Search Documents

By William N. Matheson

The major consideration in any management and engineering decision is the economic evaluation of factors. The right decision is the best cost decision, provided of course that it is the safest. Changes, improvements, and advances in mining technology, as in any business, come about through the pressure of economics in combination with the inventiveness and engineering abilities of efficient mine operators. Without proper economic evaluation and an appreciation of profit and loss of an operation, advances in mining methods would come about at a slow pace, if at all.

Jan 4, 1959

Still, the subject of iron ore is associated in peoples' minds mostly with the Lake Superior region and this is as it should be. The Minnesota Section meeting exposed the forces that over a period of two decades, 1950 to 1970, are revolutionizing the industry. The Lake Superior region has been at the center of these changes and has been deeply affected by them. The rise of foreign sources of iron ore that did not require beneficiation and taxation considerations were for a number of years a giant setback for the Lake Superior region and Minnesota in particular. However, the companies in this region have led the drive to improve the quality of ores which are culminating in the tailored pellet. This is a product in which iron content, silica, without significant impurities, is manufactured for optimum blast furnace performance. The oxide pellet does not seem to be the end as partially or prereduced pellets are closer at hand.

Jan 3, 1966

By George Rice

THE A.I.M.E. Ground Movement and Subsidence Committee, pro-posed in 1920, held its first technical meeting in February 1923, under the able chairmanship of Mr. H. G. Moulton. The following list of papers and discussions, indicating the scope of the problems and relation to various conditions and method of mining, were given at that meeting by the distinguished mining engineers: J. Parke Channing, on Subsidence at Miami, Arizona; Howard N. Eavenson, on Mining an Upper Bitu-minous Seam after a Lower Seam Has Been Extracted; J. J. Rutledge, on Subsidence in Two Oklahoma Coal Mines; Benjamin F. Tillson, on Caving Cracks at the Franklin Mine, New Jersey; Dr. F. W. McNair, on the Dome Theory; William Kelly, on disadvantageous effects in Mining and Back-filling Working Upward a Steep-Pitching Iron-ore Bed; Eli T. Conner, on the good results of Back-Filling by Hydraulic Means in Preventing Destructive Subsidence in the Anthracite District; Thomas H. Claggett, on the merit of Systematic Pillar Drawing in Preventing Cracks in an Upper Seam; Henry H. Otto, successful results in the anthracite district in Mining Lower Seams before Upper Seams; Douglas Bunting, on Leading Factors Incident to Successful Mining of an Overlying Seam; Edwin Ludlow, on there being "less damage to buildings when there was complete removal of coal with complete packing than when room and pillar is used without extracting the pillars"; Edward O'Toole on "bumps" in eastern Kentucky mines, which he thought were the result of geologic stresses: H. A. Buehler, on Upheaval of Surface in an Illinois Case in Advance of the Face; and R. Dawson Hall, on The Mechanics of Subsidence from Mining. The author of this paper discussed various kinds of problems of ground movement and subsidence in both metal mining and coal mining; the production of great landslides by workings near an outcrop, as at Turtle Mountain, Alberta, Canada; and general

