures for different lengths of time. In this method, only very small specimens can be used, in order to achieve rapid rates of heating and cooling. The technique involved in these measurements is very simple, but is time consuming. Another experimental approach to the problem consists of measuring the change in length of a compact as it is being heated at a constant rate. Although the two variables, time and temperature, are combined into a single one (rate of heating), the results obtained by this technique may be of great interest from both practical and theoretical points of view. The purpose of this paper is to describe a systematic investigation of the change in length during sintering of compacts made of one or several metal powders. Experimental Procedure The change in lenght of compacts during sintering was measured with an automatic recording dilatometer. In this instrument the change in length and the temperature are recorded as functions of time on two separate charts and the rates of heating and cooling may be varied from 1 to 20°F per min. The temperature may be kept constant at any value up to 2200°F during any predetermined length of time and the change in length recorded as a function of time. Provision is made for atmosphere control inside the fused silica tube containing the specimen. Three different magnifications are provided, so that 1 in. on the chart corresponds to 0.002, 0.005, or 0.010 in. change in length. The mechanism following the change

Jan 1, 1950

By F. S. McNicholas

IN the first big blast at Climax, a "loop back" (three-wire system) was used (Fig. 1), with the idea of securing a wiring system that would give to all series the same amount of current. Single-phase, 60-cycle, 13,800-volt FIG. 1.-Loop BACK. A THREE-WIRE SYSTEM WITH BRANCHES VARYING IN LENGTH. power was furnished to the blasting transformers, each of which has a rated capacity of 600 amp. at 480 volts. Two blasting leads were taken from these transformers, giving a blasting current of single phase, 60 cycle, 480 volts, with an instantaneous capacity of more than 600 amp. The blasting switch was on the secondary leads from the transformers, to avoid lag in energizing them.

Jan 1, 1938

By B. M. Larsen, T. E. Brower

In an earlier paper1 we discussed a simple, rapid method of taking samples of liquid steel and analyzing them for oxygen, which, though possibly not absolutely accurate (as is likewise true of all other methods), certainly yields results that are satisfactorily consistent. Accordingly, samples from hundreds of heats ill different shops were analyzed for oxygen, with results sufficient in number to be handled statistically and permit estimation of some general trends. Measurements were made in acid as well as in basic open hearths, throughout the refining period, to learn how stirring, blocking, adding ore or lime or sand, affects the oxygen content of the bath; we also followed the changes brought about by tapping, teeming, and deoxidizing. Some of these results, discussed in an earlier paper,2 bring out the dominant role of the carbon-oxygen reaction in regulating degree of oxidation both of metal and of slag, also the marked variability of concentration of dissolved oxygen [0] in excess of that which would be in equilibrium with carbon and with carbon monoxide at a partial pressure of one atmosphere. As a logical beginning on this complex problem, we take up in the present paper the variability of this excess oxygen, A[O], when other factors are as constant as they ever are in open-hearth operation; namely, toward the end of the refining period, when the bath may be regarded as in the comparatively steady state it approaches in absence of disturbing factors. This condition was chosen because by that time the significant reactions, except carbon oxidation, between slag and metal have come to substantial equilibrium; this facilitates interpretation of the complex situation and thus perinits more definite conclusions to be drawn. This excess concentration of oxygen, A[O], in the metal is needed, of course, to make the steelmaking process go at a reasonable rate; alternatively it may be regarded as a measure of the virtual pressure increment above one atmosphere, or of the degree of supersaturation, of carbon monoxide in the metal. According to our measurements, the excess concentration A[0] in liquid steel varies from about 0.009 to 0,035 per cent oxygen by weight. It is quite sensitive to change of conditions, being altered temporarily by operations such as stirring with green poles, additions of ore, lime, or sand; but as the disturbance works itself out, it resumes a level within the intermediate range characteristic of the steady state. The conditions most common in average furnace practice are covered by the following: i[O] Usual Per Cent Oxygen by Conditions Weight 1. Vigorous lime boil........... o.oog-0.01s 2. Steady state................ o.or;-0.025 3. Working of ore.............. o.023-0.035 Thus when the bath can be regarded as in a reasonably steady state—that is, when it is boiling normally, is at proper temperature with a well-shaped slag, and has been free from disturbing effects during the preceding half hour—A[0] returns from

