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By H. W. Camp

In attempting to summarize and pick out the outstanding development,s in refinery engineering during the past 12 or 13 months, one is immediately impressed by the great strides that have taken place. It is, of course, difficult to obtain the proper perspective with the old year so fresh in our minds, and if relatively minor improvements seem to be given equal recognition with those of magnitude, it will be because the nature of the human mind causes us to place the trivial incidents of today with the important things of yesterday. For instance, the Twenly-Second Semiannual Motor Gasoline Survey (December, 1930) of the U. S. Bureau of Mines shows that the total gasoline yield per barrel of crude increased from 39.6 per cent.. in 1929 to 41.99 per cent. in 1930. This latter fact may appear rather confusing since the production of high-gravity crudes has been steadily increasing. However, during the past yew there has been a reduction in gasoline end point of approximately 8" F. which roughly corresponds to about 1.0 per cent. yield from the crude oil. It appears, therefore, that, if the gasoline mere based on a corresponding distillation basis 1930 would show better yields from the crude than 1929. Cracked gasoline increased from 14.83 per cent. in 1929 to 17.46 per cent. in 1930. Percentage of total gasoline from cracking stills has doubled since 192.5, thus indicating the extent to which this phase of refining has advanced over past years. Replacement of the old shell still batteries by a pipe still and single flash system, with an efficient fractionating tower, has proved advantageous both in fuel consumption and in decreasing space required for the installation. Considerable activity has been shown in the field of vapor-phase cracking, with the development of several new types of units. The production of a product with a high antiknock value makes the vapor-phase processes particularly attractive. Cracking equipment has shown a marked advance with the development of corrosion-resisting alloys. Nickel, chromium and vanadium have shown interesting possibilities, but as yet are fairly expensive. Ceramic material for lining reaction

Jan 1, 1931

John Langton, New York, N. Y. (communication to the Secretary*):—In Mr. Firmstone's discussion (Trans., xli., 912) the formula he gives for the periodical payment—his equation (2)—may be simplified in form. The expression given by Mr. Firmstone is : q=S(l+ r) 1-(1+r)-(1+r) r which is a simpler form to use in calculations. Where S= principal to bc extinguished. Q = periodical payment or charge for sinking-fund, r = rate of interest per period (i. e., interest on $1 per period). n = number of payments or periods. Kent's Mechanical Enyineer's Pocket Book, under the heading Annuities, equation No. 5, gives : " The annuity which $1 will purchase for any number of ytlars n is * Received Jiily 16, 1911.

Jan 1, 1912

By James B. Duke, Ernest W. Greene

An idea of what is meant by "Ultrafine Minerals Or Slimes" as used in the title can be had by an examination of Fig. I. Particle size distributions are presented for a crude kaolin clay, a coarse fraction and a fine fraction. The crude kaolin as mined was blunged with water, degritted through a 325-mesh Tyler screen, and dispersed with sodium silicate. The dispersed crude was then centrifuged in the laboratory into the coarse and fine fractions shown in the figure. The distribution curves were obtained by the well known sedimentation technique of Casagrande' using a long arm centrifuge to obtain equivalent spherical diameters down to 0.2 microns on the finer fraction. Fig. 1 shows the coarse fraction kaolin to be 50% finer than 3 microns, while the crude kaolin is 50% finer than 1.4 microns, and the fine fraction 50% finer than 0.6 microns. For comparison 53 microns on the horizontal scale would represent Stokesian diameters which would just pass a 325- mesh Tyler screen, taking into consideration the Andreasen sieve correction factor of 1.22 X the screen opening.' Another helpful comparison is the surface areas, which in square meters per gram, are 0.044 for 53 micron particles, 0.77 for 3 microns, 1.65 for 1.4 microns, and 5.0 for 0.46 micron particles. Fig. 2 is an electron micrograph of a similar fine fraction kaolin showing the hexagonal plate- like particle shapes.

Jan 10, 1962

By H. S. Drinker

THE first discovery of zinc 011 the property now worked by this company was made by the celebrated mineralogist, Prof. William Theodore Rapper, in 1845. Different claimants kept the property in continual litigation till 1861, when everything was satisfactorily settled by the purchase of the land by the company, who have worked it ever since. The mines are situated in a valley, four miles south of Bethlehem, Pa., in Lehigh County. The valley is from two to three miles broad, and is bounded 011 the north and south by syenitic hills. Syenite underlies the valley, and upon this is deposited the Potsdam sandstone, then a layer of slate, and, finally, a dolomitic limestone, belonging to the Chazy or Black River formation, in which the zinc deposits are inclosed. The rocks occur in folds, and have evidently been subjected to tremendous pressure from below, the zinc deposit being found in an anticlinal axis, in nearly vertical veins. I t is supposed that the zinc was originally disseminated through the

By Herbert Clark Hoover, Lou Henry Hoover

DIFFERENT methods of parting gold from silver, and, on the other hand, silver from gold, were discussed in the last book; also the separation of copper from the latter and further, of lead from gold as well as from silver ; and, lastly, the methods for refining the two precious metals. Now I will speak of the methods by which silver must be separated from copper, and likewise from iron.1 The officina, or the building necessary for the purposes and use of those who separate silver from copper, is constructed in this manner. First, four long walls are built, of which the first, which is parallel with the bank of a stream, and the second, are both two hundred and sixty-four feet long. The second, however, stops at one hundred and fifty-one feet, and after, as it were, a break for a length of twenty-four feet, it continues again until it is of a length equal to the first wall. The third wall is one hundred and twenty feet long, starting at a point opposite the sixty-seventh foot of the other walls, and reaching to their one hundred and eighty-sixth foot.