Jan 1, 1939

By J. B. Cheatham

An analysis of indexing is given for an idealized plastic rock. Indexing refers to chipping from a bit tooth to a previously formed crater. A method is indicated for computing the dependence of chip size and shape as well as force on the tooth for various tooth shapes as a function of crater configuration and distance to the previous crater. Pressure on the tooth is found to vary with depth of penetration after a critical depth has been reached. Damp sand model experiments indicate that simple solutions using linear and logarithmic spiral slip-line fields for Coulomb plastic materials give reasonable approximations to the failure patterns. Both idealized roller cone bits and fixed blade or drag bits are considered, and methods are presented for approximating the force on the bit and the chip volume when the rock is subjected to sufficiently high pressure to behave plastically. One of the primary objectives of drilling research investigations is to develop theoretical equations which can be used to predict the behavior of rock under drilling loading conditions. Since rocks usually fail in a brittle manner at atmospheric conditions, much current research is directed to studies of the brittle failure of rock under the action of drilling too1s. 1,2 The results of triaxial test on rocks indicate that both rock strength and ductility increase at high confining pressures.3 Recent results from single bit-tooth indentation experiments on dry rocks indicate that the transition from brittle to ductile failure under a bit tooth can occur at pressures sufficiently low to be of interest in oil well drilling.4 Single bit tooth experiments and analyses are not directly applicable to the drilling process since one of the important facets in drilling is the interaction between a bit-tooth and the previously formed crater. This interaction is referred to as indexing.' The purpose of the present paper is to indicate how plasticity theory can be applied to the analysis of indexing for plastic rock. Drag bit or fixed-blade drilling is considered and a discussion is given of the drilling action of an idealized roller cone bit. Results of simple experimental studies using a damp sand model are given as a method of visualizing the failure patterns in rocks under high confining pressures. These results are compared with plasticity theory using simple solutions based on linear and logarithmic spiral slip-lines. Discontinuous stress fields are proposed for drag bit drilling based upon the soil mechanics analysis of Shields. SINGLE BIT TOOTH RESULTS Single bit tooth indentation experiments using dry rocks under pressures from 0 to 5000 psi have indicated a correlation between the shape of a force-displacement curve and the nature of the rock failure beneath the tool.4 Continuous linear curves are in-

Jan 1, 1965

By H. B. Menardi

The history, geology, ore deposits and current mining operations of the Huitzuco district have been described by C. W. Vaupelll and the current mill operations by David Segura.2 This paper completes the story of Huitzuco with a description of the reduction of the flotation concentrates. The predominant ore mineral in the Huitzuco deposit is livingstonite, a rare, mercury-antimony sulphide (HgS.2Sb2S3). Other associated minerals include cinnabar, stibnite and elemental sulphur. The gangue is largely limestone and gypsum. Former operators mined a selected grade of ore, from which the mercury was recovered by an expensive, inefficient operation. A considerable tonnage of rejects and fills, carrying mercury and antimony, was left in the mine. In order to recover both mercury and antimony from these fills and also from newly developed ore, a combination method has been developed, by means of which the livingstonite is concentrated by flotation and the flotation concentrates subsequently reduced to metallic mercury and antimony. An essential part of the flotation treatment, as developed by Segura, is the use of copper sulphate as an activating agent. A typical analysis of concentrates now being produced at the property is as follows: Hg, 7.73 per cent; Sb, 26.40; Fe, 1.20; S, 29.68; CaO, 11.50; MgO, 7.25; SiOz, 1.63; A1203, 0.65. PRocess foR Treatment op Concentrates In the development of a process for treatment of these concentrates, three different methods were investigated: (1) roasting in a Dwight-Lloyd sintering machine with an excess of air; (2) separation of mercury and antimony by distillation in a reducing atmosphere; (3) a third method, which ultimately developed into the present process. The third method involved a slow calcination at a closely regulated temperature with a limited admission of air, which separated the mercury and antimony, the raeroury passing out of the kiln with the furnace gases and the antimony remaining in the calcine or "sinter."

Jan 1, 1944

By Stephen Birch

THE Chitina mining district of Alaska is located at the headwaters of the Chitina and Copper Rivers. At present, the only producing min-ing properties are the mines of the Kennecott Copper Corpn. and the Mother Lode Coalition Co., which are situated 196 miles from Cordova the port of entry. The first claims, later acquired by the Kennecott Mines Co. and afterwards transferred to the Kennecott Copper Corpn., were discovered in 1900. The Copper River & Northwestern Ry., which connects the mines with tide water at Cordova, was completed in the spring of 1911. Contemporary with the construction of the railroad, aerial tram equipment was brought to the mines by pack train and a tramway, 3 miles, long, connecting Bonanza mine with the proposed railroad terminal, was finished, enabling shipments of high-grade ore to be made immediately on the completion of the railroad. A mill to treat the lower grade ore was begun the same year. The Kennecott company's holdings consist of 111 mineral claims. The Mother Lode Coalition Mines Co., which is controlled by the Kennecott Copper Corpn., owns 73 claims adjoining the Kennecott holdings. All data on operations and geology refer equally well to the Mother Lode property.

