Search Documents

By L. S. Darken, H. A. Wriedt

THE constitution diagram of the iron-copper system derived by Daniloff' exhibits, at the iron-rich end, phase fields similar to those of the iron-carbon diagram. At 1484° 1094°, and 850°C there are, respectively, a 6-y-L peritectic, a y-E-Lperitectic, and a a-y-E eutectoid. The phase designated E is copper-rich, containing 4 pct Fe or less according to the temperature. The composition of copper-saturated ferrite (a-solid solution) has been determined previously only twice as a function of temperature. By electrical conductivity measurements, Buchholtz and Köster2 found 3.4 pct Cu in ferrite at a eutectoid temperature of 810°C. They represented the composition of saturated ferrite by the equation, log (%Cu) = 4.32-4125 (1/T) Gregg and Daniloff3 reported that Köster later modified his opinion, raising the indicated eutectoid temperature to 833 °C and the copper content of eutectoid ferrite to 3.7 pct. The literature contains several indications4-7 that the work of Köster and his co-workers is not reliable. Norton,8 using an X-ray method, found that 1.4 pct Cu dissolved in ferrite at the approximate eutectoid temperature of 850°C and that the solubility decreased to about 0.3 pct at 700°C. Below 700°C his experimental data show almost no change in solubility. These findings, although inconsistent with the work of Ruer and Goerens4 and of Smith and Palmer,7 have been used generally since their publication. In investigations involving the observation of arrests in temperature during continuous heating or cooling the precise location of the eutectoid temperature has been obscured by a large hysteresis. The more reliable work 579 where the results are not effected by this hysteresis indicates a eutectoid temperature between 840° and 850°C. The foregoing discrepancies, along with a few observations in this Laboratory relating to the precipitation of copper from ferrite, prompted us to rein-vestigate the lower temperature portion of the Fe-Cu system. In this investigation the solubility of copper in ferrite was obtained by the chemical analysis of high-purity iron equilibrated with copper-silver alloys. This method eliminated the necessity for retaining by a quench the physical state of the specimen existing at the equilibration temperature. EXPERIMENTAL The reaction of the iron with the copper alloy was

Jan 1, 1961

By L. B. Slichter

Prospecting is certainly the world's biggest and best gambling business. It is a game where the chips cost many thousands and where many millions, even billions, can be won. An attractive feature of this gamble is the fact that the players are free to rig the odds as favorably as possible. Only the limited bounty of nature, the restrictions of the laws of the land, the competition from other players, the limited sagacity of the player himself— only these, and other factors, restrict the possibilities for large winnings. Today probably about half the land surface of the earth remains to be searched for signs of ore visible at the surface. It is fortunate that so much land still remains for exploration by relatively economical methods. Looking to the future, it is a good guess that the costs of finding ore will increase in response to increasing technical difficulties. To continue to achieve profits from prospecting, it will become necessary to analyze the business more skillfully, in a more and more sophisticated manner. Whenever feasible, qualitative concepts must be replaced by their quantitative equivalents, improvements, and extensions. THE GAMBLE The fact that prospecting is a gambling business reminds us at once of the elementary rule known as the law of gambler's ruin. I hesitate to mention so simple a rule in this company of experts. This rule expresses the rather serious chance of going broke when the odds for success are small, merely by a normal run of bad luck regardless of the long-run expectations of gain. The only sure way of avoiding this special risk of gambler's ruin is to have enough capital, and the will, to continue the play many times and thus ride out the inevitable runs of bad luck. For example, if the probability of success is one in ten for each venture, there is a 35 pct chance that ten successive ventures will fail in a row. But if one has the capital to continue the play through a run of 100 failures, then the chance of gambler's ruin is only 5 in 100,000. Almost any management, I suppose, would be under terrific pressure to cease the play during a long run of bad luck. To continue to prospect under such circumstances requires great confidence in the premises, and a firm conviction in the long-term expectation of profit. There is evidence which I shall mention that in the future even the strongest companies may not have the necessary capital to prospect on the ideal scale. Joint ventures will be very desirable as a means for extending the available venture

