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By John Fulton

Early in the year 1875, some difficulty was experienced in the " Old blast-furnaces " of the Cambria Iron Company, at Johnstown, Pennsylvania, arising from the increased use of native coke, prepared in Belgian ovens near the furnaces, in mixture with Connellsville coke, in the production of Bessemer pig-iron. This was the more surprising at that time, because a very caulplete coal-crushing and washing apparatus, after the Bradford plan, had just been put into successful operation, affording a pure coal for coke-making. It was confidently expected that this thorough cleansing of the coal would improve the quality of the coke and fit it for the increasing manufacture of Bessemer pig-iron. But the trouble with the native coke continued. Experiments in fuel charges-increasing and reducing them—failed to correct the irregular working of these furnaces. Increasing the charge of native coke produced abnormal heat at the top of the furnace, reducing it cooled the furnace below. At this stage of affairs, the Hon. Daniel J. Morrell, general manager of the Cambria Iron Company, requested the writer to make the requisite investigation to determine the nature of this coke-difficulty, and to ascertain why the Johnstown coke could not be used in blastfurnace work as well as Connellsville. This task at first did not appear to present any special difficulties, but subsequently it expanded into proportions that caused no little anxiety. The first effort was directed to procuring average samples of Johnstown and Connellsville coke. At this period of the investigation, it was assumed that the whole matter could be intelligently solved by comparing the chemical analyses of the two cokes: Johnstown Coke. Connellsville Coke. Moisture,.......... 0.30 Volatile matter,......... 0.46 Fixed carbon,.90.48 89.57 Ash,.8.96 9.11 Sulphur,.0.56 0.82 Phosphorus,..... 0.014 100.00 100.274 Analysts, T. T. Morrell. A. S. McCreath.

Jan 1, 1884

By G. R. St. Pierre, A. Ghosh

In the Si-C-0 system at temperatures of interest in process metallurgy there are four stable condensed phases—silicon, carbon, silica, and silicon carbide (Sic). Silicon melts at 1683oK1 and shows negligible solubility for carbon and oxygen at 1700°K. However, silicon dissolves approximately 1.5 at. pct C at 2500oK.2 Carbon in the form of graphite remains pure and dissolves negligible amount of silicon.' Silica (crysto-balite) melts at 2001°K' and is essentially a stoichio-metric compound. Silicon carbide is a solid and shows negligible deviation from stoichiometry up to approximately 2800oK.2 Besides the condensed phases there is a gas phase

Jan 1, 1970

By C. G. Black

The purpose of this paper is to present data obtained from an investigation conducted on the grindability of coal by the American Society for Testing Materials Tentative Standard ball-mill method. This investigation has been conducted over a period of a year, during which time several hundred tests were run on a wide selection of different coals coming from Pennsylvania, Ohio, West Virginia and Kentucky. The data herewith presented tend to show that the tentative standard ball-mill method might give results that are misleading and that a volume measure might be more indicative of the true grindability of materials of varying specific gravity. A means whereby the time element might be reduced has been studied and other phases of this work such as preparation of samples, mechanical screening, reporting results and duplicating tests have been investigated. It is hoped that all criticism offered in this paper will be taken as constructive, with the purpose ill mind of perfecting the method now tentative to show the true conditions 2nd of making optional the manner in which certain parts of the test might be conducted to meet diversified laboratory conditions. Weight and Volume The results of grindability on extraneous material from the coal scam obtained with the present tentative standard ball mill method indicate easier grinding of these materials as compared to most coals. Varying amounts of these impurities in the coal therefore may cause a raising of the grindability index and may lead to the erroneous conclusion that coals containing large amounts of these foreign materials are easier to grind. For a true comparison of grindability, it would seem necessary to grind to the same particle weight. If material were sized in a commercial mill by means of an air current set at a velocity to remove 1.60 sp. gr. float coal ground to 80 per cent minus 200 mesh, any 1.60 sink particles would be left behind, and in order to remove this 1.60 sink material by the same current it would be necessary to grind finer and produce a