Jan 1, 1924

By M. E. Wadsworth, R. D. Groves, B. W. Madsen

Experimental results obtained from large-scale (7 ton), long-term (500 days) leaching of low-tenor copper sulfide ores were interpreted by a geometric model involving the movement of a reaction zone through the ore fragments. The postulated leaching mechanism involved mixed kinetics and included diffusion of the reacting species through the host rock and a chemical reaction at the surface of the mineral particles. Ore properties which were considered were Ore particle size and shape, grade of ore, and mineral content. Excellent agreement between the mixed kinetic model and the laboratory leaching results was obtained when -1/2-in. ore particles were not included in the leach. Also, a satisfactory prediction of copper recovery from +1/2-in. ore with a different particle size distribution was made; however, it was necessary to modify the shape factor in the model. Prediction of copper recovery from an ore sample which included the -1/2-in. fine fraction and leaching for 460 days was 56.5%, and the recovery from leaching was 47.0%. The presence of fines caused solution channeling and a deficiency of oxygen which inhibited the leaching rate. Also, compaction of fines around coarser particles interfered with circulation around the ore particles, and bulk solution transport probably was a rate controlling factor.

Jan 1, 1976

By E. H. Mohr

At the McGill concentrator of the Nevada Consolidated Copper Co., all flotation operations have been carried out in Forrester cells since November, 1926. In respect to cost of operation, the new cell has proved to be superior to the Callow, which it supplanted, and its incorporation in the flow sheet has resulted in a marked improvement in the metallurgy. Before citing the advantages of the new machine over the old, it should be explained that the Forrester has not only replaced the Callow, but it has made possible the elimination of a number of tables as well. When Callow cells were in use, table concentration preceded flotation. The ball mill pulp was elevated, by means of air lifts, to roughing Wilfley tables situated two floors directly above the flotation floor. On an intermediate floor were additional tables which cleaned the concentrate from those above. The finished table concentrate constituted approximately 60 per cent. of the total mill production, and contained about 40 per cent, of the recovered copper. The table tailings were returned to the classifiers, the overflow from which went to the flotation cells and the oversize to the ball mills. As soon as the Forrester cells were installed, this tabling operation was discontinued. The ball mill pulp now discharges into the classifiers direct, without the intermediate tabling. It was impracticable to ernploy the present flow sheet when Callow cells were in use, for the reason that unless the great,er part of the heavy mineral was first removed, it accumulated on the mats and adversely affected the flotation operation. By tabling ahead of flotation, it was practicable to maintain the concentrator at 10 to 20 per cent. greater capacity (depending up on the pyrite contnt of the ore) than was possible by flotation alone. This was due to the necessity for finer grinding in the latter case. Even though the table conecntrates contained only 6 to 8 per cent'. copper content, the low smelting costs made it advantageous to operate at high mill capacity, at the cost of low grade of product. Standard Section equipment With the old flow sheet, a standard section consisted of four ball mills and classifiers, one air. lift, 20 roughing tables, 10 cleaner tables, 12 primary