Jan 1, 1961

By Palmer C. G, Rowe P. G

Ore handling improvements at Mount Isa Mines Limited were necessary to keep abreast of technological advances in all phases of sublevel stoping practice.In the early 1930s an efficient ore handling system was in operation.However in 1936 excessive water encountered in deepening the hoisting shaft necessitated the sinking of a second hoisting shaft. Financial difficulties prevented abandoning the original shaft and surface ore handling and crushing facilities, so ore was transferred underground and underground crushing deferred. It was not until 1951 Jhat multilevel crushing permitted increased efficiency in the handling of ore from stopes. Ore was then hoisted in one shaft and transferred on the surface by conveyor belt to a new crushing plant.In 1956 a planned production increase from 5,000 to 11,000 tons per mining day made possible further improvements in ore handling. Larger equipment was installed and existing equipment modified to capacitate the handling of larger sized ore at increased daily tonnages.INTRODUCTIONThe intensive exploratory programme commenced in 1954 disclosed large reserves of both lead and copper ores and in 1956 an expansion programme was initiated to increase production from 5,000 to 11,000 tons per mining day. The increased tonnage would result in decreased costs per ton and further cost reductions would result from increased mechanisation underground. The combined effect would permit larger tonnages of lower grade ores to be included in ore reserves. To achieve both cost reduction and the scheduled production increase, improvement in ore handling was of vital importance.For efficient and low cost sublevel stoping it is necessary to maintain an economic balance between...

Jan 1, 1961

By J. E. Campbell, A. W. Wesner, M. Pobereskin

This paper summarizes the results of an investigation sponsored by The Coates Steel Products Company. The objective was to determine the relative effects of certain chemical and physical properties on the wear resistance of grinding balls when grinding cement. In making this study, a number of types of balls were used in both dry and wet grinding in commercial ball mills. The radioactive-tracer technique was used in separating the test balls from the rest of the mill charge and in sorting the different types of halls that were all tested at the same time for each of the grinding operations. PROCESSING OF GRINDING BALLS In order to determine whether the properties of grinding balls, namely, composition, microstructure and hardness, can be related to the rate of wear, the halls used in this study were selected to represent various combinations of these properties. It was recognized that the mechanism of wear was not the same in the wet mill as in the dry mill. Therefore, the types of balls tested in each of the two environments were not necessarily the same. However, all of the types of balls evaluated in the dry mill were tested simultaneously for one time period, while those evaluated in the wet mill were tested simultaneously for three different time periods. All of the test balls were separated from the charge when the mill was dumped. All of the balls were of 0.940-inch diameter when charged, and all but the regular production balls were finish ground to this size to eliminate the decarburized surface layers. The steels and processing methods used, as well as the micro-structures and hardness levels of the balls, are described as follows: Regular production forged SAE 1095 steel balls were used in both the wet and dry mills. They were forged to a nominal diameter of 0.940 inch and quenched in oil from the forging temperature. Each retained a partially decarburized surface layer from the forging bar stock; the maximum depth of this layer was about 0.030 inch and the decarburization was negligible where the ball was sheared from the bar. The hardness at the surface was 40 to 46 Rock-well C in the partially decarburized area and 57 to 62 Rockwell C in the areas not decarburized. The hardened zone extended to a depth of about one-third of the radius. Below the decarburized layer, the hardened zone consisted primarily of hard as-quenched martensite with varying amounts of retained austenite. This structure changed to fine pearlite at a depth of about one-third of the radius. In order to eliminate the effect of the decarburized layer, a set of production balls of SAE 1095 steel was forged to about a 1-inch diameter, quenched in oil from the forging temperature, and ground on ball-bearing grinding equipment to a 0.940-inch diameter to remove the decarburized layer. The balls' hardened zones were similar to those in the previous set. The hardness of the finished surfaces was 52 to 60 Rockwell C. A slight amount of decarburization apparently remained after the removal of about 0.030 inch from the surface. Balls of this type were tested in the dry mill only. Another set of forged SAE 1095 steel balls of 1-inch diameter was heated to 1500 F and quenched

