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By P. K. Koh, H. A. Lewis, H. F. Graff

New data on the distribution of silicon between slag and carbon-saturated iron at 1600Oand 1700OC are presented which, in combination with previously published data, permit the determination of silica activities over a broad range of compositions in the CaO-Al2O3-SiO2 system. The distribution of silicon between graphite-saturated Fe-Si-C alloys and blast furnace-type slags in equilibrium with CO has been described in previous publications.1"3 In this past work the silica-silicon relation was established at temperatures of 1425" to 1'700°C for slags containing up to 20 pct A12O3. This paper presents the results of additional studies at 1600" and 1700° C which extend the silicon distribution data at these temperatures for CaO-A12O3-SiO, slags over a range from zero pct Al2O3 to saturation with Al2O3, or CaO.2Al2O3. The upper limit of SiO2 is set by the occurrence of Sic as a stable phase when the metal contains 23.0 or 23.7 pct Si at 1600" or 1700°C, respectively. The activity of silica over the expanded range is determined directly from the distribution data.3 Recently4-7 other investigators have studied the activities of SiO, and CaO, principally in the binary system, using different methods and obtaining somewhat different results. EXPERIMENTAL STUDY The experimental apparatus and procedure have been fully described in previous publications.1, 3 Six new series of experimental heats have been made, four at 1600° and two at 1700°C. Master slags of several fixed CaO/Al203 ratios were pre-melted in graphite crucibles, and these were used with additions of silica to prepare the initial slag for each experiment. Slag and metal were stirred at 100 rpm and CO was passed through the furnace at 150 cc per min. The initial sample was taken 1 hr after addition of slag at 1600°C or 1/2 hr after addition at 1700°C. The run was normally continued for 8 hr at 1600°C or 7 hr at 1700°C, and the final sample was taken at the end of this period. Changes in Si and SiO2 content indicate the direction of approach to equilibrium, and in a series of runs where the approach is from both sides this permits approximate location of the equilibrium line. Fig. 1 shows the results of such a series of 15 runs at 1600°C for slags of CaO/Al,O3 = 1.50 by weight. Figs. 2 and 3 record other series at 1600°C and Fig. 5 a series at 1700°C with fixed CaO/Al0 ratios. The results of the experiments at 162003°C have been reported in part in a preliminary note.3 In the experiments recorded in Figs. 4 and 6, the slags were saturated with A12O3 (or with CaO.2A12O3 within its field of stability) by suspending a pure alumina tube in the melt during the course of the run. The final slag analyses were used to establish the liquidus boundaries8 in the stability fields of CaO.2Al2O3 and of Al20,. ACTIVITY OF SILICA The free-energy change in the reaction has been calculated by Fulton and chipman2 from recent and trustworthy data including heats of formation, entropies, and heat capacities. The more recent determination by Olette of the high-temperature enthalpy of liquid silicon is in satisfactory agreement with the values used and therefore requires no revision of the result which is expressed in the equation: SiO2 (crist) + 2C (graph) = Si + 2CO(g.) [1] &F° = + 161,500 - 87.4T The standard state for silica is taken as pure cristobalite and that of Si as the pure liquid metal. Since the melts were made under 1 atm of CO and were graphite-saturated, the equilibrium constant for Eq. [I] reduces to K1 = asi /asio2. The value of this constant is 1.77 at 1600°C and 16.2 at 1700°C. Through K1, the activity of silica in the slag is directly related to the activity of silicon in the equilibrium metal.