Jan 1, 1928

By R. F. Mehl, C. S. Barrett, A. G. Guy

INTRODUCTION IN the last few years this laboratory has published a series of papers on the mecha¬nism of age -hardening.14,1,6,11,20 Briefly stated it has been proposed that hardening is caused by lattice strains created by lat¬tice coherency between matrix and precipi¬tate; much experimental evidence has been advanced to support this thesis. The system Cu-Be is an attractive one to study. It is a system that exhibits marked age hardening; it ought to be subject to the type of analysis of structural changes hitherto explored and it is a prominent example of "discontinuous precipitation," which, though studied in other cases, 6,20 remains largely unexplained. This research shows that hardening in the Cu-Be system proceeds essentially as proposed earlier.14 Guinier-Preston zones have been observed to form on the {100) planes of the matrix; no transition lattice was found. Aging and overaging were ob¬served to occur most rapidly in the region of discontinuous precipitation-the grain boundary-apparently accelerated there by strain. The Constitution of the Cu-Be System The Cu-Be constitution diagram is shown in Fig I. The structure of the ? phase was determined by Misch16 as of the caesium-chloride type with a lattice con¬stant of 2.598 Å High temperature X ray photograms of 7.2 pct Be alloy made by Kossolapow and Trapesnikow13 led them to conclude that the structure of the P phase is simply disordered body-centered cubic with [ao] = 2.79 Å at 750°C. Previous Work The initial work on the age-hardening of Cu-Be alloys was that of Masing, Dahl, Holm, and Haase. Their numerous papers were published in book form in 1929, sub¬sequently translated.18 The effects of Be content, solution temperature, aging tem¬perature, and cold work on the hardness developed in these alloys were studied. An initial decrease in the electrical conduc¬tivity was observed when strips of a 2.5 pct Be alloy were aged at temperatures from 200 to 350°C; a wire of the same analysis failed to exhibit this decrease when aged at 350°C. A volume decrease of 0.6 pct oc¬curred during aging at 250°C. Microscopic investigations of quenched and aged alloys disclosed a regularly ori¬ented "striping" (ripples) on the a and on the ß grains. In the case of the ß the cause of the striping was held to be decomposition

Jan 1, 1948

OFFICERS AND DIRECTORS (The Officers are the President, 6 Vice-Presidents, the Treasurer, and the Secretary) President and Director HENRY DEWITT SMITH' Past President and Director LEO F. REINARTZ President-elect and Director C. E REISTLE, JR

Jan 1, 1955

By M. J. Buerger

POSSIBLY the most puzzling features observed during a single-crystal deformation test are the appearance of slip striae on the surface of the crystal and the strengthening of the specimen.1 While the first phenomenon is mainly of theoretical interest, the latter is fundamental to all strain-hardening and cold-working processes in which the deformation occurs by translation. The plastic deformation of crystals has received considerable discussion2 and need not be reviewed here in any great detail. It has been observed, essentially, that after a certain amount of initial stress is applied, the surface of the crystal becomes covered with sets of parallel striae or rings which appear to divide it into thin slabs or slip blocks.3 The striae are the traces of rational crystallographic planes along which shearing movements are localized, these movements providing the strain which accompanies stress applied beyond pure elasticity. As deformation under the applied stress proceeds the slip blocks shear past one another, although the essentially single-crystal character of the specimen is unimpaired. All of the shearing apparently takes place in the immediate vicinity of the original slip striae without the appearance of new planes of movement.4 Since each increment of strain is attended by an increased stress requirement to continue the deformation, the crystal strengthens during the process.

Jan 1, 1928

By J. W. Axelson

CRUSHING has always been a major operation in the chemical and metallurgical industries, yet little is known about the theory of crushing, and today, the design of crushers is still based almost entirely on empirical knowledge and accumulated practical experience. In view of the increasing national need for the economic working of poorer grades of ores, the lack of a fundamental under- standing of this unit operation hardly presents a satisfactory situation. Basic investigations of crushing have been concerned mainly with three phases of the problem, 1-the mechanism of the fracture process itself, 2- the particle size distribution of the crushed product, and 3-the relationship between the energy input and the amount of new surface produced. Probably adequate information on these phases will be required for a comprehensive understanding of the process of crushing. A bibliography covering these phases will be given as well as a review of the work done in recent years at the University of Minnesota on the relationship between the energy input for crushing and the amount of new surface produced.

Jan 12, 1951

By Charles B. Dudley

It is evident that, as the size and weight of cars have increased, the demands on the cast-iron car-wheel have become more and more severe. Fortunately, the factor of safety in the cast-iron wheel, as originally made, was so large that it was only when freight-cars approached the capacity of from 80,000 to 100,000 lb. that the question of wheels began to give some anxiety. A moment's thought will show that, as freight- and passenger-cars have increased in weight and size, a much greater duty is required of the same eight wheels than was formerly the case. In other words, although the cars have increased in capacity from 15 tons to 50 tons each, the number of wheels under a car has not been generally increased. It is true, some cars have been equipped with 12 wheels, but this construction is not the rule. Hence the duty which the wheel must perform has become much more severe, and although it was clearly able to stand some increase, there must evidently be a limit somewhere which cannot be exceeded. Meanwhile, it is the part of wisdom to make better wheels if possible, and accordingly, no little study is now being devoted to the requirements for cast-iron car-wheels; and modifications are being made, both by increasing the weight of the wheel and by improving the quality of the metal of which it is made. The desire to improve the quality of the wheel has led to the preparation of specifications for its manufacture; and within the last 15 years a more rigid scrutiny of the wheel is char-