Jan 1, 1950

By K. Mukherjee

A quasi-thermodynamic calculation is presented for estirt~ating vacancy Parameters in fcc metals. For fcc metals the equilibrium vacancy concentration at the melting point can be approximated by a constant mole fraction. The Gibbs free energy of formation of the vacancies can then be expressed as a simple function of the absolute melting temperature alone. From this description the formation energy Ef, the vibrational entropy Sv, and the formation volume of a vacancy can be calculated in terms of known thermodynamic parameters. The calculated values are in good agree-rnent with available experimental results. IT appears that many of the physical parameters associated with lattice defects can be empirically related to thermodynamic melting parameters.'-7 Besides the implicit correlations between melting phenomenon and point defects, the melting temperature itself at atmospheric pressure can be used as a reference temperature for the purpose of comparison of vacancy concentrations in a given crystal class. The concentration at the melting point then corresponds to the maximum vacancy super saturation in the crystalline state at a fixed pressure. For pure fcc metals such as Au, Cu, Al, Ag, Pb, and Pt the mole fraction of vacancies at the melting point8 is of the order of 10*"3 to 10"1. The equation expressing the equilibrium vacancy concentration at the melting point contains three parameters: the formation energy, the vibrational entropy, and the absolute melting temperature, see Eq. [2]. Since these three parameters cover a wide range of values8 in the above-mentioned metals, it can be assumed that in other fcc metals, for which no reliable data are available, the mole fraction of vacancies at the melting point most likely lies in the range 10~3 to lO"4. This is then the basis of the calculations described in this article. Within this range the mixing entropy can be linearized and the vibrational entropy, the formation energy, and the vacancy formation volume can be calculated in terms of known thermodynamic constants. 1} LINEARIZATION OF MIXING ENTROPY The Helmholtz free energy of a crystal containing n vacancies at the melting point can be writteng as where Ef, T,, k, and S, are the formation energy of a vacancy, the absolute melting temperature, Boltz-mann's constant, and the vacancy formation entropy, respectively. After applying Stirling's approximation in Eq. [l] and minimizing with respect to n where (n/N), is the equilibrium vacancy concentration in mole fraction at the melting temperature at 1 atm pressure. If instead of setting (6~/6n) = 0 immediately after Stirling's approximation in Eq. [I], the configurational partition function is simplified, then some useful and significant results can be obtained by utilizing the experimental inforrnation After applying Stirling's approximation and rearranging is plotted as a function of (n/N) in Fig. 1. The points, not all of which are shown, are fitted by a straight line substituting from Eq. [5] into Eq. [I] and minimizing with respect to n combining Eqs. [6] and [2] It must be emphasized that Eq. [7] is only a crude approximation based on an oversimplified model of

Jan 1, 1967

By K. E. Brown, A. R. Hagedorn

A 1,500-ft experimental well was used to study the pressure gradients occurring during continuous, vertical, two-phase flow through 1-in., 1 1/4-in. and 1 1/3-in. nominal size tubing. The test well was equipped with two gas-lijt valves and four Mathak electronic pressure transmitters as well as with instruments to measure the liquid production rate, air injection rate, temperatures and surface pressures. Tests were conducted for widely varying liquid flow, rates, gas-liquid ratios and liquid viscosities. From these data, an accucrate pressure-depth traverse was construcred for each test in each of the three tubing sizes. From the results of these tests, correlations have beer] developed which allow the accurate prediction of flowing pressure gradients for a wide variety of tubing sizes, flow conditions, and liquid properties. Also. the correlations and equations which are developed satisfy the necessary condition that they reduce to the relationships approprate to single-phase flow when the flow rate of either the gas or the liqutid phase becomes zero. All the correlations involve only dimensionless groups. which is a condition usually Sought for in similarity analysis but not always achieved. The correlations developed in this study have been used to calculate pressure gradients for pipes of larger diameter than those upon which the correlations are based. Comparisons of these calcuclaed gradients with experimetally determined gradients for the same flow conditions obtained from the literature indicate that extrapolation to these larger pipe sizes is possible with a degree of accuracy sufficient for engineering calculations. The extent of this extrapolation can only he determined with additional data from larger pipe diameters. INTRODUCTION The accurate prediction of the pressure drop expected to occur during the multiphase flow of fluids in the flow string of a well is a wideiy recognized problem in the petroleum industry. The problem has been brought even more into prominence with the advent of tubingless or slim-hole completions which use small-diameter tubing. Many of the correlations which give reasonably accurale results in the lager tubing sizes are greatly in error when applied to small-diameter conduits. Small-diameter conduits are defined as 1 1/2-in. nominal size tubing or smaller. The study of the pressure gradients which occur during multiphase flow of fluids in pipes is exceedingly complex because of the large number of variables involved. Further difficulties relate to the possibility of numerous flow regimes of widely varying geometry and mechanism and the instabilities of the fluid interfaces involved. Consequently, a solution to the problem by the approach normally used in classical fluid dynamics based on the formulation and solution of the Navier-Stokes equation has not been forthcoming. This is primarily the result of the nonlincarities involved and the difficulty of adequately describing the boundary conditions. As a result of the foregoing, most investigators have chosen semi-empirical or purely empirical approaches in an efford to obtain a practical solution to the problem. Much of the previous work in this area was done in short-tube models in the laboratory. A number of problems arise, however, when attempts are made to extrapolate these laboratory results to oilfield conditions where a much longer tube is encountered. In those few studies where data taken in long tubes were utilized, the data covered only a limited range of the variables, and as a result inaccuracies are introduced when the correlations are extended outside the range of the original data. Also, as a consequence of the limited amount of data available for these studies, the effects of several important variables were overlooked. The problem of predicting the pressure drop which occurs in multiphase flow differs from that of single-phase flow in that another source of pressure loss is introduced, namely, those pressure losses arising from slippage between the phases. This slippage is a result of the difference between the integrated average linear velocities of the two phases, which in turn is due to the physical properties of the fluids involved. In contrast to single-phase flow. the pressure losses in multiphase flow do not always increase with a decrease in the size of the conduit or an increase in production rate. This is attributed to the presence of the gas phase which tends to slip by the liquid phase without actually contributing to its lift. Many investigators have attempted to correlate both the slippage losses and the friction losses by means of a single energy-loss factor analogous to the one used in the single-phase flow problem. In an approach of this type, hourever, many of the important variables, such as