Jan 1, 1960

By R. L. Ash, R. R. Rollins, C. J. Konya

An investigation was performed to determine conditions for optimizing the spacing of simultaneously initiated multiple explosive columns. This was done by using models of mortar, dolomite, and Plexiglas with 10-grain mild detonating fuse as the explosive charge. It was desired to simulate blastholes with multiple primers initiated by detonating fuse or when high-velocity explosives are used in low-velocity materials. It was found that optimum spacing between multiple charges was strongly influenced by charge length. At less than optimum charge length, the spacing at which complete shearing was possible between adjacent charges decreased exponentially with a subsequent loss of broken material volume. For charges fired simultaneously, larger burdens and spacings were possible as compared to those necessary for single-crater charges. For each material studied, there was a characteristic optimum charge length and a maximum attainable spacing at any given burden. Proper selection of the spacing distance between charges is fundamental to successful blasting. Its value directly affects the cost of drilling and explosives used per unit of broken material. In addition, the choice of a spacing that is Compatible with a given set of blasting conditions aids in the control of fragmentation sizing, ground vibrations, overbreak, and throw which in turn, influence other production costs. For example, normally loaded blastholes that are spaced too closely invariably promote overbreak and usually give coarse fragmentation. Unless care is taken, airblast and violent flyrock will occur and under certain conditions cutoffs and misfires may result. Too large a spacing, on the other hand, frequently leads to conditions that form bootlegs or toes. The choice of a particular spacicg to use, however, is largely a matter of individual experience and judgment, usually based on trial and error. Very little is known or can be found in the literature with regard to how the spacing between charges is related to field conditions and charge geometry. As a general rule, the firing time sequence of adjacent charges and properties of a material are thought to have the most significant influence on the spacing distance best suited for any given field condition. For example, delayed initiation of adjacent charges usually always requires a closer spacing than when charges are fired at the same time. This should be expected if one considers that the energy normally dissipated and lost in the surrounding ground from charges fired independently would be captured and utilized for breaking material between charges when they are initiated together. Spacing can be extended also when charges are aligned with structural planes of a material, such as jointing, along which shearing is relatively easy. It is customary to relate the spacing (S) between charges to their common burden (B) in the form of a spacing ratio, or SIB. The burden normally is considered as the optimum depth or distance from any single charge perpendicular to the nearest free or open face at which the desired fragmentation and maximum crater yield are obtained. For production blasting, value of the ratio is generally considered to vary from 1 to 2, depending on conditions.1-6 When adjacent charges are fired independent of one another, the value varies from 1 to about 1.4, the closer amount being employed to square corners or produce craters having the ideal 90" apex angle. The larger ratio is the geometric balance value for craters having an apex angle of 135". The basic ideal crater forms in the plane of the charge diameter for charges fired independently are shown in Fig. 1. In the event charges are fired simultaneously, geometric balance in the plane of their charge diameters suggests that a spacing ratio near 2 would be appropriate, as illustrated by Fig. 2. In practice, however, some compromise ratio value must be selected to conform with the specific ground conditions. An example would be where the jointing planes tend to produce 60° or 120° crater angles, the appropriate geometrically balanced charge arrangement being given by Fig. 3. In this condition, the spacing ratio is 1.15, not 1 or 1.4 as suggested for the 90° cratering of independently fired adjacent charges. In view of the foregoing, it would seem logical to assume that whenever charges all having the same burden are fired at the same time, spacing distances always can be greater than those permitted by charges fired independently. In practice this is not the case, however.

Jan 1, 1970

By C. M. Loeb, A. M. Gaudin, J. H. Brown

THE object of most size reduction operations in the mineral industry is to liberate the grains of valuable minerals in the ore from those of the gangue. This is usually accomplished by crushing and grinding the entire mass of ore until there is only a small probability that any single particle contains more than one mineral. During this size reduction only limited control exists over size or composition of the particles exposed to the breaking action, and there is no control over the paths followed by cracks generated during the operation. This lack of control usually results in overgrinding and in production of large quantities of very fine material. The first detriment, overgrinding, is costly in itself, but when combined with the second factor it is doubly so. Not only is the fracture of a free particle unnecessary—the fracture of these particles may also make subsequent separation operations difficult, inefficient, and wasteful. It has been pointed out' that if the object of size reduction is to liberate the valuable mineral component of the ore then, ideally, fracture should follow intergranular paths to the exclusion of trans-granular ones. This would result in liberation of the valuable minerals with as little size reduction as possible. This ideal comminution operation is referred to as intergranular comminution, and it was the object of the investigation to determine the extent to which it could be developed by heat treatments. There are many indications in the literature that heating rocks prior to crushing may be favorable. Reports by Holman,2 Yates3 and Myers' are pertinent. These investigators showed that heating certain rocks prior to crushing them did, in fact, improve their crushing characteristics in that fewer fines were produced, although the fact that intergranular comminution was being effected apparently was overlooked. In addition, Sosman noted that if there is appreciable anisotropism in the thermal coefficients of expansion of even a pure mineral, then considerable permanent separation of the grains of the rock can be expected as a result of heating the rock to a high temperature.' By the same token, if there are ap- preciable differences in the thermal expansion coefficients of the various minerals of a multi-component rock, similar results should be obtained by heating this rock. This has been tested, partially, by Brenner," who obtained patents covering the heat treatment of some pegmatitic rocks in order to facilitate comminution of these materials. It has also been demonstrated that this may occur in taconite." Also, the possibility of causing decomposition of one mineral in a rock as a means of promoting intergranular fracture has been considered. Seigle2 and Schiffman et al. have obtained patents on such processes as applied to calcareous iron ores. These reports all indicate that heat treatments prior to crushing may contribute materially to intergranular comminution, but they also indicate that no organized attempt has been made to determine the controlling factors of the method or to determine its applicability in general. The present article is a report on the initial phase of such an investigation. The authors have reviewed the claims of prior investigators and have attempted, also, to establish the factors that might determine the applicability of heat treatments in the mineral industry. In this work 2000-g samples of various rocks were heated in a small laboratory furnace and crushing and sizing operations were carried out in standard laboratory equipment. All samples of each rock were as nearly identical as possible in particle size, grain size, and composition and contained only lumps coarse enough to contain many grains each. Tests on Granite A number of tests were made on a coarse grained Finnish granite obtained in the form of coarse chips from a local monument yard. This rock exhibited little variation from piece to piece in either composition or grain size. The minerals contained were quartz, orthoclase, small amounts of hornblende, and minute quantities of mica. Grain size ranged from about 1 mm to about 3 mm. Temperature of the Heat Treatment: In some cases the granite was heated to a particular temperature and crushed, hot, immediately upon withdrawal from the furnace—in others the rock was allowed to cool before crushing, but without quenching to room temperature after heating. In most tests on granite the heating period was about 2 hr with the furnace at the highest temperature for about 1 hr. Cases in which these periods were varied greatly will be presented separately.