Jan 1, 1905

By B. H. Caudle, B. M. Hickman, I. H. Silberberg

Secondary recovery projects often are not started in oil reservoirs until dictated by rising GOR's or declining oil production. Such circumstances require a well dispersed injection pattern to prevent the oil bank from bypassing the production wells. Many times this purpose is served by using the five-spot or line-drive patterns, in which half of the wells are made injection wells. A regular injection pattern is proposed that is based on the same well spacing as the five-spot but in which only a third of the wells need be converted to injection wells. Fluid flow model studies on this proposed pattern show that production histories will be as good as those for a five-spot pattern at mobility ratios of three and above. The model data are presented in a form that may be used to predict recoveries from both homogeneous and stratified formations. Introduction The injection of fluid into a petroleum reservoir usually is prompted by an actual, or predicted, loss in the productivity of the reservoir. Often, an oil reservoir is produced by pressure depletion until high GOR's dictate a change in policy. Fluid injection also causes increased ultimate recovery, but this usually is of less importance to the operator than is the maintenance of the production rate. It is a fact that the ultimate recovery in any particular fluid injection project is relatively independent of the number and spacing of the injection wells as long as the reservoir voidage through the production wells is balanced by the total injection rate. However, the response of a production well is directly affected by the distance it lies from an injection well. In places where the well productivities are declining, the injection wells should be spaced as evenly as possible among the producing wells ("dispersed" injection pattern). More particularly, if the production response depends upon the formation (and subsequent production) of an oil bank, each production well should be surrounded by injection wells. Otherwise, the oil productivity will increase very little as the oil bank moves past the well.'.' An evenly spaced injection pattern has the best chance of generating an oil kick in the production wells. If reservoir voidage and injection rates are to be equal, the ratio of injection wells to production wells must be equal to the ratio of the average injectivity to the average productivity (in reservoir volumes). This ratio is in strumental in choosing the injection pattern. Table 1 shows the production well to injection well ratio for the dispersed injection patterns that have been described in the literature.'-' All of these patterns will have comparable recoveries and response times (for the same basic distance between wells). The choice of pattern there fore depends upon the relative injectivity and productivity of the wells and upon the type of drilling pattern employed. Another important factor may be the ability of the proposed production wells to make up the lost productivity from those wells converted to injectors. These factors become more pronounced at the higher oil viscosities (low gravity oils). All other things being equal, the probable ratio of single well injectivity to single well productivity will be equal to the water-oil mobility ratio of the proposed flood. Thus, high mobility ratios (due to low gravity oil) should call for fewer injection wells — as long as these injection wells are "dispersed" to obtain more even pro duction response for the remaining wells. Secondary re covery projects in oil reservoirs seldom, if ever, will have mobility ratios requiring more injection wells than production wells. Thus, only the last five of the patterns listed in Table 1 will be of significant interest in oil recovery. Almost every development drilling program calls for the wells to be drilled along equally spaced rows and