Jan 1, 1966

By R. M. Hays

At the present time, oxidized taconite and semitaconite are not processed on the Mesabi Range except in vely limited tonnages. However, it is estimated that by 1990 shipments from the Range will approach 75 million tons of iron ore per year. To achieve this and future production, either additional magnetic taconite reserves must be proven or oxidized taconite and semitaconite must be processed. One beneficiation process which uses high-intensity magnetic separation and flotation for concentrating oxidized taconite and semitaconite is described. Detailed metallurgical results are given and an economic evaluation of the process is presented with comparison to other processes. The results of the present study indicate that the high-intensity magnetic separation and flotation process appears nearer commercialization for the processing of Mesabi Range semitaconite. The disadvantages of the process are nearer solution and are being solved by well-defined programs. During the twentieth century, the Mesabi Range in northeastern Minnesota has been the main source of iron ore for American steel mills. The Mesabi Range, which is about 120 miles long and up to 3 miles wide, is the pre-glacial outcrop of a group of sedimentary iron-bearing rocks known as the Biwabik formation. The formation, which is 200 to 650 ft thick and dips 50 to 10° to the southeast, is divided into four stratigraphic units. From top to bottom, these are the Upper Slaty, Upper Cherty, Lower Slaty, and Lower Cherty. Since its deposition in Precambrian time, most of the original Biwabik formation has been subject to alteration. Thus, although the iron content of the original formation was quite uniform, major variations in mineral composition now occur both laterally along the length of the formation and vertically within its units. Details of the mineralogical variations in the Biwabik formation have been described by Wolff,' Gruner, 2 and White.3 The unaltered portion of the Biwabik formation includes the magnetite taconites which are being commercially processed by Reserve Mining, Erie Mining, Eveleth Taconite, National Steel, Butler Taconite, and U.S. Steel. These six operations have a capacity of 32 million tons of iron ore pellets per year and have produced approximately 170milion tons of pellets since 1956. Capacity for another 30 million tons of pellets per year is being considered by the industry for the next ten years. This increase will be obtained by constructing new plants and by expanding present operations. Thus, by ! 978, it is estimated that approximately 60 million tons of iron ore vellets will be produced each vear from the magnetite taconites of the Biwabik formation. The portions of the Biwabik formation subjected to intense alteration resulted in direct shipping, wash, and heavy-media ores. Since 1892, 1.8 billion tons of direct shipping ore have been shipped from the Mesabi Range. This ore has been the major source of iron ore for American steel mills. In 1967, about 11 million tons of direct shipping ore and 13 million tons of wash, jig, heavy-media, and other gravity concentrates were shipped from the Mesabi Range.4 These products generally have poorer physical and chemical properties than the iron pellets produced from magnetic taconite. It is predicted that the shipment of these materials will decrease from 24 million tons in 1967 to 6 to 8 million tons in 1975.' This decrease will

Jan 1, 1970

By F. D. Weeks

The Salida smelting plant, owned by the Ohio & Colorado Smelting & Refining Co., is situated at an altitude of 7,000 ft., about 2 miles west of Salida, Colo., and 215 miles southwest of Denver. Salida is a division point on the Denver & Rio Grande R. R. where the narrow-gage line serving the San Juan region joins the through standard gage. This narrow-gage road also serves the coal and coke companies at Crested Butte and the lime quarries and mines at Monarch and Garfield, which are only 20 miles distant. These facts, together with the possibility of an appropriate site, doubtless influenced the builders to locate at this point. The smelting plant and yard occupy 80 acres of land on the northeast side of the Arkansas River. The slope of the ground is such as to permit of the ore being delivered by cars at a point 25 ft. above the feed floor; and in front of the furnaces there is a fall of 50 ft. for the slag dump. A recent purchase of river bottom has provided slag room for 20 years. As originally constructed, the plant consisted of four blast furnaces, 40 by 144 in. at tuyéres, two copper-matting furnaces, dust chamber, stack, power plant, blowers, sampling works, and one railroad trestle about 16 ft. from bed floor to rails. Later 12 hand roasters of excellent design were added and the practice of smelting leady matte in a copper-matting furnace was abandoned. Since the plant passed under the present management the equipment has been increased so that the following is a brief description of the same at this date: Two lead blast furnaces, 48 by 180 in. at tuyéres, 17 ft. 6 in. tuyéres to feed floor; four lead blast furnaces, 40 by 144 in. at tuyéres, 17 ft. 6 in. tuyéree to feed floor; two copper-matting furnaces, not used; blast-furnace flue, 900 ft.; area 245 sq. ft.; blast-furnace stack, 150 ft.; above furnace floor, 190 ft. When smelting a 500-ton charge per day, the gases approximate 63,700 cu. ft. per minute and travel at a speed of 260 ft. per minute. The slag is hauled by a 10-ton electric locomotive; the slag shells and settler refuse are elevated to a bin on the feed floor by an electric hoist. The matte is tapped in 1,600-lb. cakes and loaded into cars