Jan 1, 1961

By P. D. Rao, H. B. Charmbury, D. R. Mitchell

As a result of a changing market pattern for anthracite, there is now an increasing demand for steam size and a decreasing demand for domestic sizes. To help the producer meet this new demand, the authors present the technical aspects of the problem of crushing domestic sizes to steam fineness and size. The market pattern for anthracite coal has changed considerably in recent years. The demand for domestic sizes has been gradually decreasing and the demand for the steam sizes has been increasing. The companies however aim for a maximum production of domestic sizes, since these sizes command premium prices. To meet the market demands, particularly in certain periods of the year, further size reduction of the coarse clean coal is necessary. The present practice of multistage crushing in toothed rolls leads to complicated flow arrangements and is high in cost. Data presented herein indicate that crushers with high reduction ratios could be used with a resulting simplification of crushing flows and an appreciable reduction of capital and operating costs. The different types of crushing equipment investigated included the rotary-type breaker, the jaw crusher, the impactor-type crusher, and the Aerofall mill. CURRENT PRACTICE The standard sizes of anthracite are presented in Table I and market specifications require that any given size be maintained with only a limited amount of oversize and undersize material. The 1958 prices of the various sizes also are presented, so it may be seen that size reduction is only performed when dictated by market demands. At the present time the general practice of crushing anthracite consists of removing the top size and crushing it in a toothed double roll crusher to the next largest size. Consequently, to reduce the domestic egg, stove, and nut sizes to steam sizes of pea, buck, rice, and barley, the original material must be screened to produce the three top sizes and each size must be crushed in three separate rolls to the next largest size. Each crushed product must then be rescreened to remove the oversize for re crush ing. Thus a typical operation would consist of scalping the egg size from a mixture, crushing it to stove size, and rescreening the product. This same cycle would be repeated for the stove and nut sizes requiring three screens and three crushers as well as conveying equipment to handle the crushed products. This results in an expensive series of operations to make a product which sells for a lower price. If the three top sizes are crushed in one operation to the smaller sizes then the yield of steam sizes is low and the production of Nos. 4 and 5 sizes is large. These sizes sell for lower prices than the steam sizes. A typical operation is presented in Fig. 1. From the figure it may be seen that the original cleaned anthracite contains 67.4 pet domestic sizes, 26.4 pct steam sizes, and 6.2 pct fines. After multistage crushing of the domestic to steam sizes the composition of the product is 0.6 pct domestic sizes, 89.6 pct steam sizes, and 9.8 pct fines. From a technical standpoint these results are reasonable but the handling and crushing costs are

Jan 1, 1961

The development and introduction of a preventive maintenance scheme in the coke ovens department, coal preparation section, of The Broken Hill Proprietary Co. Ltd., Newcastle steelworks, is described together with its operation and the results achieved during the first 12 months.'fhe scheme consists essentially of an engineering analysis of the plant to establish maintenance schedules, and to record the details on reference cards.Items are reviewed progressively and a weekly work schedule is prepared, from which work orders are passed on to the maintenance foremen. These foremen in turn originate action orders and issue inspection sheets as required. Results are then reported back to the maintenance planner and used to revise norms, practices, and cycles.INTRODUCTIONWhat is maintenance? Some people consider maintenance to be the work involved in putting things together again after they have fallen apart. The author regards it as the mechanical care and repair of plant machinery and equipment to keep it operating in sound mechanical condition, without undue interference with production processes.Maintenance may be organized or haphazard, and, too often, what has been fondly imagined to be nicely controlled and organized maintenance proves, under pressure, to be more correctly described as haphazard. Haphazard maintenance deals with breakdowns as they occur, and frequently finds itself confronted with more breakdowns than can be handled. Even organized maintenance, if not based on sound planning, can, at times, be found in a chaotic state, because of unexpected and unanticipated breakdown.Properly planned and scheduled maintenance permits an orderly and controlled approach to the problem of maintain...

Jan 1, 1961

By H. J. Modi, D. W. Fuerstenau

pH effectively regulates the flotation of corundum through its control of the surface charge. Since collector ions function as counter ions in the double layer, the collector must be anionic when corundum is positively charged and cationic when it is negatively charged. Even a long-chained xanthate will float positively charged corundum. Association of the hydrocarbon chains of the adsorbed collector ions gives rise to increased adsorption and increased rate of flotation. Because of this phenomenon, neutral organic molecules aid flotation although the unionized collector cannot promote flotation independently. Na2SO4 was found to depress the flotation of corundum with animic collectors through ionic competition whereas the same salt activates corundum for flotation with cationic collectors. A close connection between ion association, electrokinetic behavior, and floatability is suggested. As part of the increasing interest in the physical chemistry of flotation, strong experimental evidence has been obtained which indicates that organic ions are adsorbed as counter ions by non-metallic oxide minerals mainly because of the electrical forces which exist at interfaces(1-3). Thus flotation of such minerals without activation should then be appreciable only when the mineral and the collector are oppositely charged. Since only limited work has been reported on the effect of surface charge upon flotation behavior(3) this general hypothesis on the flotation of insoluble oxide minerals has not been tested thoroughly. Investigating the flotation of a mineral for which extensive electrokinetic data exist yields a way to test this hypothesis more extensively. The electrokinetic behavior of corundum in solutions containing both organic and inorganic electrolytes has been studied quite thoroughly' Electro-kinetic measurements have shown that in the case of corundum, simply by controlling pH the solid surface can be charged either positively or negatively. For this reason corundum was chosen for study. The object of this paper is to present a study of the flotation of corundum and to explain the results in the light of our knowledge of adsorption mechanisms gained from streaming potential investigations. This paper is concerned mainly with the effect of surface charge on flotation with anionic and cationic collectors, but in addition such flotation phenomena as collector hydrolysis, Van der Waals attraction between hydrocarbon chains, ionic competition, and activation are discussed. While the present paper is confined to a reporting of the data obtained and some of the more immediate conclusions, it will be seen that the results, in combination with those on other systems, do promise a