Jan 1, 1969

By H. Rush Spedden, Arvid Thunaes

R. R. Knobler and F. E. Albertson—Following the testwork done by Thunaes and Spedden, a Sullivan deck plant was built for the Colquiri mill. This plant started to operate in April 1945 and continues in successful operation. Since May 1950 the Colquiri mill has operated with a completely revamped flowsheet at a rate of 1000 tons per day. The old Sullivan deck section was retained with this new flowsheet. The "slime plant" in Colquiri consists of one 12-cell Denver No. 18-Sp. Sub-A flotation machine, one 16-ft and one 12-ft hydroseparator, six five-deck Sullivan-frames and six shaking tables for cleaning up the Sullivan deck concentrates, as well as 12 Deister plus three Plat-O tables on which the fine sands, which are not treated on the S-decks, are concentrated. Only —325-mesh fines are treated on the Sullivan decks. The combined hydroseparator overflows are fed to the S-decks. From 1946 to 1949, i.e. with the "old" flowsheet and an average mill feed of 800 tons per day an average of 45 tons per day was treated on four Sullivan rougher frames. The two additional frames were used as cleaners. At times, up to 90 and 120 tons of slimes were treated successfully on this Sullivan deck installation. The average operating data for the S-deck section in Colquiri for the year 1949 were as follows: The tests reported in this paper indicated a tin recovery of 42 pct. The actual plant results for 1949 show a 54 pct tin recovery from the Sullivan decks. The total average quantity of tin produced from the S-deck section in 1949 was 15 tons of tin per month. This corresponds to 3 pct of the total mill recovery. The difference between the 6 pct overall recovery-increase expected after the tests in 1943 and the actual plant results given for 1949 is explained as follows. According to the tests, a total of 170 tons was expected to be routed to the Sullivan decks, but because of certain changes in the flowsheet and mill amplifications,

Jan 1, 1951

By H. Rush Spedden, Arvid Thunaes

R. R. Knobler and F. E. Albertson—Following the testwork done by Thunaes and Spedden, a Sullivan deck plant was built for the Colquiri mill. This plant started to operate in April 1945 and continues in successful operation. Since May 1950 the Colquiri mill has operated with a completely revamped flowsheet at a rate of 1000 tons per day. The old Sullivan deck section was retained with this new flowsheet. The "slime plant" in Colquiri consists of one 12-cell Denver No. 18-Sp. Sub-A flotation machine, one 16-ft and one 12-ft hydroseparator, six five-deck Sullivan-frames and six shaking tables for cleaning up the Sullivan deck concentrates, as well as 12 Deister plus three Plat-O tables on which the fine sands, which are not treated on the S-decks, are concentrated. Only —325-mesh fines are treated on the Sullivan decks. The combined hydroseparator overflows are fed to the S-decks. From 1946 to 1949, i.e. with the "old" flowsheet and an average mill feed of 800 tons per day an average of 45 tons per day was treated on four Sullivan rougher frames. The two additional frames were used as cleaners. At times, up to 90 and 120 tons of slimes were treated successfully on this Sullivan deck installation. The average operating data for the S-deck section in Colquiri for the year 1949 were as follows: The tests reported in this paper indicated a tin recovery of 42 pct. The actual plant results for 1949 show a 54 pct tin recovery from the Sullivan decks. The total average quantity of tin produced from the S-deck section in 1949 was 15 tons of tin per month. This corresponds to 3 pct of the total mill recovery. The difference between the 6 pct overall recovery-increase expected after the tests in 1943 and the actual plant results given for 1949 is explained as follows. According to the tests, a total of 170 tons was expected to be routed to the Sullivan decks, but because of certain changes in the flowsheet and mill amplifications,

Jan 1, 1951

AS I told you in the chapter on salts, saltpeter is a mixture composed of many substances extracted with fire- and water from arid and manurial soils, from that growth which exudes from new walls or from that loosened soil that is found in tombs or uninhabited caves where the rain cannot enter. It is my belief that it is engendered in these soils from an airy moisture that is drunk in and absorbed by the earthy dryness. Considering its effects, I do not know how to decide and say exactly what the nature of saltpeter is. The learned and very wise physicians find it salty and with a subtle sharpness both in taste and in their medical experiments. Considering its strong biting qualities, they decide to say that it is of a hot and dry nature; on the other hand when they see that it is engendered from the air and that when touched by fire it becomes inflammable and vaporous, and rises with a fearful impetus when mixed in powder for military guns, it manifestly shows itself to be of a hot and moist airy nature. And then when they see that it is shining and transparent with whiteness and is fusible in every fire in a manner corresponding to a thing of a watery nature, and since it. is heavy, it seems that it can be said to be watery in nature. To this is added the experience of touch and of the great cooling action which it has on water in summertime when wine is to be cooled, and since it crushes and breaks with but little pressure, it might be thought to be earthy in nature; especially when it is seen to change into a hard, white stone when burned with an equal amount of sulphur. Thus it seems that every one of the elemental qualities is predominant in it. Now the ancients called this, or a similar thing, nitrum. [150] Pliny, in Book XXXI of his Natural History, has said that it was a thing not much different from salt, that doctors do not seem to have yet known its nature, that it is found in various places, and that the best is found in Macedonia. But among the moderns and especially in our regions, it is not known today what it is that Pliny and other writers refer -to, since according to the aforesaid Pliny and other ancient writers it is a natural mineral. Some say that in place of, the natural kind the artificial will serve for the same medicinal purposes, since it has the same nature and is perhaps more powerful.