Jan 1, 1916

By W. Hurst

What is entailed here is the extension of the sinzplified material balance formulas to encompass interference between oil fields. As previously reported, the ex-plicitness as so revealed for the cunzulative pressure drop as a function of all factors contributing to its change in the material balance equation, is now transcribed through the inter-conzmunicating aquifer to effect an increased pressure drop on an adjoining field by interference. Such is performed by mathematical analyses and the application of the Laplace transformations. What is accomplished is that the reiteration problems previously associated with interference studies are nullified, since volumetric changes for the fluids in situ are automatically adjusted by the explicitness so expressed, and what pertains to the superposition principle applies only to the impeded water drive upon a su.bject field, which is likewise incorporated in the over-all pressure drop that results. The mathematics treats with the rigorous solution of the problem, as well as with methods easily amenable to numerical interpretation by the practicing engineer. In all cases, however, what is deduced for areal extents, reveal interference between wells when time becomes large that further substantiate the analyses. Probably of equal significance, all variables and for/nation characteristics are accounted for. This applies to the differences in PVT analyses that occur from field to field and the physical parameters associated with the lithology of the reservoirs. The latter is deduced from a rigorous interpretation of the unsteady-state flow problem for sands of different permeabilities in series. Thus, what purports to be a trial-and-error calculation to include variations in sand conditions within the intercommunicating aquifer to define water drive has little importance compared to the designation of such paranzeters at the reservoir to constitute the essential criteria. INTRODUCTION In a recent publication, the reader has been introduced to the simplified material balance formulas,' It has been shown that the complexities formerly associated for identifying and determining reservoir pressure in the material balance relationship can be resolved by treating with the Laplace transformations, the inversion of which reveals pressure as an explicit function of all factors contributing to its change. For an under saturated oil reservoir, this constitutes an integrated effect from the inception of production; and for a saturated oil reservoir, such represents a survey traverse. However, what is most important, for the first time we are able to express the pressure change in a reservoir directly to the oil, gas and water produced, and to the physical parameters of the formation, in the same manner that reservoir pressure has served in the past to define well interference — which opens up this avenue of endeavor for investigating interference between oil fields. The occurrence of interference is by virtue of the juxtaposition of two or more oil reservoirs producing in the same aquifer. Thus, in any one field, the extended production so incurred creates a pressure lowering in the field, as well as in the aquifer, which can now be transcribed by these simplified material balance formulas to the adjoining reservoir to observe the increased pressure change so induced. These effects are retroactive from field to field, to reveal a pressure drop far in excess of normal depletion to evidence the interference pattern that can result. Such is the purpose for the present undertaking — to develop the mathematical physics treating with this phenomenon of interference between oil fields, employing the simplified material balance formulas. Since Ref. 1 often will be referred to in the text, it is identified as "The Simplification", in the same connotation that another paper has been referred to as the "Laplace Trans-formation. However, the amount of material to be covered in these passages is substantial; therefore, the total emphasis will be placed upon developing these fundamental ideas treating with interference rather than attempting to offer work curves, which always can be prepared. For the practical application of these simplified material balance formulas that form the framework of the present development, the reader will find a comprehensive treatment in the Appendix of Ref. 1, illustrating a factual field example. In essence, what is proposed by this presentation is to make available to the practicing engineer a simplified means to treat with interference between oil fields, independent of the complex machine computations that have attended such undertakings in the past. As the problem unfolds, a common pattern will be discernible, whereby the many reiteration processes that previously required computing devices are eliminated.

By Ian Wark

IN the paper on the development of the flotation process at Broken Hill (Australia) prepared by the Broken Hill Branch of the Australasian Institute of Mining and Metallurgy and published in its Proceedings for 1930 (N. S. 80), it is stated that the concentration of ores by flotation has been established on a commercial basis in Australia since 1904 and that differential flotation was practiced at Broken Hill in 1912. Since then, progress in flotation practice throughout the world has been rapid, and during recent years the knowledge of the reactions between comparatively modern so-called chemical promoters or collecting agents and the mineral surfaces has been broadened by the published accounts off many investigators, particularly those working in American universities. Because of the benefits the Australian mining industry has derived from the flotation process, it was felt that the industry should take some part in the efforts that are being made to extend the knowledge of the principles on which the flotation process is based. Accordingly, early in 1930, three Broken Hill mining companies-the Broken Hill South, the North Broken Hill, and the Zinc Corporation-together with the Electrolytic Zinc Co. of Australasia, Ltd., and the Mount Lyell Mining & Railway Co., established a laboratory in the chemistry department of the University of Melbourne for the purpose of investigating the principles underlying flotation. Subsequently, the Burma Corporation, Ltd., expressed a desire to cooperate. The two authors are directly employed and controlled by these six organizations, and the cost of the investigation is divided equally among them. The research has been carried out under the general direction and in close cooperation with Harry Hey, chief metallurgist for the Electrolytic Zinc Co. of Australasia, Ltd., and the authors desire to place on record their indebtedness to him for his ever willing help and constructive criticism.