Jan 1, 1959

By K. Tanaka, J. D. Mote, J. E. Dorn

It) an attempt to ulLcoce.lP the operative strain-rate-contl-olliy: dislocation nieclzanistns, specially oviented sizgle clystals of the intel-nzediate 1zexagonal phase containing Ag plus 33 at. pct A1 were tested in tension over a wide range of temperatures. Slip was observed to take place by the {0001} <1120> {l100} mechani fracture took place across the(i100) plane and winning occurred by the (i01Z) ?lechanisn. Basal slip exhibited a strong yield point over the -alzge from 77 to 450°K, the upper ,esolved shear st]-ess having the exceptionally high value of 10,500 psi over this entire ?-a?zge of tenzpei,atuves. The critical 9-esolved shear stress for prismatic slip decreased f7-om 48,000 psi at 4.3"K to 23,000 psi at 170°K (Region 1) follozcirg zt:lzich it decl>eased sloz&ly to 21,500 psi at 475°K (Res&apos;on II); from 475" to 575°K (Regioz III), the c7-itical esolced shear stress dec&apos;-eased precipitously to 2000 psi; and from 575" to 750°K (Region IV) it decreased less afi&apos;dly to a low value of about 500 psi. Pvistintic slip in Region I was pobably controlled by the tliel-nally activated riecharzisui of nucleation and g,-ozcth of kinks in dislocations lying in Peierls potential troughs. In Region II for prismatic slip the critical 1-esolved shear stress was slzocn to be deteemined by sh0l.t-range 01-dering, Overall the forgiorz fo basal slip, 7.c.lre1-e a Strong yield-point phenorlienu ia7as observed, the critical vesolved slzea?-stress was shoztn to be determined by n conibirzation 0-f Szizuki locking and short-range-order Izavderzizg, The precipitous decrease in the critical resolved shear stress with increase in ter,/pe7-atrir-e over Region HI was tentatively ascribed to a decrease in the degree of slort-)ange 07-del;iqq (0)- clusteing) and also the effect of fluctuations the degree of o?der, It is at pgreser2t zrtzce)taitz as to 1t1hethe1- these or other possi1)le effects are also ,esponsible. fo- the data obsel-ved 172 Region IV. 1NTEREST in inter metallic compounds stems not only from their role in dispersion hardening of polyphase alloy ystems but equally from their potentialities for high strength, hardness, and stability not only at atmospheric temperatures but especially at elevated temperatures. As summarized in a re- cent symposium of the Electrochemical Society on "Mechanical Properties of Inter metallic Compound", most of the experimental evidence regarding the mechanical behavior of intermetallic compounds centers about the effect of temperature on the hardness and ductility of polycrystalline specimens. The available data reveal that the plastic behavior of intermetallic compounds might be rationalized in terms of the usual dislocation mechanisms appropriate to a solid solutions providing the additional complexities arising from crystal structure, long-range ordering, short-range ordering, and defect lattices are taken into consideration. It is apparent, however, in terms of the history on a solid solutions, that a complete detailed mechanistic rationalization of dislocation processes may not be possible until the deformation processes are studied in single crystals of intermetallic compounds. The present paper contains a preliminary report on the plastic behavior of single crystals of the hexagonal Ag-A1 intermetallic phase over a wide range of temperatures. The results confirm the thesis that single crystal data provide a most effective method of identifying operative dislocation mechanisms in intermetallic compounds. EXPERIMENTAL TECHNIQUES Several factors prompted the selection of the hexagonal Ag-A1 intermetallic phase for this preliminar investigation on the plastic properties of single crystals of intermetallic compounds: 1) This phase has a wide solubility range5 which would permit future investigations on the effect of composition and axial ratios on slip mechanisms. 2) Although it undoubtedly exhibits short-range ordering (or clustering) this intermetallic phase is free from complexities arising from long-range ordering.6 3) Since the atomic radii of aluminum and silver are practically identical, the possible complications due to Cottrell locking are minimized. 4)Whereas the dislocations on the basal planes are expected to dissociate into Shockley partials and are thus susceptible to Suzuki locking, those on the prismatic planes probably remain complete. 5) The axial ratio, being 1.61, is almost ideal, suggesting that short-range ordering may be almost spherically symmetrical. The present investigation was conducted exclusively with the hexagonal Ag-A1 alloy containing 33 at. pct Al. Preliminary investigations revealed that this alloy undergoes basal slip by the (0001)