Jan 1, 1942

By W. O. Philbrook, K. M. Goldman, G. Derge

EQUILIBRIUM conditions for steelmaking reactions have been studied extensively over the past two decades by a .number of investigators, with gratifying results. Equilibrium data are essential to the understanding of any process, and such knowledge has the practical utility of placing a limit beyond which control measures cannot succeed. When, however, the driving rate of an industrial process is so fast that there is not time for reactions to reach equilibrium, it then becomes necessary that the factors which control the rates of reaction be known in order to establish full control of the process. The desulphurization of molten pig iron by slags within the iron blast furnace is one such process in which the actual degree of sulphur absorption by the slag does not approach the equilibrium distribution obtained for similar slag-metal systems in the laboratory.' A study of the rate-controlling factors governing the desulphurization of iron by slags is becoming increasingly important as the sulphur content of raw materials increases with depletion of higher grade ores and coal. Following up an initial study on kinetics,' this paper reports some results of a continuing laboratory investigation on the mechanism of sulphur transfer between molten iron saturated with carbon and slags of the CaO-SiO,-A12O3 system. Laboratory experiments cannot reproduce in detail the conditions which prevail in the hearth of a blast furnace, but they will contribute to an understanding of the smelting reactions and perhaps will explain some apparent inconsistencies in the factors controlling the equilibrium distribution of sulphur between iron and slags under reducing compared with oxidizing conditions. The mechanism of a reaction—that is, the individual steps and the sequence in which they occur to produce the overall reaction which is ordinarily observed—must be elucidated to learn which steps are the slow ones, or "bottlenecks," that fix the pace of the process as a whole. In the desulphurization reaction, for example, the overall rate might, in principle, depend upon the rate of diffusion of sulphur in the metal or in the slag, upon the speed of some homogeneous reaction taking place entirely within one of the phases, or upon some heterogeneous reaction taking place across the slag-metal interface. Even some side reaction or the back reaction might have some influence. Evidence on all of these points will be presented in this paper. The rate of diffusion of sulphur in both iron and slag is so slow that it seems necessary to assume that convection provides the mechanism of carrying