Jan 1, 1932

By W. Rostoker, R. P. Elliott, B. W. Levinger

Phase equilibria in the Ti-Mn-Mo system have been investigated in the composition range 100 to 60 pct Ti and in the temperature range 550 to 1150°C. Three out of ten isothermal sections are presented as well as seven vertical sections, projections of the beta space and the surface of incipient melting. A STUDY of the system Ti-Mn-Mo was undertaken as part of a program to aid in the development of titanium-base phase equilibrium diagrams of potential technical importance to alloy development. The scope of the investigation included the composition range 100 to 60 wt pct Ti (all compositions refer to weight percentages) and the temperature range 550" to 1150°C. A surface of incipient melting was constructed from measurements of a large number of alloys. In the binary system Ti-Mo,' the high temperature modification of titanium generally spoken of as the p phase is stabilized to successively lower temperatures with increasing alloy content. There are no known intermediate phases, and complete miscibility exists between body-centered cubic titanium and molybdenum. The u solid solution is limited to less than 0.8 pct Mo at temperatures between 885" and 600°C. An equilibrium diagram for the binary system Ti-Mn has been published.' Manganese behaves as a p stabilizer but undergoes a eutectoid transformation at about 20 pct Mn and 550°C to a plus an intermediate phase. The structure of this phase has been found to be isomorphous with the a phase FeCr." Accordingly, the phase is assigned the formula TiMn, although its actual location on the composition scale has not been determined and may well be not exactly at the equiatomic composition. In disagreement with ref. 2, the origin of the TiMn phase has been shown' to be a peritectoid transformation at a temperature between 900" and 1000°C. A second intermediate phase occurs at the atomic proportions TiMn,. The structure of this phase was originally established by Wallbaum" as having a hexagonal lattice (C 14 type, 12 atoms per unit cell) isomorphous with MgZn,. In this work, the structure determination was confirmed and lattice parameters determined to be: a = 4.815 kX, c - 7.901 kX. Because of the extensive range of the P phase, the ternary system was suited to the determination of the phase boundaries by the parametric method of Andersen and Jette." During the course of the investigation, this method was found applicable only to the P/a+P boundaries but not to P/P+ compound boundaries. Metallographic methods were used alternatively for delineating these phase boundaries. Preparation and Treatment of Alloys Over 70 alloys were prepared for the study of this system. Ten gram alloy buttons were arc melted in a nonconsumable electrode furnace using inert atmosphere protection, and a water cooled copper crucible. The construction and operation of this type of melting unit has been described.' To insure complete solution of alloy additions, especially molybdenum, the buttons were remelted alternately on top or bottom as many as five times. This was accomplished by a suitable mechanism for "flipping" the ingots over. It was very difficult to obtain homogeneous melts in the region of less than 75 pct Ti because of the very high melting point of molybdenum and Ti-Mo master alloys and the comparatively low boiling point of manganese. Thus, the arc conditions for insuring complete solution of molybdenum were incompatible with the prevention of heavy losses of

Jan 1, 1955

By Joseph Singewald

INTRODUCTION FOR a number of years Russia, India and Brazil have outranked .all other countries as producers of manganese ores. During the 5 years immediately preceding the European war, the average annual production of Russia was 844,000, of India 694,000 and of Brazil 200,000 long tons. Since the outbreak of the war there has been a considerable falling off in the Russian and Indian production, particularly in the former, whereas the production of Brazilian ores has been greatly increased, amounting in 1914 to over 250,000 tons and in 1915 to nearly 350,000, with conditions favorable for a still larger production in 1916. Manganese is one of the few industrially important metals that are not produced in the United States in quantity commensurate with our needs, so that we have been compelled to import annually about 300,000 tons, having a value of over $2,000,000, and these ores have been obtained from the three countries mentioned above. Their rank as contributors to our imports of manganese ores was India, Russia, Brazil; and in 1913, Brazil contributed only one-fifth. During the past 2 years Brazil has furnished rapidly increasing quantities to this country, and, with the falling off of imports from India and Russia, has become our principal foreign source. In 1914, Brazil furnished two-fifths of the imports and more than India or Russia, and in 1915 over nine-tenths. The manganese ores of Brazil are consequently of more than usual interest to us at this time.

Jan 10, 1916

AS provided in Article VIII, Section 4 of the By- laws, the Papers and Publications Committee has formulated the following, rules of pro-cedure and submitted them to the Board of Directors for consideration and approval. In preparing these rules the committee, and an active subcommittee con-sisting of Messrs. Norris, Sharpless; Fohs and Wither-ell, have kept in mind the desirability of codifying the practice that has grown up through the now sev-eral years of experience of the committtee, without casting them in a mold so rigid as to be hampering. The problems of the Institute change from year to year and the methods of publication must change also to meet each new situation as it arises. "The estab-lishment of divisions and special volumes devoted to the papers. read before each required certain changes from former practice. These rules have been framed with this in view. The committee suggested, however, in its letter of transmittal, that the system will be made all the more workable if the Board will, as far as practicable, in making appointments to this committee, arrange for representation from each of the professional divisions and technical committees. It is understood that the chairman of the committee will be named by the Board. The rules regarding spe-cial volumes have been framed, keeping in mind the present policy of distribution, as already approved by the Directors under which the special volumes go only to the special. groups except on request. The committee has not attempted to provide rules governing consideration of any appeal made to the Board from one of its decisions, but its services and files will necessarily be at the full service of the Board in any such event. The rules which were ap-proved by the Board at its meeting on March 30, 1928, are as follows: Signed by C. S. Witherell, Chairman. (A) The authority of. the Committee on Papers and Pub-lications, hereinafter called the Committee, is defined in Art. VIII, Sec. 4 of the By-laws of the Institute, which reads as follows: "Papers and Publications: The Committee on Papers and Publications shall consist of the Secretary of the Institute, and at least twelve members of the Institute, to assist in passing on all papers offered for publication. They shall be appointed annually by the Board of Direc-tors, on recommendation of the President, and shall have authority, subject to appeal to the Board, to accept or reject papers for publication in the Transactions and Technical Publications of the Institute. The Commit-tee shall formulate a set of rules of procedure which, when approved by the Board, shall have the force of By-laws until revoked or amended." The policy of the Committee will be to uphold the ob-jects of -the Institute as set forth in the Constitution (Arti-cle 1, Sec. 2).