Jan 1, 1948

By B. M. Larsen, T. E. Brower

In an earlier paper1 we discussed a simple, rapid method of taking samples of liquid steel and analyzing them for oxygen, which, though possibly not absolutely accurate (as is likewise true of all other methods), certainly yields results that are satisfactorily consistent. Accordingly, samples from hundreds of heats ill different shops were analyzed for oxygen, with results sufficient in number to be handled statistically and permit estimation of some general trends. Measurements were made in acid as well as in basic open hearths, throughout the refining period, to learn how stirring, blocking, adding ore or lime or sand, affects the oxygen content of the bath; we also followed the changes brought about by tapping, teeming, and deoxidizing. Some of these results, discussed in an earlier paper,2 bring out the dominant role of the carbon-oxygen reaction in regulating degree of oxidation both of metal and of slag, also the marked variability of concentration of dissolved oxygen [0] in excess of that which would be in equilibrium with carbon and with carbon monoxide at a partial pressure of one atmosphere. As a logical beginning on this complex problem, we take up in the present paper the variability of this excess oxygen, A[O], when other factors are as constant as they ever are in open-hearth operation; namely, toward the end of the refining period, when the bath may be regarded as in the comparatively steady state it approaches in absence of disturbing factors. This condition was chosen because by that time the significant reactions, except carbon oxidation, between slag and metal have come to substantial equilibrium; this facilitates interpretation of the complex situation and thus perinits more definite conclusions to be drawn. This excess concentration of oxygen, A[O], in the metal is needed, of course, to make the steelmaking process go at a reasonable rate; alternatively it may be regarded as a measure of the virtual pressure increment above one atmosphere, or of the degree of supersaturation, of carbon monoxide in the metal. According to our measurements, the excess concentration A[0] in liquid steel varies from about 0.009 to 0,035 per cent oxygen by weight. It is quite sensitive to change of conditions, being altered temporarily by operations such as stirring with green poles, additions of ore, lime, or sand; but as the disturbance works itself out, it resumes a level within the intermediate range characteristic of the steady state. The conditions most common in average furnace practice are covered by the following: i[O] Usual Per Cent Oxygen by Conditions Weight 1. Vigorous lime boil........... o.oog-0.01s 2. Steady state................ o.or;-0.025 3. Working of ore.............. o.023-0.035 Thus when the bath can be regarded as in a reasonably steady state—that is, when it is boiling normally, is at proper temperature with a well-shaped slag, and has been free from disturbing effects during the preceding half hour—A[0] returns from

Jan 1, 1948

By E. A. Slover

THE usual reverberatory system of smelting cop--1- per concentrate or calcine has for its component parts a furnace and one or two waste heat boilers. These parts are operated on a basis of compromise, since the furnace can send gas to the boilers at too high a temperature and the boilers by plugging, due to dust or slag, can place a definite limit on the amount of fuel the furnace can burn. Over the years the copper concentrate smelting furnace has had few advances in design. The simple rules of design such as the flame should wipe the bath and the speed of the gases should be reasonably low for dust carrying purposes seem to cover the main features. In the construction of the individual furnaces some innovations are always being introduced. Among these are charging so that the work of smelting is a complete bath process, the use of suspended brick arches in place of sprung arches, the use of basic brick, not only in the crucible, but also in roof and sidewalls, the use of various means to feed the charge, the use of magnetite or other heavy material to construct the hearth, water cooling of bridgewall and slag skimming bay, the smelting of raw charge instead of calcine, the use of preheated air, and possibly the use of oxygen-enriched air for combustion. But the general outlines of the furnaces have not changed much except as to size. Furnaces at Hurley As shown on Fig. 1, the furnace at Hurley is 126 ft long between the longitudinal buckstays and 32 ft wide at the skewback plates. The foundation is a concrete retaining wall with piers at intervals that go deeper into the earth. Purposely the wall at the burner end of the furnace is not backed-up as tightly as the other parts of the foundation so that movement due to expansion may take place here rather than into the boiler foundations. Within these foundation retaining walls of concrete, the earth has been removed to allow the placement of the crucible brick base inside of which a silica hearth is laid 4 ft 6 in. in depth. No expansion is left in the brick base and crucible where they are in contact with the hearth. The hearth itself is of quartzite crushed to 1 in. size with fines left in the product. An 8 in. layer is laid and tamped with paving tampers to about 6 in. in thickness. Then a layer of silica flour is spread and vibrated into the hearth. This operation is repeated until a depth of 4 ft 6 in. is occupied by the silica mass onion-skinned in layers of approximately 6 in. Before firing the entire hearth is covered with broken slag to a depth of 4 in. so that a seal may be formed on the hearth. The crucible is completely faced with magnesite chemically bonded brick while the outside, against the foundation, is made of silica brick. The side-walls are carried up with silica brick in which expansion joints are left at intervals. Above the crucible the sidewall is corbelled to form a shelf on which the charge may build up along the side-walls, see Fig. 2. The arch of the furnace is sprung 20 in. silica brick, with the longitudinal centerline horizontal the length of the furnace, and some 9 ft in the center above the bath. Both straight and wedge brick are used in the construction and a thin silica mortar is troweled for joints. After the arch under heat has assumed its permanent shape, a silica slurry is spread over the arch to fill any cracks that have formed, thus giving bearing surface to the brick and preventing dust from entering the body of the arch to act as a future fluxing agent. The uptake of the furnace slopes up to the boiler entrance where a brick pilaster divides the gas stream for the two boilers. Over this flared uptake is a suspended flat arch of firebrick. The pilaster and sidewalls are constructed of firebrick but the bottom of the uptake is lined with silica brick and fettled through holes in the roof with siliceous fettling. Close to the entrance of each boiler is a brick covered slot through which water-cooled dampers may be lowered in event of boiler trouble. These water-cooled dampers are hung permanently in position ready to be lowered when needed. Flexible hoses to follow the dampers as they are lowered are connected at all times and individual chain blocks are used to lower the dampers. A pump supplying water is started before the dampers enter the heat. Charging of the furnace along the sidewalls for some 80 ft from the bridgewall is accomplished by electric vibrating conveyors fed by belt from charge storage bins above the furnace. These conveyors