Jan 1, 1936

By C. G. Black

The purpose of this paper is to present data obtained from an investigation conducted on the grindability of coal by the American Society for Testing Materials Tentative Standard ball-mill method. This investigation has been conducted over a period of a year, during which time several hundred tests were run on a wide selection of different coals coming from Pennsylvania, Ohio, West Virginia and Kentucky. The data herewith presented tend to show that the tentative standard ball-mill method might give results that are misleading and that a volume measure might be more indicative of the true grindability of materials of varying specific gravity. A means whereby the time element might be reduced has been studied and other phases of this work such as preparation of samples, mechanical screening, reporting results and duplicating tests have been investigated. It is hoped that all criticism offered in this paper will be taken as constructive, with the purpose ill mind of perfecting the method now tentative to show the true conditions 2nd of making optional the manner in which certain parts of the test might be conducted to meet diversified laboratory conditions. Weight and Volume The results of grindability on extraneous material from the coal scam obtained with the present tentative standard ball mill method indicate easier grinding of these materials as compared to most coals. Varying amounts of these impurities in the coal therefore may cause a raising of the grindability index and may lead to the erroneous conclusion that coals containing large amounts of these foreign materials are easier to grind. For a true comparison of grindability, it would seem necessary to grind to the same particle weight. If material were sized in a commercial mill by means of an air current set at a velocity to remove 1.60 sp. gr. float coal ground to 80 per cent minus 200 mesh, any 1.60 sink particles would be left behind, and in order to remove this 1.60 sink material by the same current it would be necessary to grind finer and produce a

Jan 1, 1936

By Andrew Mayer

Early in 1942 National Lead Co. was requested by the War Production Board to construct and operate a plant for the Government to produce magnesium by the ferrosilicon process which had been developed by Dr. E. M. Pidgeon in the laboratories of the Canadian National Research Council. A contract with the Defense PlantCorporation was concluded in May 1942 and Magnesium Reduction Co., a wholly owned subsidiary of National Lead Co., was formed to carry out this project. The capacity of the plant was rated at I0,000,000 lb. of magnesium per year, equivalent to an average production of about 14 tons per day. The Process The process, in brief, is as follows: The raw materials are dolomite and ferro-silicon, 75 per cent grade. The dolomite is calcined, the calcine and ferrosilicon are ground and mixed and the mixture is briquetted. The briquets are charged into tubular retorts of chrome-nickel steel set horizontally in a furnace with the open ends projecting outside the front wall. The retorts are then closed and evacuated. Magnesium is liberated according to then reaction 2(MgOl CaO) + Si = 2Mg + zCaO, Si% It is distilled from the charge and con- densed on a removable sleeve in the throat of the retort. Sodium and potassium, if present, are also liberated, distilled and condensed on a "sodium condenser" in the end of the retort. At the expiration of the distillation period the retorts are discharged and the cycle of operations is repeated. For the rated daily capacity of 14 tons of magnesium there are required approximately 170 tons of dolomite, equivalent to about 85 tons of calcine, and 16.5 tons of ferrosilicon. The retort residue, a bulky powder which at present is waste, amounts to about 87 tons. Choosing the Site The investigation to select the location of the plant was conducted by the Research Laboratories of National Lead Co. This work included geologic and economic surveys of a number of districts, examination of dolomite samples by chemical, spectrographic and petrographic methods and large-scale tests of dolomite from several of the more promising localities. in the last-named tests 2-ton samples were put through the ferrosilicon process in the pilot plant of the Canadian National Research Council, under the guidance of Dr. Pidgeon. The conclusions drawn from the investigation, in regard to the quality of dolomite to be used in the ferrosilicon process, were: First, the dolomite should contain not less than 21 per cent MgO and not more than 2 per cent acid-insoluble material. Second, the dolomite should

Jan 1, 1944

By Thomas D. Henderson, Donald E. Crowell

INTRODUCTION This paper presents a "case history" of the steps taken to estimate capital and operating costs for a typical porphyry copper concentrator of ±9,070 metric tons (10,000 short tons) per day capacity. The concentrator described herein is hypothetical, but is based on experience and is, in effect, a composite of several actual projects. The numbers used are illustrative only and do not represent actual costs at any particular concentrator. These numbers are, however, representative of typical milling costs at one point in time, and. are sufficiently accurate to illustrate the cost estimating procedures described in this paper. It is not possible to publish actual, current milling costs and, in any event, actual cost figures quickly become meaningless in today's economic situation with various costs all rising at different rates. The paper is divided into three parts: (1) order-of-magnitude estimate, (2) feasibility study estimate, and (3) discussion of differences between estimated and actual costs.