Jan 1, 1961

By Kenneth M. Haley

THE development of Reserve Mining Company's pelletizing practices began in 1949 in a pilot plant located at Ashland, Kentucky. Three years later the participating companies -Armco Steel Corporation, Cleveland Cliffs Iron Mining Company and Republic Steel Corporation concluded the work and the small plant was shut down. Oglebay Norton and Company was the operating manager and Armco Steel Corporation furnished labor and services. Arthur G. McKee and Company were the designers and constructors. The equipment flow sheet that was employed in that pilot operation is shown in Figure 1. This was Reserve's first step from the laboratory work per-formed by the Mines Experiment Station of the University of Minnesota in developing a commercial operation. The plant used New York state concentrate, purchased from Jones and Laughlin Steel Corporation's Benson Mine. This concentrate, minus 14 mesh, 64.3 per cent Fe, 5.7 per cent Si02, was processed at Ashland to 80 per cent minus 325 mesh in conventional ball mill, classifier, thickener, filtering

Jan 1, 1961

By Lionel A. York

THE first application of the cementation process took place in the year 1864, according to H. Standish Ball, when a break in the brickwork lining at a depth of 270', in one of the Rhine Preussen pits near Homburg, was sealed by running in milk of cement from the surface by means of a small hand pump. At about the same time Monsieur Portier adopted the method in France for repairing breaks in shaft lining. Figure 1 is descriptive of the Portier system and is of interest when compared with modern practice. During the next thirty years numerous attempts were made to improve the process; the general method used in sinking being to put clown bore holes outside the perimeter of the shaft for the purposes of injection. The cementation process was first recognized as an important as-set to mining in 1896 when Monsieur A. Francois developed his method of boring and injecting inside the shaft below the permanent water level. The Francois cementation process rapidly attained fame on the Continent and many shafts were sunk by its use in France, Belgium and

Jan 1, 1961

By H. H. Wade, N. F. Schulz

The large quantities of iron-bearing materials, including taconite, semi-taconite,* and other low-grade ferruginous materials occurring in Minnesota and elsewhere, constitute an important potential source of iron. Satisfactory grade and recovery of iron ore concentrate cannot ordinarily be obtained from these ferruginous materials by simple methods of beneficiation. Iron is distinctive in that the free metal and two of its oxides [magnetite (Fe2O3) and maghemite (gamma-Fe2O3) ] are strongly ferromagnetic. The successful magnetic separation of natural magnetite from taconite has already been established on a commercial scale. The iron in semi-taconite, however, occurs largely as goethite (Fe2O3-H4O) and hematite (Fe2O3) which do not respond to ordinary magnetic separation methods. Magnetic roasting converts these iron oxides to magnetite, which can, after liberation, be magnetically concentrated. MAGNETIC ROASTING Hematite undergoes a stepwise reduction when exposed at moderately elevated temperatures to a reducing atmosphere containing hydrogen or carbon monoxide: Fe2O3? Fe3O4 ? FeO ? Fe Since the desirable product in magnetic roasting is Fe3O4, the conversion is halted after the first stage by controlling the reduction potential of the gas mixture with water vapor or carbon dioxide. The reducing gas mixture must be such that the oxygen partial pressure at the solid-gas reaction interface will be maintained intermediately between the oxygen partial pressures in equilibrium with Fe2O3 and Fe3O4 at the prevailing temperature. Because of the rapid increase in decomposition pressure of Fe3O4 with increase in temperature, the practical upper limit of useful reaction temperature is about 1400°F, above which over-reduction usually occurs in atmospheres containing hydrogen or carbon monoxide. On the other hand, useful reaction rates usually occur only above 800°F. Goethite reacts similarly except that the combined water probably has to be driven off thermally before the reduction to magnetite occurs. Iron carbonate must also be decomposed thermally to an oxide before conversion to magnetite. Iron silicates are apparently not amenable to magnetic roasting. The rate of penetration of the reaction toward the interior of a given ore fragment is dependent on the chemical composition and physical structure of the solid. Most ore fragments tested here reduced satisfactorily to 1/4-in. depth in less than 14 min at 1200°F. Some very dense fragments, however, were reduced to less than 1/32-in. depth under the same conditions. Determination of the optimum temperature, gas composition, ore fragment size, or time of exposure