Jan 1, 1954

By J. J. Scott

This investigation utilizes three dimensional photo-elastic techniques to determine the stress distributions in the pillars and the roof of a room and pillar mine model. Castolite plastic was used for the construction of the model in which a stress pattern was developed by the stress freezing process. Analysis of the stress pattern was accomplished by observation of points in slices from the model with a polariscope and solving the data thus obtained by the shear-difference method. Slices through the centerline, edge of pillar and centerline of span were investigated. The results of this work provide information on the overall stress distribution in points of stress concentration in compression, shear and tension, show relative magnitude of stress throughout the model and indicate directions of the most important stress values. A generalized roof bolt pattern is developed for room and pillar mines based on the stress distribution. A digital computer was used in the reduction of data and proved to be extremely useful for investigation of this type. Underground excavations disturb the equilibrium of the earth's crust and a core of fracture, or altered stress distributions exists around the openings. This fact is well known by mining operators and investigators in the rock mechanics field. Analysis of the stress distribution in the areas around mine workings is a most difficult problem and one which has no simple and direct solution. Some of the approaches used by investigators in attempts to study stress distribution around mine openings involve mathematical analysis, photoelastic techniques, statistical analysis, model studies on brittle materials and observation of fracture patterns. Models involving imbedded strain gauges, gelatin models using imbedded strings to show internal movement, and many other unique laboratory techniques are employed. Work is also being done in mining operations where load cells, strain gauges, photostress material, convergence plugs, dynamic plugs, etc., are used to measure strain which can in turn be used to obtain stress. Some of these devices measure strain from which stress may be derived. All approaches to the solution of stress distribution are hampered by the fact that the earth is not a homogeneous, isotropic and elastic media. Geologic faults, fractures, bedding planes, vugs, variations in material composition, variations in grain size and structure, etc., make it extremely difficult to obtain results which can be universally applied to mining operations. The purpose of this investigation will be to demonstrate the use of three-dimensional photoelastic techniques to obtain the stress distribution in the pillars and roof of a room and pillar mining operation. This method utilizes full analysis of interior points, and thus it may be that small tensile or shear stresses found in the investigation will prove to be of greater significance than the more obvious points of stress concentration found by two-dimensional techniques. A digital computer was employed to simplify and to shorten the time necessary to obtain final results. Application of the computer to problems of this type may enable future investigators to attempt more elaborate analyses than have been previously considered feasible. THEORY Three Dimensional Stress Analysis-Shear Difference Solution:' Photoelastic studies are based on the fact that an optically isotropic transparent solid becomes optically anisotropic when subjected to forced deformation; the degree of optical anisotropy being proportional to the deformation of the material. There are two fundamental laws which govern photoelasticity: 1) Light is polarized in the directions of the principal stress axes, and is transmitted only on planes of principal stress. 2) The difference in principal stresses is proportional to the difference in the indices of refraction along the planes of principal stress. One method used in three-dimensional photoelastic work is to examine a slice of the object by normal incidence and oblique incidence. In normal incidence,