Jan 1, 1978

By M. J. Taffanel

I AM very much impressed by this manifestation of international brotherhood; the mining engineers on both sides of the ocean have similar subjects to deal with, meet with the same difficulties, expose themselves sometimes to similar dangers; they are like companions in arms in the battle fought by mankind in order to subjugate the natural forces and to draw profit therefrom. There is no other field where this fraternity must manifest itself more actively than in the studies pursued with a view to improving the safety of mines and to rendering impossible in the future these mining disasters, which every year claim hundreds of victims. These are very difficult problems, the solution of which requires numerous and laborious experiments accompanied by very arduous scientific investigations; it is extremely desirable that the various mining countries, which have undertaken to solve these difficult problems, should unite their efforts in order to succeed as soon as possible and as well as possible. This idea has been fully grasped by your Bureau of Mines, which has brought me to Pittsburg; the daily exchange of views with the distinguished chief of-the mine investigation division, George S. Rice, and with his excellent collaborators, the comparison of the results obtained at Lievin and Bruceton, the checking of these results by means of new tests, the profound study of the measurements, diagrams, measuring and laboratory apparatus, the cooperative investigation of new apparatus, or the preparation of new investigations, are, I hope, of such a nature as to bring about great progress. I think that it will be interesting to you to know what has been done in France along this line.

Jan 9, 1914

By T. B. Counselman

AFTER several years study, the Universal Atlas Cement Co. decided to rebuild its plant at Leeds, Alabama. The entire old plant, which was to operate during the new construction, was then to be scrapped with the exception of two kilns. The quarry was to be mechanized, new and larger kilns installed, with the most modern type of equipment all through the plant. Raw grinding operations at all of the Universal Atlas plants were dry. However, the dust hazard of this operation and the growing emphasis on occupational diseases made wet grinding attractive. Coal at Leeds is cheap, so that a slightly higher coal practice was unimportant. Power was cheap, so that waste-heat boilers were not desirable. Since new kilns were to be installed, these could be made long enough to make filters unnecessary. Wet raw grinding could be either the old practice of open-circuit compartment mills or closed-circuit grinding in accordance with metal-lurgical practice. The latter saved power and steel wear and saved initial investment. This had been demonstrated in cement plants. Also, the elimination of tramp oversize was a most important advantage. Ease of chemical control was another factor. Blending in the thickener, with elimination of slurry mixers had been found practicable. Therefore, if wet raw grinding were used at all, the advantages were clearly in favor of closed-circuit grinding. Detailed layouts of the most modern possible closed-circuit dry-grinding plant and of a closed-circuit wet-grinding plant were prepared, and careful estimates were made of initial invest-ment and operating costs. Full consideration of all the factors involved resulted in a decision to install wet closed-circuit grinding.

Jan 1, 1939

By Harrison Schmitt

DURING the past decade the writer has studied and mapped certain ore deposits and their structural associations in the states of Chihuahua, Durango, New Mexico and Arizona, and he believes that these deposits, with the exception of the manto-chimney, are representative of the principal types in the area of the states named. In this paper the deposits are described; then generalizations are drawn so far as seems appropriate. HANOVER, NEW MEXICO The Hanover; New Mexico, zinc deposits were briefly described by Lindgren and others,1 Paige2 and recently in more detail by Schmitt.3 These, the kind ordinarily called contact metamorphic, border the southern end of an oval area of granodiorite intrusive 3/4 mile wide and 2 1/2 miles long (Fig. 1). The granodiorite has intruded, folded, and metamorphosed sedimentary and igneous rocks, which range in age from pre-Cambrian to Upper Cretaceous. About 90 per cent of the ore has replaced coarse, recrystallized, pure, fossiliferous, lower Mississippian, nonmagnesian limestone (Hanover limestone) 110 ft. thick;4 about 75 per cent of it occurs in the upper 50 ft. of this horizon just below an 18-ft. lower Pennsylvanian shale bed, which lies on a locally inconspicuous unconformity representing upper Mississippian time. The orebodies show five shapes or attitudes, each dependent on a special structural situation: 1. Most of the ore occurs as podlike; thick lenslike or other "three-dimensional" shoots extending from top to bottom of the Hanover limestone and lying adjacent to the outer contact of a silicate zone largely of andradite, which follows the intrusive contact (Figs. 1 and 2).