Jan 5, 1928

By J. T. Crawford

Although nearly 10 pct of the total tonnage of coal produced annually within the United States is handled by bulk freighters on the Great Lakes, very little of the detail connected with it has been published other than occasional newspaper stories and publication of tonnage statistics. Of the total tonnage floated on the Lakes each year some 10,000,000 is stored and distributed from the port of Duluth-Superior, at the western end of Lake Superior commonly known as the Head of the Lakes. This port has the largest single area concentration of coal docks in the world. Since this area contains the largest ore docks, the largest movable material handling bridge, the largest and highest grain elevator and the largest coal bri-quetting plant in the world, it is entirely fitting and proper that here also should be located the largest coal dock and what we believe to be the world's largest clam shell. Of the sixteen coal docks operated by ten companies, five are owned and operated by the North Western-Hanna Fuel Co. which has two docks on the Superior, Wis., waterfront and three docks in Duluth, Minn. It is with these five docks that we are primarily concerned, General History In the summer of 1871 a small sailing vessel entered the harbor of Duluth-Superior with the first commercial coal cargo. All the coal brought up that first year did not amount to more than 3000 tons. During the year 1877 the first dock equipped for handling coal was built in Duluth. Coal receipts increased to 52,785 tons in 1879 the first year for which an official record was kept. Since then the volume of water-borne coal to the Head of the Lakes steadily increased to a maximum of 12,688,321 tons in the year 1923. This tonnage was nearly equalled in the year 1927 and the next highest tonnage recent year was in 1946 when 10,105,703 tons were unloaded. The average annual bring-up over a ten year period 1938 to 1947 was 8,605,231 tons. Approximately 30 pct of the coal unloaded at the Head of the Lakes is handled over the docks of the North Western-Hanna Fuel Co. Competition of other fuels coupled with expansion of coal fields in the mid-west have held coal receipts for Duluth-Superior at a relatively constant figure during the last eight years although the total tonnage of coal floated on the Great Lakes has more than doubled in the past 25 years. From the shovel and wheelbarrow method of unloading early cargoes to the horsepowered windlass derrick with a wooden tub was but a short step. The first movable coal handling, steam operated,

Jan 1, 1949

By Samuel Storchheim

IT was noted that variations existed in the magnetic properties, namely, coercive force, He, and rema-nence, Br, of 18/8 stainless steel wires of the same analysis when these wires were given the same percentage of reductiori of area at different rates of reduction. In order to determine the extent of these variations, a program was initiated whereby commercial 302 and 304 stainless steels were drawn with a wide variation of speeds. Ultimate tensile strengths and specific resistances, as well as the Hc-BT measurements, were determined. If variations of considerable magnitude were found, it was thought that the knowledge accrued from the investigation could probably aid in determining the particular drawing speed required to obtain desired properties for a given specification for a particular analysis of wire. Drawing Procedure Two specimens (set I, No. 302 and No. 304), the chemical analyses of which are given in Table I, were drawn at different speeds, namely, 45 to 3900 ft per min, directly from 0.016 in. diam wire to 0.004 in. diam wire, a 93.7 pct reduction of area. To obtain this high percentage of reduction, the wire at 0.016 in. was softened by annealing at 1150°C. It was then strung in a No. 16CF variable speed Vaughn drawing unit and passed through 13 B.&S. gage dies of successively smaller diameter, as shown in the accompanying table. Following this step, several hundred feet of the wire was drawn, at a particular speed, through the dies and wound upon a spool. Once the sample of wire was drawn at the desired speed, the spool with the finished sample was removed, tagged for later identification, and an empty spool was placed in the machine. Another wire sample was drawn at a different speed, after making the necessary adjustments upon the Vaughn unit. At the greater speeds, several thousand feet of sample were drawn to insure that the wire to be used for testing would be uniform. This was done because the desired speed was not reached instantly; the wire taking a short while to attain the required drawing rate. All speed measurements were made by placing a tachometer against one of the machine's pulleys during the actual drawing of the specimens. 84.4 pct reduction of area. The 0.072 in. diameter wire was also annealed at 1150°C. The process of obtaining samples was the same as described above. The only difference between the two was that the rates involved were not as extensive, set II covering a range of 625 to 1150 ft per min. Tensile tests were carried out for the 0.004 in. diam wires on a horizontal Scott Tensile Tester, while tensile tests for the 0.028 in. diam wires were made on a vertical Scott Tensile Tester. Each value listed in the tables represented the average of five samples with a maximum deviation of approximately 3 pct among the samples. Resistance per foot of the wires was determined by use of standard bridges accurate to the nearest hundreth of an ohm for the 0.004 in. wire, and to the nearest thousandth of an ohm for the 0.028 in. wire. The resistances per foot were then converted to resistivities by use of the following equation: P = RxCM where R is the resistance per ft in ohms per ft; CM, the area in circular mils; and P, the resistivity in ohms—CM. The values listed in the tables likewise