Jan 1, 1964

By J. P. Nichols

Modern vibrating screening dates from about 1910 at which time the forerunner of the present day two bearing and four bearing circular motion inclined vibrating screens were first introduced. These simple machines consisted of a screen box holding some kind of a screening surface to which was attached an off-center shaft or a shaft with an off-center weight which was rotated. The rotation of these shafts caused a shivering motion to the screen box and produced the first vibrating screen. In the 1920's and early 30's, the box designs were improved and the mechanisms designs were also improved and by 1940, vibrating screens were an important part of most process plants. In the early days when the screen sizes were dictated by how much could be shoveled onto the screen box and by the force that a particular mechanism could produce, not much thought was given to the proper sizing of the particular units. As mechanism and body design improved the ability to predict vibrating screen capacity at various openings became more important. In the late 1930's, a set of empirical capacities came to be used. with few modifications, these capacities are still in use today, both by those companies that use a formula involving only the amount of fine material in a feed, and by those companies using a formula in which the total feed to a particular screen is used. In either method of calculation, the empirical values are based on these ideal conditions: (1) The material being fed to the screen contains 25% oversize. (2) The feed to the screen contains 40% half size material. (3) The material has a bulk density of 1.6 T/Cu.M. (100 lbs/cu.ft.). (4) The screening surface has an open area of 50%. (5) The material is dry, free flowing and relatively cubical. (6) The maximum screening efficiency required is between 90 and 95%. It is obvious that these conditions do not occur in natural or crushed size distributions. Therefore, modifications must be made to these empirical values to allow for the calculation of a vibrating screen for a particular set of conditions. Before we consider these various modifications, let us look at what happens to material as it is acted on by a vibrating screen so that we can better understand why the modifications are nec-

Jan 1, 1982

By B. G. Slay, J. P. Pritchard, J. T. Pierce

A short discission is given of the cryotron us a supercozductitzg- switch. The parameters of interest such as gaiz, critical gate current, critical control current, and critical surface current density ave described. The photomask-photoresist process makes possible the study of much narrower dimensioned cryotrons than previously possible using mask stencils. Data is given for the above parameters as a function of width, thickness, temperuture, and crossing ratio. Photographs of the various circuits used are shown. The techniques involved in making connection to the circuits and the precautions taken to avoid side effects in the measurements are discussed. In particular data is given for critical currents of sipercotzducting. lines down to 2X 104 in. A vise in critical surface current density is obserced jor the strallev lines in both lead and tin. A discussion of the ihact of these measurements in the design of future cryotron circuitry is given. In addition to the improvement in device density achieved by simple scale reduction. additonal gains in density may result through improved device form factor. THE cryotron is a superconducting current switch, invented by D. A. Buck.' Fig. 1 shows a typical crossed-film cryotron. Usually the gate is made of tin and the control is made of lead. The critical gate current, is the current that drives the gate resistive with zero current flowing in the control. The critical control current, I is defined as the current in the control necessary to switch the gate resistive with a small monitoring current in the gate. The gain of the device, G, is defined as Since for uniform thin films G should be proportional to the gate width divided by the control width, it is convenient to define the critical surface current density, Jcg, to be the critical gate current divided by the gate wldth. The data which follows will show that, for a given control width, I, is relatively constant. Thus, by measuring Jcg the change in gain as a function of gate width may be measured without measuring I,.