Jan 1, 1933

By T. B. Counselman

AFTER several years study, the Universal Atlas Cement Co. decided to rebuild its plant at Leeds, Alabama. The entire old plant, which was to operate during the new construction, was then to be scrapped with the exception of two kilns. The quarry was to be mechanized, new and larger kilns installed, with the most modern type of equipment all through the plant. Raw grinding operations at all of the Universal Atlas plants were dry. However, the dust hazard of this operation and the growing emphasis on occupational diseases made wet grinding attractive. Coal at Leeds is cheap, so that a slightly higher coal practice was unimportant. Power was cheap, so that waste-heat boilers were not desirable. Since new kilns were to be installed, these could be made long enough to make filters unnecessary. Wet raw grinding could be either the old practice of open-circuit compartment mills or closed-circuit grinding in accordance with metal-lurgical practice. The latter saved power and steel wear and saved initial investment. This had been demonstrated in cement plants. Also, the elimination of tramp oversize was a most important advantage. Ease of chemical control was another factor. Blending in the thickener, with elimination of slurry mixers had been found practicable. Therefore, if wet raw grinding were used at all, the advantages were clearly in favor of closed-circuit grinding. Detailed layouts of the most modern possible closed-circuit dry-grinding plant and of a closed-circuit wet-grinding plant were prepared, and careful estimates were made of initial invest-ment and operating costs. Full consideration of all the factors involved resulted in a decision to install wet closed-circuit grinding.

Jan 1, 1939

By R. H. King

This paper describes the results of an underground coal mine trial installation of a shielded trailing cable. The cable was a three conductor, round, type SHC-GC and was installed on a shuttle car in a conventional room-and-pillar mining section. The purpose of the study was to evaluate proposed legislation that would require shielding on low-voltage trailing cables. Such legislation would require machine modifications to cable reels, packing glands, and couplers; increase cable splicing complexity; and necessitate additional splicing materials and training. Furthermore, fault location might be more difficult. Even though shielding will cause some problems for the mine operator, it can reduce electrical shock hazard to mine workers. This paper presents information that mine operators can use to solve most of the problems encountered if they are required to install shielded cables. The in-mine evaluation was funded by the US Bureau of Mines; however, the views and conclusions in this paper are those of the author and should not be interpreted as necessarily representing official policies of the US government.

Jan 1, 1982

By S. P. Goplen, M. Henriksen, T. F. Barron

In the past decade, numerous large scale drilling projects have made use of reverse circulation air lift methods. Reverse air lift pumping can be established by injecting gas into a column of liquid; the subsequently decreased density of the gas-liquid mixture allows atmospheric pressure outside the lift conduit to establish flow. The rate of flow can be controlled by varying the gas to liquid ratio. A typical reverse circulation air lift system is depicted in [Fig. 1]. Allen 1 has described a number of reverse air lift drilling projects. Although cited examples are generally related to large diameter holes, the method has seen use in other applications such as sand dredging, chemical pumping, water well pumping. and continuous coring. One particular benefit obtained with enclosed reverse circulation drilling is elimination of lost circulation. The required air/water ratio for pumping at a given submergence and lift has been established experimentally by Gibbs.2 His data, however, do not reflect sensitivity of flow rate to air/water ratio. Further, in Gibbs' work, air apparently is injec-