Jan 1, 1952

By W. H. Smith

IN connection with some recent work on the effect of impurities on the ductility of chromium, it appeared desirable to know the solid solubility of carbon in chromium. A literature survey indicated that this information was not available. Although considerable work has been done on the Cr-C phase diagram,&apos;.&apos; previous investigators have been more concerned with the structure and phase boundaries of the carbide phases than with the terminal solid solution. The phase diagram shown in Fig. 1 is taken from the work of Bloom and Grant&apos; and represents the most recent determination. As indicated by the dashed line, the a solid-solubility limit was not determined. Experimental Procedure Alloys of chromium were prepared from hydrogen-treated and vacuum-degassed electrolytic chromium plus spectrographic grade carbon. The oxygen and nitrogen content of the alloys was <0.002 pct. After melting, analysis of the alloys showed them to contain 0.02, 0.08, 0.15, and 0.55 pct C. Pieces of the alloys were heated in a protective atmosphere to various temperatures and then quenched after holding for a time sufficient to insure equilibrium. Microscopic examination of the as-quenched alloys for the presence of a second phase was used as a measure of the solubility limit. The heats were made in a multiple hearth arc-furnace using a zirconium-gettered static argon atmosphere. A zirconium melt was made before each Cr-C heat. Triple melting was used to insure ingot homogeneity as shown by microscopic examination. The alloys were prepared by adding portions of a 4.5 pct C-Cr master alloy to high purity chromium. The carbon contents listed previously were those obtained by analysis. The nitrogen and oxygen contents after arc melting were both <0.002 pct. Sections 1/8X1/4X1/2 in. were cut from the 100-g ingots and a hole drilled in one end in order to suspend the sample from a molybdenum wire. After the surface was carefully cleaned, a sample of each melt was hung in a mullite tube heated externally by a platinum resistance furnace connected to a vacuum system. The lower portion of the mullite tube was sealed to Pyrex and closed off several inches below the furnace. This was filled with sili-cone oil kept cold by circulating cold water around the outside of the Pyrex. Quenching into the oil bath was achieved by melting a fuse wire supporting the sample. It required about 4 sec for the sample to cool from 1400" to 600°C. This severity of quench was considered satisfactory to freeze-in the high temperature equilibrium. For tests made at temperatures of 900" to 1200°C, heating was done in vacuum; for tests above 1200°C, an argon atmosphere was used. The holding time employed ranged from 12 hr at 900°C to 6 hr at 1400°C. Experiments were performed at temperatures of 900°, 1000°, 1100°, 1200°, 1300°, and 1400°C. Microscopic examination for evidence of a second phase was done at X1500. Experimental Result The microstructures of a 0.08 pct C-Cr alloy as-cast and after quenching from 1300°C are shown in Figs. 2 and 3. A 0.15 pct C-Cr alloy quenched from 1300°C is shown in Fig. 4. The data obtained from the quenching experiments is shown graphically in Fig. 5. If the Van&apos;t Hoff equation is obeyed, a plot on a logarithmic scale of the mol fraction of solute vs the reciprocal of the absolute temperature should give a straight line. For dilute solutions the weight percentage can be substituted for the mol fraction without introducing any appreciable error. The Van&apos;t Hoff equation can then be written as where H is the heat of solution in calories per mol. The slope of the straight line on the log pct C vs 1/T plot gives the value of AH. Assuming that the Van&apos;t Hoff equation is obeyed, which is probably justified for the dilute solution of carbon in chromium, the heavy straight line shown on Fig. 5 represents the best fit of the data. This line was obtained as follows. On Fig. 5 the results of the microscopic examination of all alloys following quenching were plotted and designated as to whether one or two phases were seen. Below 1100°C all alloys showed a second phase on quenching. The heavy vertical lines shown in Fig. 5 therefore represent the possible range of the ter-

Jan 1, 1958

By C. D. Thompson, D. C. Violetta, C. A. Czako

The mechanics of the continuous sintering process are briefly reviewed and the application of this process for the beneficiation of principal minerals used in making cement is shown. Pilot plant tests using pyrobeneficiation techniques were conducted with mixtures of limestone and shale in various ratios. Results are illustrated and show the percentage removal of carbon, carbonates, sulfur, alkalies, and volatiles as a function of the limestone-shale ratio. Tests with varying quantities of solid fwls, added to mixtures containing limestone and shale in a 4 to 1 ratio, were made to attain optimum elimination of the volatile constituents. Experimental results show that 90 pct of these constituents were removed during the calcination sintem&apos;ng process. The advantages of using the beneficiated material for making cement are discussed with regard to prepam&apos;ng, handling, and clinkering operations. A sintering process for treating raw materials or kiln waste products which contain high amounts of alkaline materials is described. Tests were performed with a high alkaline waste dust mixed in a 1 to 1 ratio with raw cement meal. The data from these tests and others using different halides for aiding alkali removal are evaluated by the alkali content of the cement clinker. Test results showed that the Dwight-Lloyd process could be applied with facility in beneficiating raw materials for cement making. Substantial quantities of deleterious volutile constituents were removed by calcining and sintering techniques. The continuous sintering process has been applied for a number of years as a means of beneficiating minerals for subsequent conversion to merchantable products. The process was developed and first used for nonferrous metallurgical applicatfons. However, it has risen to prominence for ferrous metallurgical applications where last year in the U.S. and Canada nearly 40 million tons of iron ore were treated by grate firing processes. Through the years of process development many investigators have realized the versatility of the sintering process and have directed experiments toward the use of it for applications with nonmetallic minerals. These have included calcining limestone, carbonizing coal, expanding shale, and burning cement raw meal. Today commercial sintering plants produce cement clinker from mixtures of limestone and shale, lightweight aggregate from shale or clay, and calcined products from phosphate shale. The mechanics of the downdraft sintering process have been presented by many authors. In general practice the materials to be agglomated are mixed as moistened fines with a solid fuel and this mixture is spread uniformly on the grates of the sintering machine. The charge is then momentarily directed under an ignition furnace. After ignition a continuous downdraft of air promotes combustion of the fuel and through heat transfer particles become partially fused into a coherent mass. This is cooled from continued draft induced after combustion. Many investigators have shown that distinct zones are formed during the process of sintering iron ores. At the midway point of the sintering process these are prominent and are, respectively, from the top of the bed to the bottom, comprised of a cooling zone of sintered material, a combustion zone of sinter formation, a calcination zone, a drying zone, and a zone of unreacted materials. Completion of sintering occurs near the discharge end of the machine where the firing zone approaches either the grates or grate dressing material called hearth layer. Here the peak draft temperature usually can be noted in the windboxes. The zonal pattern has been illustrated with the use of graphs which show thermal gradients as they progress through the bed of materials during sintering. Fig. 1 shows, in simplified form, thermal gradients existing when clinkering cement on a sintering machine. The thermal gradients illustrated are for periods of initiation, midway and completion of the clinkering process. The occurrence of these zones has importance for the beneficiation of cement raw materials. A research program was performed at the Dwight-