Jan 1, 1967

By P. G. Carpenter, R. G. Nisle, F. W. Bubb

It is the purpose of this paper to describe an electrical computer which has been constructed to solve the equations for vapor-liquid equilibrium in multi-component systems. The instrument consists of seven component-computing units each with proper indicating means and power supplies. Each unit is a resistance network with a voltage matching servomechanism, and each provides an output voltage proportional to the mol-fractions for vapor and liquid phases. These voltages are summed and matched with a reference voltage to provide the solutions. Any reasonable number of such units may be put together to make a computer. The theory and operation of the computer is discussed. A number of applications and examples of computer results are given. The computer yields the over-all vapor or liquid fraction to a probable error of 0.002. An interpolation method is described which reduces the probable error to 0.0002. INTRODUCTION The process known as Flash Vaporization may be described as follows: A mixture of known composition of relatively volatile components is allowed to come to thermodynamic equilibrium at some given temperature and pressure by any path whatsoever. In general, a vapor and a liquid phase will be present at equilbrium. The problem is to determine the fractions of the mixture in the vapor and liquid phases, and to determine the mol fractions of the various components in each phase. The relations that exist between these various quantities at equilibrium are well known and will be given later. These equations are difficult to solve, yielding only to trial-and-error methods. The virtues of the analog computer are its speed and the automatic character of its calculations. The device is operated by turning a crank which varies the value of v, the total vapor fraction. This actuates a number of servomechanisms which perform automatically all computations. This computer facilitates the solution of such problems as (to mention a few] : Analysis of separator operation, studies of changes in composition of reservoir fluids with pressure decline, and analysis of natural gasoline plant operation. Examples of some of these problems are given in detail later. STATEMENT OF THE PROBLEM The essential equations are xm = 1+(km-l)v . . (1) Kmzm ym =------------------------. . (2) 1 + (Km-l)v ; n n S S xm= --------------------=1 1 + (Km-l)v m = 1 m=1 ......(3) n n S = S Km Zm =1 S ym 1 + (Km-1)v m=1 m = l ......(4) where ym = Vm/V = Mol fraction of m-th component in vapor phase. xm = Lm/L = Mol fraction of m-th component in liquid phase. Zm = Fm/F = Mol fraction of m-th component in mixture. KM = ym/xm = Equilibrium constant for m-th component at the given temperature and pressure. v = V/F = Mol fraction of vapor in the mixture. F = Total mols of a mixture of n components. V - Total mols of vapor in the mixture. L = Total mols of liquid in the mixture. Fm = Total mols of m-th component. Lm = Mols of m-th component in liquid phase. Vm = Mols of m-th component in vapor phase. Assuming the total mols of the mixture F and its composition (all the Fm) to be known — either from a quantitative analysis or from the amounts put together to make the mixture — the zm = Fm/F may be regarded as known quantities. From the given temperature and pressure of the mixture the K., are known (principally from experimental data). The primary problem is then: having given all the zm and Km to calculate v and all the fractions xm and ym. Having calculated these ratios, it is a simple matter to calculate V. L. Vm, Lm from their defining equations. SOLUTION OF THE PROBLEM Basic Computer Unit The computer consists of n units, each of which provides voltages xEo and YE, proportional to x and y as in

Jan 1, 1950

By Robert Schorr

PAGE 1. Introduction,........... 82 2. Characteristics of Briquettes,........ 87 3. The Manufacture of Coal- and Coke-Briquettes,.... 89 Binders, Organic and Inorganic,...... 90 Mixing,...........94 Presses, Continuous and Intermittent Action,....95 Cost of Briquettes and Sales-Prices in Various Countries,.. 99 4. The Manufacture of Peat-Briquettes,......101 Solid Peat-Slabs,.........101 Solid, Carbonized Peat-Briquettes,......103 Estimates of Cost of Operating Peat-Briquetting Plants,... 105 5. Manufacture of Mineral-Briquettes for Metallurgical Processes,.. 108 6. Manufacture of Artificial Stone-Masses,......112 7. Bibliography,...........115 Introduction. Briquetting is apparently a very simple operation, yet it has offered so many mechanical and chemical difficulties that it required almost a century to place it on a commercial basis. Since 1860, however, the progress has been steady and satisfactory, and to-day, for most of the European coal companies, the briquette-plant is a very profitable investment. Among the many applications of scientific knowledge to further economy by recovering and utilizing by-products and raw materials, there is possibly none which deserves the attention of Americans to a higher degree than briquetting. The recent miners' strike in Pennsylvania, which resulted in a scarce and irregular supply of coal at a high price, has advanced the possibilities for briquetting, and as a consequence public interest has grown rapidly. While the manufacture of coal-briquettes is not new in the United States, adverse conditions and sometimes lack of