Jan 1, 1980

By R. G. Mallon, R. L. Braun

Primary copper-sulfide ore deposits that are well below the water table can be chemically mined by in-situ high-pressure leaching. The leaching is accomplished by pumping oxygen gas into the bottom of a large underground flooded column of rubblized ore. Part of the injected oxygen dissolves under the high hydrostatic pressure in the column. With a high enough injection rate, the excess oxygen gas induces a convective circulation of oxygenated solution within the rubble column, which transports dissolved oxygen throughout the system. At elevated temperatures, this results in a relatively rapid dissolution of the copper minerals. A method of calculating the leaching rates is presented. The calculation is done by means of a finite-difference computer code that couples all of the processes relating to copper leaching, convective circulation, and the injection, dissolution, and consumption of oxygen. Results of the calculation for a typical case show that over 70% of the copper can be extracted in less than six years with good efficiency in oxygen utilization.

Jan 1, 1976

By J. M. Snook, F. C. Langenberg

FOR a better understanding of the melting and refining processes, the oxygen content of the steel bath should be determined. The oxygen content influences many process and product variables such as the efficiency of alloy recoveries, the ingot structure (through reaction with carbon in the steel), certain reaction rates, cleanliness (formation of indigenous oxide inclusions), and mechanical properties. Many methods of sampling liquid steel for oxygen have been devised. These methods can be divided into two main groups: 1) taking samples directly in the furnace or ladle;'-'0 2) removing the metal from the furnace or ladle in a sampling spoon and pouring it into a mold, or drawing the metal from the spoon into a sample holder.11-l4 The advantages and disadvantages of the two general sampling methods are presented in Table I. The bomb-sampling technique described by Huff, Bailey, and Richards8 has been used by the Crucible Steel Co. in certain research projects where oxygen contents were needed. It was recognized that a localized boiling condition could result when the mold is introduced into the steel bath, although the work of Gilbert and Bailey9 seems to indicate that the carbon boil occurring in the bomb-sampler cavity has a negligible effect on the oxygen content. In order to confirm the reported negligible effect of a carbon boil, additional work was done with adaptations of the conventional bomb sampler. The methods used are shown in Fig. 1 and are as follows: 1) Type A: aluminum wire (0.125-in. diam) coil in cavity; steel cap over mouth 2) Type B: aluminum cup in cavity; aluminum cap over mouth, 3) Type C: aluminum cup in cavity; steel cap over mouth, 4) Type D: fine aluminum wire (0.013-in. diam) in cavity; steel cap over mouth. The amount of aluminum wire in the Type A and D samples varied from 0.8 to 1.2 pct of the sample weight. The aluminum cups averaged 2.0 pct of the sample weight; the latter were used to ensure that the steel contacting the walls of the bomb cavity was fully killed. Samples were taken from plain carbon heats made in open-hearth furnaces. This grade of steel before silicon additions is essentially an Fe-C alloy and the effects of other alloying elements on the oxygen activity can be neglected. The results of sampling with the bomb samplers is shown in Fig. 2 where the percent oxygen is plotted against the percent carbon on log-log paper. Spoon sample results on the same grade of steel where the sample was killed in the spoon are shown in Fig. 3. The oxygen contents were determined by the vacuum fusion method. A comparison of Figs. 2 and 3 shows that the oxygen concentrations obtained with the spoon-sampling method (metal killed in the spoon after slag removed