Jan 1, 1961

By S. M. M. Safvi, A. Jowett

Small-scale continuous flotation tests are described in which the influence of 1) pulp density and 2) feed rate on the rate of flotation are investigated. The results provide further evidence of first-order kinetics for flotation and indicate that tests at different feed rates provide a particularly simple method of determining rate constants. Flotation kinetics are discussed in general terms. Over a period of many years numerous research workers have studied the rate of recovery of valuable materials from flotation pulps with a view to deter-mining the mechanism of the froth flotation process. Froth flo ation is fundamentally a collision process involviug at some stage collision and then mutual attachment between air bubbles and mineral particles. It is possible to draw analogies therefore between the flotation process involving mainly macroscopic particles and the ordinary processes of chemical reaction kinetics involving collisions between atoms, molecules and ions. By adopting this analogy we may postulate a general equation representing the kinetics of the flotation process: -dC/dt =K.CC ,Aa (1) The rate constant, K, will be a complex function involving reagent concentrations, particle and bubble sizes, induction times and no doubt flotation cell design. However, these are factors which can be kept constant in any series of experiments, permitting investigation of the exponents a and c, which determine the order of the reaction occurring in flotation. The various research investigations carried out so far have all aimed at evaluating a and c and in the majority of cases only since the concentration of air has usually been kept constant in a series of tests, so that a relationship of the following type has been investigated: -dC/dt = K1 Cc (2) The techniques of both batch testing and continuous testing have been used to investigate this relationship. The technique of continuous testing, in which steady-state conditions are established, permits direct investigation of the above equation. In the case of batch testing, C is decreasing continuously

Jan 1, 1961

By S. A. Falconer

The benefication of non-sulfide or non-metallic minerals by froth flotation has been practiced commercially for at least 30 years. In the pioneer stages of development of this flotation art, most of the promoters investigated were fatty acids or soaps that had been found to float sulfide minerals. When commercial application of froth flotation started to assume importance in the late 1920' s and 1930's, a great deal of fundamental research work was done in the laboratory (11-14, 19, 22, 23, 26, 29, 32). Among other subjects, this work included studies of the relative effectiveness of the different fatty acids and their salts on the flotability of various non-sulfide and non-metallic minerals. The successful present-day use of fatty acid and petroleum sulfonate type promoters results from better understanding of the principles of non-sulfide flotation plus the developnent of effective selectivity modifying agents. Many investigators have contributed to a greater appreciation of the numerous interrelated factors involved in fatty acid and soap flotation of non-sulfide minerals. Of these, mention should be made of recent fundamental studies by Tagart and Arbiter (31); Wolstenholme and Schulman (35); Schulman and Smith (28); Wark and his co-workers(3,4) Du FIietz(l5); and J. Leja(24). According to J. Rogers, K. L. Sutherland and E. E. and I. W. Wark(27), "the basic requirements for a 'collector' in flotation are that it shall form a film over the surface of the mineral to be floated and that this film should be oriented so that hydrocarbon (non-polar) groups form the effective surface." When modified in that way, the mineral surface possesses some degree of air avidity. All investigators would agree with Taggart, Taylor and Ince (33)that such a film can be produced only if the collector compound contains an active (polar) group to be attracted to the mineral surface, as well as the non-polar group. The commonly used anionic-type promoters for non-sulfide, non-metallic and metallic oxide ores include: 1. Fatty acids, and their soaps and emulsions. 2. Fatty acids residues — "pitches" and "foots." 3. Sulfonated and sulfated fatty acids and so-called "petroleum sulfonates." 4. Paraffin chain salts — sulfonated and sulfated higher alcohols. 5. Cyclic organic acids or salts, such as naph-thenic acid and naphthenates. Certain fatty acids are more effective than others as promoters for particular minerals under certain specified conditions. Some of the more costly fatty acids or collectors of the paraffin chain salt type may give improved selectivity in some instances and are being used commercially in a limited number of operations. Local availability may dictate the choice of the fatty acid used — for example, the use of palmoil fatty acid for flotation of copper carbonate ores in the Belgian Congo. Coconut oil fatty acids have also found limited use.