Jan 1, 1961

By F. O. Shuck, F. E. Butler

A diaphragm diffusion cell suitable for use with liquid metals was designed and tested by measuring the diffusion coefficients for a Hg-10 pct In alloy at 25°C. The binary diffusion coefficient and the coefficients for In and Hg tracers were measured at this composition. The measured coefficients for both tracers were less than the binary diffusion coefficient. When these coefficients were combined with the activity correction term of Darken, the predicted binary coefficient was somewhat higher than the experimental value. In recent years a number of investigators have been studying diffusion of liquid metals in order to gain a better understanding of the liquid state and to examine possible applications to liquid metal processing. The bulk of this work has involved either self-diffusion or diffusion of low concentrations of solute metals in an essentially pure solvent metal. The most common experimental technique employed has been the capillary-reservoir method of Anderson and saddington1 since it is well-suited to measurements involving low concentrations of solutes. As an expansion of previous studies, an investigation of the variation of diffusion coefficient over large composition ranges has been undertaken in this laboratory. Although it is possible to modify the capillary-reservoir method for such measurements, the technique loses its principal advantage of requiring only small quantities of solutes. Therefore, it was felt desirable to investigate other possible experimental techniques applicable to measurements over large composition ranges. A technique which has been used widely in measurements of other liquid systems and seemed potentially applicable to studies of liquid metals was the diaphragm cell. This method has been found to combine relative simplicity of operation with accurate and reproducible results. In use with normal liquids, the tyiical diaphragm cell utilizes a porous glass disk. However, because of the high surface tension, it is not possible to use a diaphragm of such fine pore structure with liquid metals. Furthermore. projected use at high temperatures makes the use of glass undesirable. Consequently, the first problem was the search for an appropriate cell design. EXPERLMENTAL Diaphragm Cell. A number of possible cell designs were examined. Prototypes of several were constructed and tested. The final design is shown in an exploded view in Fig. 1. The cell consists of two identical chambers, one on either side of a 0.20-in.-thick partition. The chambers are 0.15 in. deep and have a volume of 3.45 cu cm. Each chamber is provided with two small threaded holes for filling and emptying. The filling holes may be sealed with stainless-steel screws. Drilled lengthwise through the center partition is a 0.250-in.-diam hole. Inserted into this hole is a rod having seven 0.059-in.-diam holes drilled radially through it. The total "hole" volume is 0.078 cu cm. When the rod is properly aligned, the holes in it constitute the "diaphragm" through which diffusion occurs. The center section was of Teflon and the rod and face plates were of Plexiglas. Four such cells were constructed and used in this study. Subsequent to this investigation, a similar cell constructed entirely of iron was fabricated and tested at higher temperatures. Several considerations led to this final design. It is desirable to minimize the volume of the cell chambers to decrease the time required for a measurement. However, the total volume of metal in the diaphragm must be small in comparison with that in the chambers. To facilitate filling and emptying it is desirable to have a diaphragm which may be closed.

Jan 1, 1970

By L. D. Fetterolf

The New Jersey Zinc Zompany operates at Depue, Illinois, an integrated zinc smelting plant using tie vertical retort reduction process. The overall operation comprises green concentrate roasting, sintering, vertical retort smelting and metal refining. In the roasting operation, the fine concentrate is pelletized, the dried pellets fluidized and autogenously roasted with air in a “fluid column” with minimal formation of dust requiring recirculation. The roasted pellets and any associated roaster dust are agglomerated in a rotating drum and laid down, in combination with coal for fuel, on a Dwight-Lloyd sintering machine which converts to a semi-fused mass - a physical state more suitable for vertical retort smelting. The smeltins plant of 27 vertical retorts rated at 8 T zinc each per day on average requires the following major processing steps - mix preparation and briquettine, briquet coking, vertical retort reduction, vapor condensation and slab casting of metal. A portion of zinc production is converted to 99.99+ percent purity by boiling and fractional distillation in columns consisting of externally heated, superimposed trays of high-conductivity refractory.

Jan 1, 1970