Jan 1, 1960

By P. R. Swann, R. M. Fisher, W. R. Duff

Observations by transmission electron microscoPy of the microstructures in quenched Fe-A1 alloys provide the basis for revision of the Fe-A1 phase diagram. The diagram is characterized by first-order transitions between the disordered ferromagnetic a! solid solution and the two ordered phases FeAl and Fe3Al. However, second-order transitions showing corre-luted fluctuations in the degree of order near the critical temperature occur between the paramagnetic a! and FeAl and the FeAl and FeAl phases. ELUCIDATION of the true nature of the Fe-A1 equilibrium diagram has proven to be an extremely challenging problem as evidenced by the fact that six quite different versions of the iron-rich portions have been published since 1958.&apos;-6 The most recent diagram shown in Fig. 1 was proposed by Rimlinger6 based on X-ray diffraction analysis and dilatometry. He shows the single-phase fields of the ordered phases of the types FeAl(L20) and Fe3A1(D03) and the disordered bcc Fe-A1 solid solution, a, separated from one another in every case by large two-phase regions. Thus, he implies that in Fe-A1 alloys ordering always occurs via a classical phase transition. In contrast to this view, theoretical studies3&apos;" have concluded that the transitions a — FeAl and FeAl — Fe3A1 should be second- or higher-order in character. A further revision of the Fe-A1 diagram of state is presented in Fig. 2 based on our previous8,9 and current observations by transmission electron microscopy of the microstructures found in quenched specimens. The average quenching rate was 50,000°C per sec and although this was sufficiently rapid to prevent Fe3A1-type ordering below -550°C it was not sufficient to suppress ordering of the type FeAl in the temperature range -650" to 800°C. However, ordering during the quench is easily recognized since the FeAl phase formed during holding at temperature rapidly develops large antiphase domains whereas FeAl formed during quenching has a very fine domain size, -50Å diam. Each symbol in Fig. 2 indicates the result of a micro-structural examination of specimens heat-treated (-1 hr) until equilibrium was established. The results will be discussed in the order of the points, A - Dl marked on both diagrams with reference to the phase fields, critical temperature, and "transition" ranges, respectively. The dark-field micrograph corresponding to point A is shown in Fig. 3 and was obtained using a 111 reflection of the Fe3A1 structure which reveals both a&apos;/4<111> and a&apos;/2<100> antiphase domain boundaries.10 The structure observed in this specimen is typical of single-phase Fe3A1 and shows no evidence for FeAl or any other additional phase. The dark-field micrograph shown in Fig. 4 corresponds to point B on the two diagrams. It was taken with a 100 FeAl reflection which reveals a/2(111) domains and exhibits the typical structure of single-

Jan 1, 1970

By Donald R. Mason, Daniel F. O’Kane

DIFFERENTIAL thermal analysis measurements have been used to determine the ZnTe-In2Te3 pseudo-binary phase diagram. Since ZnTe and In2Te3 are both semiconductors, any intermediate compounds formed in this system should also be semiconductors according to the rules delineated by Mooser and pearson.1 The compound ZnIn2Te4 was first reported by Hahn et a1.2 and it has subsequently been prepared and characterized electrically by several workers. A brief summary of this work has been presented by Mason and O&apos;Kane3 who also report that the compound appeared to melt peritec-tically. The Zn-Te phase diagram has been investigated by Kulwicki4 who found that ZnTe has a melting point of 1290° + 2°C. The In-Te diagram published in Hansen&apos; contains significant errors, and has been redetermined by Grochowski et al.6 However, both sources agree that the melting point of In2Te3 is 667°C, with the latter authors reporting the previously observed solid-solid transformation occurring at about 550°C. Hahn ef al.2 reported lattice-parameter values for the ZnTe-In2Te3 system. The 67 mole pct ZnTe-33 mole pct In2Te3 composition was the only

Jan 1, 1965

By J. S. Koehler, F. Seitz

THE dislocation theory of plastic flow originated in 1928, when Prandtl1 suggested that the internal damping of metals of plastic origin might arise from the motion of a crystalline imperfection. Following this, Dehlinger2 employed a similar imperfection in a theory of recrystallization. The first detailed theoretical discussions were those of E. Orowan3 and G. I. Taylor4 in 1934. The resulting theory was gen- eralized by J. M. Burgers5 in 1939. Mott and Nabarro, in England, and Seitz and Read,6 in the United States, started an extensive examination of the atomic theory of dislocations in the period between 1939 and 1941. However, this development was arrested by the mar and was not taken up again until after 1945. Since 1946, the field has been the object of very intense activity, particularly in the United States and England. Dislocations are beautiful and intricate imperfections; they are sufficiently complicated to account for the great variety of phenomena encountered in the field of plastic flow. Above all, the dislocation concept enables us to understand how a metal or other plastic solid can deform inelastically under very small stresses at temperatures far below the melting point. Suppose we attempt to shear a perfect crystal plastically. This can be done by sliding one entire atomic plane simultaneously through one atomic distance with&apos; respect to the near-