Jan 1, 1961

By K. K. Landes

The north central and northeastern states supply over 50 pct of the chemical and metallurgical limestone produced annually in the United States, and Ontario is the leading source of this material in Canada. About three fourths of the chemical and metallurgical stone produced in this area comes from Michigan, Pennsylvania, Ohio, and Ontario. The specifications, physical and especially chemical, for chemical and metallurgical stone depend upon the end use, but in general the tolerance for impurities is so low as to rule out most limestones and dolomites. Both high calcium limestone and dolomite are used in the metallurgical and chemical industries. The principal metallurgical use is as a flux in iron ore blast furnaces and open hearth steel furnaces. The major chemical uses are in lime burning and soda-ash manufacture. Other important uses include glass manufacture, in making calcium carbide, and in sugar factories. All deposits of high quality carbonate rock are definitely finite. Invariably and inevitably they diminish in either thickness or quality laterally so as to become unworkable as chemical and metallurgical stone. The size of these high quality lenses ranges from a few thousands of tons to hundreds of millions of tons. The best hunting for new supplies is in the geologic formations known to contain premium quality deposits. However, because of high transportation costs, carbonate rocks, to be workable, must be close to market, dollarwise. Definitions: Chemical stone is limestone or dolomite used as a raw material by the chemical industry. The major chemical uses are in lime burning and soda-ash manufacture. Other important uses include glass manufacture, making calcium carbide, and in sugar factories. The specifications for chemical stone depend entirely upon the end product. Some uses require dolomite, some require high calcium limestone; in a few instances, notably in lime burning, either type of carbonate rock can be used as raw material. As a general rule, a fairly pure stone is necessary and for certain particular uses the tolerance for some elements such as sulfur, iron, and phosphorus may be extremely small. Metallurgical stone is used as a flux in furnace operations, especially in the blast furnace where iron ore is converted into pig iron and in the open hearth furnace where pig iron is made into steel. Either limestone or dolomite can be used in the blast furnace, but only high calcium stone is charged in the open hearth (except for rice-size dolomite used as a refractory). There are no rigid chemical specifications for blast furnace stone, but low silica, alumina, sulfur, and phosphorus are desirable. For open hearth use, in addition to low magnesia, a very low sulfur content is also specified by many furnace operators. For furnace use, fluxstone also must be sufficiently shatter resistant that it will remain in lump form during transport and charging into the furnace. Commercial stone is the term applied in the trade for crushed stone used for roadstone and in concrete aggregate. Many of our best chemical and metallurgical limestones are too soft to qualify in the Los Angeles abrasion test for use as commercial stone. However, some limestones, and a higher proportion of dolomites, do qualify for commercial stone and so can be sold to either market. Place Value: Even the highest grade limestone is still a bulk commodity with a relatively low value per ton. As a matter of fact there is very little difference in the average price per ton between high grade chemical and metallurgical stone and crushed limestone used commercially. Most high quality stone is produced in large operations with consequent low cost per ton whereas most commercial stone is produced in relatively small local quarries. As in the case of other bulk commodities most limestone buyers pay more for transporting the raw material than the cost of the stone at its source. For this reason steel furnaces and chemical plants

Jan 1, 1961

By N. A. Gokeen, M. Ohtani

EFFECT of elements on the solubility of carbon in liquid iron is useful in calculating and correlating a number of thermodynamic properties as shown elsewhere in detail.' Kitchener, Bockris, and spratt2, and later Turkdogan and Hancock3 have investigated the effect of sulfur on the solubility of carbon at various temperatures and concentrations. For any chosen sulfur concentration, the solubility of carbon reported by Kitchener et al. is considerably lower than that by Turkdogan and Hancock. It was therefore felt that a reinvestigation of this problem was necessary. EXPERIMENTAL METHOD An alloy of Fe-C-S .was prepared with electrolytic iron and ferrous sulfide melted by induction in a graphite crucible, 4 in. high 1 1/2 in. ID under a CO-atmosphere. The temperature, held at 1500" ± 5°C in all experiments, was measured with a Pt-Pt + 10 pct Rh thermocouple immersed into the melt. In order to assure saturation with carbon, the melt was stirred with a graphite rod. After approximately 1 hr at 1500 °C, a sample was obtained by sucking the melt into a silica tube, 4 mm ID. The sample was then analyzed for carbon and sulfur by usual methods. RESULTS The results are presented in Table I and plotted in Fig. 1 with the individual values of Turkdogan and Hancock.3 Scattering in the data is equivalent to 3 pct or less in carbon analysis. The broken line represents the empirical equation obtained by Kitchener et al. from their data. It is evident that the present data are in good agreement with those of the former investigators but in disagreement with the latter. The cause for this discrepancy cannot be readily offered since the experimental methods in the three investigations were nearly identical. Additional evidence corroborating the author's results may be presented in the following manner. The difference between the solubilities of carbon in the ternary Fe-C-X melt with a small concentration of any element X, and in the binary Fe-C melt is %?C = %CIII-%CII [1] where %?C is the change in solubility, and %CIII and %CII are the solubilities in the ternary and the binary systems respectively. The value of %Cii