Jan 1, 1951

By Charles Prasky

ABOUT five years ago the Bureau of Mines began a study, at Minneapolis, of the various methods for beneficiating the low-grade manganese deposits of the Cuyuna range. Several samples of green carbonate slate from the Cuyuna range, containing about 7 pct Mn and about 28 pct Fe, were obtained for this study. The component minerals in the carbonate slate formation were so intimately associated that mineral dressing methods were not suitable for concentrating manganese to a ferrograde product. Although upgrading of manganese and iron minerals has been effected by flotation methods, it appears that manganese cannot be separated from iron except by chemical means. Because the flotation studies at Minneapolis were carried out simultaneously with the hydrometallurgical studies, only crude slate was available for process testing. An important economic consideration in selecting any one of the several chemical methods proposed for recovering ferrograde manganese is the availability of the chemical reagent. Fortunately, large tonnages of sulfur are available on the Cuyuna range in low-grade pyrite and pyrrhotite deposits that are readily concentrated by flotation to a suitable feed for a sulfide roasting furnace as indicated by a Bureau of Mines investigation in 195 0-1951.1 The roasted residue from the sulfide concentrates is a potential source of marketable iron oxide. Therefore, if these sulfide deposits were employed in recovering manganese, utilization of two heretofore unused mineral deposits would be accomplished. Laboratory investigations of several proposed leaching processes were made as a preliminary step in the investigation of hydrometallurgical methods for the treatment of Cuyuna carbonate slate using sulfur as a chemical reagent. From results obtained in the laboratory tests, these leaching processes did not appear entirely suitable for treating Cuyuna carbonate slate. The method developed at the North Central Experiment Station and selected for a more intensive pilot-plant investigation has been termed the differential high-temperature sulfatizing process. Essentially, this sulfatizing process consists of the following steps, described in the order in which the small-scale pilot-plant investigations were made. The first and perhaps most significant step in the process consists of a sulfate roast. Carbonate slate is roasted in an atmosphere of sulfur dioxide gas and air. The overall chemical reaction for converting manganese carbonate to a sulfate at a high temperature in a sulfur dioxide atmosphere follows MnCO8 + SO2 + 1/2 0, = MnSO4 + CO2 (H077°O = —64,180 cal per g mol) .' The need for concentrated sulfuric acid is avoided. Fine grinding is not essential. Removal of carbon dioxide from the slate during roasting produces a more porous structure and permits easier permeation by the sulfatizing gas mixture. The temperature limits for this conversion of manganese have been established as 600" and 850°C. Iron sulfate is unstable and is not formed at these temperatures; phosphorus is not converted to a water-soluble form; and polythionates are not formed. The second step in the process consists simply of a water leach of the sulfatized product. Manganese sulfate in the sulfatized product is dissolved in water. Iron, phosphorus, and calcium remain in the leach residue as water-insoluble products. The third and final step consists of the recovery of ferrograde manganese oxide from manganese sulfate leach solution. Water is evaporated from the leach solution, and impure monohydrate crystals of manganese sulfate are recovered. The moist crystals are formed into pellets or extrusions and thermally decomposed to produce a ferrograde manganese

Jan 1, 1958

By C. D. King

THE key to maximum production of ingots for the war effort is maximum production of pig iron. For any given furnace and ore, the most important single influence on blast-furnace production is the quality of the furnace coke. It is generally recognized that undoubtedly deterioration of Pittsburgh coking coals will be accelerated by the depletion of some of the better grades of coal; the extension of mine mechanization, with attendant degradation of coal quality; and the opening of new coal fields, some of which are known to be fairly high in sulphur. All of these factors, contributing as they do to higher ash and sulphur, and not infrequently to inferior quality of coke, made it desirable that an extensive study be undertaken to determine the effect of washing some coals that closely approximate the average chemical composition of coking coals to be expected in the future by the subsidiaries of the United States Steel Corporation in the Pittsburgh district. It was deemed advisable to study the improvement obtained by washing such coals and to determine the need for supplementing the existing coal washery at Clairton to assure continued maximum production of pig iron from presently available furnace stacks. The existence of the coal washer at Clairton made it possible to conduct tests on coals used in raw form and on similar coals after washing. This study covers an investigation of the quality of furnace coke as affected by washing coking coals, and is confined to the use of IOO per cent high-volatile Pittsburgh seam coals from Fayette and Greene Counties in the manufacture of furnace coke. It covers the improvement in chemical and physical properties as well as the degree of uniformity in coke obtained by washing these coals, and includes the effect on blast-furnace production and practice. The test period involved is four months; the plants, Clairton by-product coke plant, and the Carrie No. I and No. 2 blast furnaces at Rankin, Pa., Carnegie- Illinois Steel Corporation. Clairton By-Product Coke Plant And Coal Washer The Clairton by-product coke plant carbonizes 30,800 net tons of coal daily in 22 batteries containing 1482 ovens. Three oven sizes are used, commonly known as wide, medium, and narrow types, recognizing the difference in oven width. These ovens also differ in height and length. The present coal washer has a daily capacity of approximately 14,000 tons, leaving 16,800 tons to be carbonized in raw form. Washed coal is used on all the narrow ovens and on three of the six batteries with ovens of medium width, raw coals being used on all the remaining ovens. The customary arrangement is

Jan 1, 1944