Jan 1, 1954

By Leon Y. Sadler, E. K. Landis

Froth flotation has been described by several authors" as being analogous to a first-order rate process. Although a few investigators"," have found orders other than one fit their data best, the rate is described most often by the first-order equation, [ ] where r is rate of disappearance of floatable mineral by flotation, C(t) is concentration of floatable mineral in the cell, t is time, and k is first-order flotation rate constant. One of the difficulties in using this rate equation to develop a mathematical description of an actual flotation process is the determination of a suitable value of the constant k. The value of k, which reflects the floatability of the floatable mineral, is dependent on the hydrodynamic conditions in the cell, type and concentration of flotation reagents used, pulp density, and pH, method of froth removal, aeration rate, particle characteristics, and many other factors. Therefore, it is reasonable to expect that each particle of floatable material has a distinct value of k and that a unique distribution of k values exists for given feed and cell conditions. This has been observed and discussed by several investigators, and mathematical descriptions of the batch cell and steady-state continuous cell with distributed flotation rate constant have been developed." The mathematical description in the form of a transfer function of the continuous or flow cell under unsteady-state conditions with a distributed flotation rate constant has not been presented in the literature. The purpose of this note is to make such a presentation. The basic features of the perfectly mixed continuous

Jan 1, 1974

ROBERT M. KEENEY,* Portsmouth, Ohio (written discussion?).- Mr. Coffin has written a most interesting paper that covers most of the uses to which water-cooled devices are subjected; but to meet with the operating department&apos;s approval, these devices must effect a saving in cost per ton of ingots produced in the furnace. The making of steel in the open-hearth furnace is dependent entirely on the personal element. It is not possible to charge the furnace with so much raw material, turn on a valve for heat, and by use of recording thermometers, clocks, etc. empty the furnace when a steel is made. Competent men trained to the work must be in charge all the time, and the quality and quantity of steel produced is quite as dependent on the care and working qualities of the furnaces as on the raw material and working of the charge itself. Steel making is hot, hard work and the inducement offered to technical men to take up this work is not sufficient to induce many to enter the field. The result is that most of the workmen in an open-hearth plant, the men who actually operate the furnace and make the steel, are not technical men and do not understand the "whys" and "wherefores " and care less. All this means that a furnace must be built as nearly fool-proof as possible. By this I mean that it is necessary to protect those parts of a furnace most subject to abuse by water-cooled devices whenever possible, and these water-cooled appliances must he as nearly fool-proof as they can be made. The control of the gas entering a furnace is of the utmost importance. This means a properly designed port must have the same design throughout the whole run of the furnace. This is not possible with the ordinary

Jan 6, 1919

By E. P: Mathewson

Discussion of the paper of E. P: Mathewson, presented at the Butte Meeting, August, 1913, and printed in Bulletin No. 78, June, 1913, pp. 1033 to 1037. PROF. JOSEPH W. RICHARDS, South Bethlehem, Pa. :-Speaking to the question of the development of the basic lining, and speaking from a somewhat considerable study of the history of the process, ,I think Mr. Mathewson&apos;s conclusion (in the last paragraph. of his paper) is perfectly justified and is very fairly stated. Messrs. Smith and Pierce worked at developing their process for many years ; copper men had been trying their level best to get away from the silica lining of the converters and had uniformly failed. On one page of Mr. Mathewson&apos;s paper we read that at the old works of the Anaconda Company, at Anaconda, the experiments " were abandoned on the score of cost and lack of advantages." On the next page, after describing the work of Messrs. Smith and Pierce, we have the statement that the Anaconda Company lined a shell with mangnesite brick and the ?results were excellent." I think Mr. Mathewson&apos;s statement that to Smith and Pierce belongs the credit of developing a long-discredited idea, is correct. C. D,. DEMOND, Anaconda, Mont. :-It seems to me that the matter of the lining is important; my impression is that they used tar in their basic lining and that did not seem to be at all successful at Anaconda, but silicate of soda was very successful. MR. RICHARDS :-I did not know that they- had abandoned the use of tar. MR. DEMOND :-I said I thought that was the practice. At Anaconda that was not at all satisfactory in the basic linings. Silicate of soda was very satisfactory and is so.

Jan 11, 1913