Jan 1, 1961

By C. H. G. Bushell, G. Brown, C. J. Krauss

1N FLOTATION, copper sulphate is used universally for the activation of sphalerite and marmatite and, to some extent, for the activation of other minerals. The copper activated surface readily adsorbs standard collectors such as the xanthates and dialkyl dithiophosphates whereas untreated zir~ sulphide does not. With the host of reagents available, making a mineral float is seldom difficult. The problem is to achieve selective flotation. In zinc flotation, the great value of copper sulphate results from its ability to activate zinc sulphide preferentially to gangue minerals. At the Sullivan Concentrator of The Consolidated Mining and Smelting Company of Canada Limited, the most abundant gangue mineral in the lead-zinc ore is pyrrhotite, and the zinc mineral is marmatite of somewhat variable zinc to iron ratio. For this Cominco study, selectivity between marmatite and pyrrhotite was therefore the most interesting aspect of copper sulphate activation.

Jan 1, 1961

By R. A. Kearns

Biennially the Bureau of Mines issues a list of equipment that has been accepted as approved or permissible during the two calendar years immediately preceding the date of publication of the listing. By so doing the mining industry is kept informed of the availability of the latest equipment for safe use underground. This list supplements Bulletin 543, Permissible Mine Equipment Approved to January 1, 1953; Information Circular 7722, Permissible Mine Equipment Approved During the Calendar Years 1953-54; Information Circular 7840, Permissible Mine Equipment Approved During the Calendar Years 1955-56; and Information Circular 7957, Permissible Mine Equipment Approved During 1957-58. This list includes equipment meeting those standards that have been revised since Information Circular 7957 was .issued. Schedule 16E,4 a revision of Schedule 160, was approved May 19, 1960.

Jan 1, 1961

By Evan Just

?We are certain there is need for research in every phase of mining." This statement by R. M. Crump of the Jones and Laughlin Steel Company (047) paraphrases opinion widespread among mining people. It is undoubtedly true. In preparation for this presentation, 193 organizations were asked to describe their mining research efforts. 60% responded, The meaning of "research" was taken in several different ways. The continual efforts of progressive companies to improve themselves were reported as research. We will call this and any effort to solve specific and immediate problems "practical research". The continuing investigation of subjects of long range interest will be called "basic research".

Jan 1, 1961

By L. Adler

Two basic cases involved in the design of an opening in bedded rock are: 1) where the beds deflect from each other so as to be separated; and 2) where the beds deflect onto their lower neighbor, loading the latter while reducing their own load. Analysis of case 1 is relatively simple, while the interference of the original deflection curves in case 2 complicates analysis of the latter. Figures and formulae are presented for determination of interference loading in this situation. The design of an opening in bedded rock involves two basic alternatives.1'2 In one case the beds deflect away from each other, so as to be separated from their neighbors; in the other case the beds deflect onto their lower neighbor, loading it, while reducing their own load. This is shown in Fig. 1. The first case lends itself to a simple analysis, since the upper and lower boundaries of each beam remain a free surface. In the latter case, however, due to the interference of original deflection curves, the analysis is more complex. Earlier investigators of this interference problem solved it by making the tacit assumption that this type of loading was uniformly distributed along the beam's length. While there is no intuitive justification for this assumption, its use does lead to an approximate solution. Still, a more exact solution does not require this apparently unrealistic restriction. To achieve initial simplicity, the analysis will at first be limited to a two-bedded sequence, as in Fig. 2. Since both beams have identical end conditions and spans, the original deflections of their center lines are members of the same family of curves, as shown in Fig. 3. Therefore, if interference occurs at one point along their center lines, it will occur at every point. For convenience, the midspan point is that usually examined. Deflection at midspan for a uniformly loaded beam regardless of end conditions is: where c is a constant depending on end conditions; w is the body load and equals yh, lb per ft; y is the specific weight of rock, lb per ft'; L is the span, ft; E is Young's Modulus, lb per ft2; I is moment of inertia (for a rectangular section), I = bh3/12 = h3/12, (ft4); b is the width of beam and equals one; and h is the thickness of beam, ft; or 3U > yB an Substituting the appropriate values into Eq. lb, interference between the two beds can be predicted. Should it occur, the additional loads along the lower boundary of the upper bed and the upper boundary of the lower bed must be determined. From the principle of action and reaction, these added loads are known to be equal and opposite, as indicated in Fig. 4. Since the final deflection curves for both beds are identical (Fig. 3), their successive derivatives of y with respect tox must also be identical. That is:

Jan 1, 1961

By R. C. Neavel

One of the most important requirements to be met by any coal producer is delivery of size consists tailored to specific applications. Generally these specifications include a minimum as well as a maximum size limit. The customers of a major Indiana coal producer noticed that some shipments of sized coal contained a large proportion of undersize pieces. Also, in these shipments, the pieces which remained within the specified limits were cracked and, when handled, crumbled readily. The producer, certain that these shipments had been properly sized at the tipple, suspected that degradation had occurred in transit. The Indiana Geological Survey was requested to examine the coal from one of these shipments to attempt to determine the cause of breakage. Two basic facts about coal and coal seams led to the employment of petrographic techniques in this investigation. First, all coal seams consist of accumulations of diverse organic and inorganic constituents; and second, most coal breakage, either desirable or undesirable, can be attributed to the inherent weakness of some of these constituents or to poor adhesion between certain constituents. Analyses of degraded coal Sampling: Samples were secured from 2 stock piles of 2-by 4-inch coal in a dealer's yard. The coal in both piles had received the same treatment from mining to storage, although it had been mined and shipped on two different days. The size consist of one pile, however, appeared smaller than that of the other. The sample from the degraded pile is referred to as the slack sample, and the other is called the unslack sample. Size analyses: In order to determine the actual size consist, the samples were screened, with a minimum of rough handling, on square 2-inch, 1-inch, 1/2-inch, and 12-mesh screens. It is assumed that 100 percent of the 2-by 4-inch coal is retained on a 2-inch round screen at the tipple. Analyses of several loads at the tipple indicate that approximately 26 percent of the 2- by 4-inch coal, 100 percent of which was assumed held on a 2-inch round screen, passed through a 2-inch square screen. Twenty-six percent may be taken as the normal difference between round and square screens. Results of the analyses of stockpile coal are shown in Table 1. Cumulative percent versus size are plotted on a Rosin-Rammler chart as Figure 1. Both samples approximately follow normal size distribution; this indicates that breakage is not size preferential and also that the samples are representative of the stockpiles. Although both samples exhibit the same general size distribution, the median size of the slack sample is considerably smaller than that of the unslack sample. Fifty-five percent of the unslack sample was undersize (29 percent greater than the normal 26 percent difference), and 83 percent of the slack sample was undersize (57 percent greater than the normal 26 percent difference). Both piles contained a greater percentage than normal of undersize coal; this indicates that they were both affected by degradation, but to a different degree. Bedding-plane analyses: Coal, because of its laminated nature, will break preferentially along bedding planes, especially where adjacent constituents are poorly cemented to each other. Subsequent breakage perpendicular to the

Jan 1, 1961

By Marvin L. Whisman

Several parameters controlling the reproducibility of tritium determina¬tion in organic compounds by liquid scintillation counting were evaluated by the Federal Bureau of Mines. These included optimum volume of scintillator solution for maximum counting efficiency, reproducibility of counting a sealed commercial standard and comparison with duplicate counts on laboratory-prepared samples. The error introduced by using microliter pipets, optical differences in counting vials, reproducibility of counting organic samples where quenching is a factor, reproducibility of counting laboratory-prepared samples with an added internal standard, optimum ratio of internal standard activity to sample activity, effect of sample volume upon counting efficiency, count rate at which coincidence losses become apparent, and reproducibility of organic compound radioassay by the standard addition technique are outlined. From the evaluation of these parameters, a technique that can be applied routinely to organic compounds containing tritium has been established, and details are included in this report.

Jan 1, 1961