Jan 1, 1905

By J. S. Kirkaldy, G. R. Purdy

THE formation of ferrite in supersaturated aus-tenite has been the subject of considerable speculation and investigation. Aaronson has recently reviewed the data pertinent to this reaction.' Since these data in all cases refer to ternary or high-order alloys it is not possible to put them to the test of simple binary diffusion formalism. It has not, therefore, been possible to make a certain statement as to the controlling mechanisms in the process. An assessment of the difficulties inherent in the conventional statistical metallographic approach suggested that a study of the problem by means of a macroscopic two-phase diffusion couple might contribute useful information. Accordingly, such a couple was synthesized, and the rate of motion of a ferrite-austenite phase boundary (the growth of an austenite layer at the expense of a ferrite layer) was determined. Wagner originally produced a diffusion solution applicable to the motion of a planar phase interface in a binary diffusion couple, assuming local equilibrium throughout the diffusion zone. This solution, which is stated free of typographical error* in Ref. where a is the position of the interface in A space (A = x/fi), C; and CL1 are the equilibrium inter-facial concentrations in phases I and 11, C: and C*1 are the initial bulk concentrations in phases I and 11, and & and are diffusion coefficients (assumed constant). If a diffusion couple consisting of homogeneous layers of ferrite and austenite in equilibrium is raised to a higher temperature, the austenite layer will grow at the expense of the ferrite. Provided that the thickness of the ferrite layer is much less than the diffusion length of carbon in ferrite, we can set , so that Wagner's solution reduces to

Jan 1, 1963

Aamot, Olav Crone, Chem. Engr., Norsk Elektrokemisk, Kongensgt. 18, Olso. Norway. '29 Abadilla.-Quirico A., Dir., Bureau of Mines Manila, P. I. '38 Abbott, Clarence E., V.P., Charge of Raw Materials, Tennessee Coal, Iron & R.R. Co., 1212 Brown-Marx Bldg., Birmingham, Ala. '04 Abel, Raymond L., Prof., Dept. of Pet. Refin., Univ. of Pittsburgh Pittsburgh, Pa. '29 Abel. Walter D., Cons. Engr 1839 Holmby Ave., W. Los Angeles. Calif. '32 Abele, Louis T R. F. D. 5. Susquehanna Trail, N., York. Pa. '31 Abell, Leo M., Ore Dressing Investigator, School of Mines Adelaide, S. Australia. '31 Abenius. Hakan V., Mine Supt., Bolidens Gruvaktiebolag Boliden, Sweden. '37 Aborn, Robert H., Asst Met., Research Lab., U. S. Steel Corp. Lincoln Highway, Kearny, N. J. '19 Abraham, W. E. V., East India United Service-Club, 16 St. James' Sq., London, S. W. 1, England. '28 Abrams. Charles Jackson 1624 17th St., Denver, Colo. '36 Abstein, Henry T., Engr. & Mgr., Profile-Tamarack Mines Co Yellow Pine, Idaho. '27

Jan 1, 1938

By S. Skowronski

The melting of cathode copper, ususally containing 95.98+ per cent. • Cu, would appear to be a simple matter. Owing to the well known affinity of copper for sulfur, however, so much sulfur is absorbed by the copper during the operation that a long and expensive refining process is required to remove it, and the final product is inferior in purity and lower in electrical conductivity than the original cathode copper. The furnace operation, as now conducted, includes the following stages: 1. Charging the cathodes into the furnace. 2. Melting the cathodes. 3. Oxidizing period—saturating the copper with cuprous sub-oxide and bringing the metal to the condition known as "set copper;'' this is now done by blowing air into the molten bath of metal. 4. Reduction period, otherwise known as "poling," whereby the oxygen previously added is removed by covering the molten bath completely with coke or charcoal and introducing large poles of green wood; this reduces the cuprous sub-oxide, bringing the copper to the "tough pitch" stage ready for casting. 5. Casting Period. Many experiments have been made to obviate the long periods of oxidation and subsequent reduction, with no success. While the copper first cast will be good, as indicated by the set surface of the bars, it is impossible to hold it to the "tough pitch" condition, and long before the whole charge has been poured the copper will become "overpoled," entailing numerous difficulties. Sulfur has always been blamed by the furnace men as the reason why this refining process is necessary, yet the literature on the subject does not sufficiently emphasize how small a proportion of sulfur will ruin copper for practical purposes. The sulfur absorbed by copper during the melting period comes from two sources: (1) the sulfate left on the cathode sheets in the tank houses; (2) the fuel. The percentage of sulfates on the cathodes is variable, and depends on the care taken in washing the cathodes after removing

Jan 1, 1919

By Frank Foote, Eric R. Jette

In a recent article,1 the authors reported the results of an X-ray investigation on the range of solid solutions that occurs in the iron-oxygen system in the vicinity of the compound FeO, which frequently is called "wüstite." The object of that investigation was primarily to study the unusual type of solid solution phenomena involved in this range. Briefly summarized the results were: 1. The NaCl type of lattice found by earlier workers was confirmed. 2. The lattice dimensions decrease rapidly with decrease in iron content. 3. No alloy that was a homogeneous single phase was ever found at compositions beyond 76.80 per cent Fe, while the theoretical iron content of FeO is 77.73 per cent Fe. 4. Measurements of the densities and lattice constants led to the conclusion that the form of substitution was complex, an FeO group being replaced by oxygen. Such a substitution, of course, is equivalent to removing an iron atom (or ion). The ideal unit cell of FeO can be considered as containing four groups of FeO and, thus, when this type of substitution takes place, there remain in the unit cell three FeO groups and an oxygen, or a total of seven atoms with a gross formula of Fe3O4 instead of the original Fe4O4. 5. This was the only interpretation that would fit the experimental results and at the same time be consistent with (a) the previously determined fact that the FeO lattice is ionic (i, e., consists of Fe++ and O? ions) and (b) the chemical fact that as the oxygen content of the substance increases, ferrous iron becomes oxidized to ferric iron. The experimental results of this investigation are included in Fig. 1. The purpose of the present investigation was to determine the limits of this homogeneity range by means of the X-ray structure analysis. This method has been applied frequently to the study of solid solutions by Westgren and his coworkers as well as by other investigators. The principle is simple: if the change in lattice constant with composition is

Jan 1, 1933

By M. J. Marcinkowski, A. S. Sastri

A detailed compressive stress-strain analysis and transmission electron microscopy investigation has been made of the deformation behavior occurring in a 50 at. pct Ni-Ti (hypoeutectoid) alloy and a 54.5 at. pct Ni-Ti (hypereutectoid) alloy. In the case of the hypoeutectoid alloy, three stages of work hardening are observed. Stage I occurs at a very low stress and is associated with plastic deformation via martensite formation. Stage 11 is characterized by very rapid work hardening and is due to difficulties in causing further deformation in the fine martensite aggregate produced in Stage I. Stage III which occurs at very high stress levels is characterized by smaller work hardening rates and is due to the plastic deformation arising from alternate reconversions of the original martensites to martensites of varying orientation. Rapid quenching of the hypereutectoid alloy leads to very high yield strengths and is related to a fine precipitate dispersion that such treatment brings about. The present investigation represents the final phase of a three-part study directed toward an understanding of the solid-state transformations in near equi-atomic Ni-Ti alloys as well as the deformation mechanisms associated with these alloys. In the first part,"2 to be henceforth referred to as I, it was found that alternate simple shears on {112} planes and in (111) directions convert the parent B2 structure in the equiatomic NiTi alloy into two distinct close-packed monoclinic martensites. All of the marten-sites were of this type, whether they were formed by cooling or by plastic deformation, whether induced to form in bulk samples or in thin foils, or whether examined in the electron microscope at room temperature or below. On the other hand, in the second part of this investigation,3 to be reffered to as 11, it was shown that upon slow cooling to about 640°C. alloys in the neighborhood of NiTi which possess the B2 structure transform eutectoidally into their equilibrium phases Ti2Ni and TiNi3. However, preceding the formation of these equilibrium phases a series of metastable intermediate phases are formed. This paper will set as its goal the elucidation of the remarkable deformation behavior exhibited by NiTi. In particular, Buehler and Wiley4 have found equiatomic NiTi to be surprisingly soft, while Buehler et al.5 have shown this alloy to possess a memory effect: i.e., upon bending at room temperature it will revert to its original shape when heated to above about 50°C. In I it was shown that NiTi was soft in the sense that the yield stress was low; nevertheless, the alloy work-hardened at an extremely rapid rate to very high stress levels. On the other hand, the hypereutectoid alloys with somewhat higher nickel, say 54.5 at. pct (60 wt pct) have enormously increased yield strengths compared to those of the equiatomic alloys. In order to determine the atomistic processes giving rise to the above behavior, it was decided to examine samples that were wafered from bulk specimens deformed in compression to various strains using the techniques of transmission electron microscopy. EXPERIMENTAL TECHNIQUE All of the alloys used in the present investigation contained either 50 at. pct Ni (55.06 wt pct) or 54.5 at. pct Ni (60 wt pct) and were arc-melted in the form of a finger using the same techniques described in I and II. The finger was capsulated in a stainless-steel jacket and swaged at 850°C into rods. Compression specimens 0.300 in, long and 0.200 in. in diam were machined from these rods. In order to completely re-crystallize the samples and remove residual stresses, all of them were capsulated in evacuated quartz, annealed for 1/2 at 1050°C. and then furnace-cooled. Compression tests were carried out in an Instron tensile testing machine covering a range of temperatures from —196° to 200°C using procedures described previously.6'7 In all cases crosshead speed was 0.02 in. per min. Wafers 0.015 in. thick were spark-cut from the cylindrical samples at 45 deg to the compression axes after they had been deformed to the desired strain. These specimens were then spark-planed to about 0.005 in. and then electrochemically thinned for examination by transmission electron microscopy as described in I.

Jan 1, 1969

By F. R. Hewitt

The methods of mining talc are simple, and in western North Carolina are almost entirely by open cut and quarry. The larger part of the talc of this section lies in various-sized "veins''. inclosed in quartzitic walls, the majority of which have by folding been thrown into a perpendicular position, or nearly so, and the exposed edges of which have been covered by debris from erosion of the mountain above. This covering of loose rocks and earth is from 5 to 25 ft. thick, and in most cases has to be removed because it is difficult to support. Sometimes the covering is heavy enough to "catch" and timber successfully, and in this case the talc can be worked out by the usual method of following the "vein," using stulls for holding up the walk until the vein is worked out, and then allowing the cut or drift to fall in. In some cases it proves more convenient to sink shafts and run drifts, following the talc "veins" until exhausted. The talc deposits of this section are badly broken and faulted both laterally and perpendicularly, and in some cases are found many feet below the level of drainage, necessitating pumping at considerable expense. My own experience in mining talc has been that the chief point is to get a reliable "vein," and that then the mining is not as difficult, or any more so at least, than that of any common ore. If the "veins" are pure talc it is simple; if the talc is admixed with foreign matter, such as tre-molite in excess, or stains of iron or manganese, as is often the case, one had better abandon the deposit. Methods of Manufacture Talc, being a soft mineral, is not difficult to reduce to the condition desired by the trade. Powdered talc is admitted under low duty from France, Italy and Austria, and some is being imported from Asiatic points. Talc is usually ground and bolted in any simple reduction mill. That mined in this section is hand-sorted and the quality that will make good powdered talc is ground for use in the manufacture of talcum powder, cosmetics, etc., while all that is hard enough and has sufficient strength is used for crayons and blanks for gas burners and electrical work, being sawed into the shapes desired. The rest that is clean and pure is ground

Jan 1, 1917

By E. W. Davis

IF the tonnage of merchantable iron ore remaining in the Lake Superior district is divided by the average of the annual shipments for the past 20 years, it will be found that this ore supply will be exhausted in approximately 35 years. This computed exhaustion period would be an indication of the future activities of the iron-mining industry if it were not for the fact that merchantable ore is being manufactured from low-grade ore, of which there is an almost inexhaustible supply in the Lake Superior district. It is true that at present only the simplest of these low-grade ores are being concentrated to any considerable extent, but progress is being made with the treatment of the more complex ores, and now, in addition to 20 washing plants on the Mesabi Range, three plants are equipped with jigs for the treatment of ore that cannot be concentrated simply by washing. The latest addition to the ore-treating plants of the Mesabi Range is a magnetic roasting and concentration plant at Cooley, Minn., in which ore that cannot be concentrated either by washing or by jigging is roasted to the magnetic state and concentrated on magnetic separators. This plant is beginning its third season of operation, and while it is too soon to make definite predictions, there are strong indica-tions that this new process may make commercially possible the manu-facture of high-grade, merchantable iron ore from large quantities of low-grade ore now considered worthless. After the iron oxide contained in an ore has been converted to mag-netite, almost any desired grade of concentrate can be produced, using magnetic-concentration equipment of the types now in general use for the concentration of the natural magnetites of New York, New Jersey and Pennsylvania. The conversion of hematite (Fe203) to magnetite (Fe304) is easily accomplished by heating the ore to a dull red tempera-ture in a reducing atmosphere. The reaction is slightly endothermic, and while it requires 200 B.t.u. to heat to 900° F. one pound of hematite ore containing 45 per cent Fe, it requires only 100 B.t.u. to convert the oxide into magnetite. In the laboratory, using small quantities of ore, this conversion is very simple, but large furnaces suitable for commercial

Jan 1, 1936

By C. R. Van Hise

PART I.-GENERAL PRINCIPLES. [ ]

Jan 1, 1902