Jan 1, 1962

By M. Bender, H. D. Goodfellow

"Belt conveyors, because of their low capital and operating costs and inherent flexibility in handling a wide variety of products, have numerous applications in the mining industry. With the increased emphasis on improving the workplace environment, dust emissions from belt conveyors at transfer and other points have become a significant problem for many applications.The paper reviews belt conveyor dust control technology and compares dust emissions from belt conveyor systems with other possible materials handling systems. Technical information on existing belt conveyor installations which have been designed to minimize or eliminate dust emissions is summarized. The important features of good design as well as the necessary input from operating and maintenance personnel are highlighted. The status of research and development work on dust control on belt conveyor systems on a world-wide basis is discussed. A proposed methodology for field measurements of emission factors from belt conveyor systems is presented. Data collected from such a test program can be used to establish a design basis for new dust-free belt conveyors.IntroductionBelt conveyors have proven themselves over the year s in the mining industry as a reliable and low-co st method of moving bulk material at high tonnage rates. Because of the above, as well as their inherent flexibility in handling a wide variety of products, belt conveyors are often called the ""workhorses"" of bulk material handling systems. For many processing plants, belt conveyors serve as the ""vital link"" in providing feed to the plant and the removal of the finished product. As would be expected, there has been considerable engineering effort in designing rugged and reliable belt conveyor systems.It is also well known that belt conveyor systems are a primary source of respirable dust emissions. The major sources are from the tail pulley where material is received and the head pulley where material is discharged, from the return idlers, due to the fine material attached to the return belt, and from the take-up pulley. It is usually considered that for dust particles less than 10 microns, these particles will remain airborne and thus can have a profound effect on the workplace environment."

Jan 1, 1982

By Edward H. Robie

MANY engineers currently are working harder than usual, in part because of the demands being made upon them for increased production in the war effort, and in part because engineers are in short supply so that there is more to do for those who are available. In some instances the question has come up as to what additional payment, if any, should properly be made for overtime work, and to what extent is such payment legally permitted under the rules of the Salary Stabilization Board. Most engineers employed in an executive, administrative, or outside salesman capacity are not paid for overtime. They are paid for doing a job instead of for the hours they put in, and are free to do all the unpaid home work they feel they should do after dinner, as well as to do a bit of thinking and worrying during wakeful hours in bed. But engineers employed in a strictly professional capacity normally have standard office hours. Some employers pay for overtime and some don&apos;t. Their practice as it existed on Jan. 25, 1951 may be continued. In the absence of such practice at that time, the Salary Stabilization Board says that the engineer may now receive additional compensation up to his straight-time rate for an extended work week. No formal approval of the Board is required. "A professional engineer," in the mind of the Board, "is a person employed in a professional capacity, who, by reason of his special knowledge of the mathematical and physical sciences and the principles and methods of engineering analysis and design, acquired by professional education and practical experience, is qualified to practice engineering." Inducted Engineers A letter from a young man just inducted into the service and fearful that his engineering training is not going to be utilized reached us last week from Fort Riley, Kan. The AIME, cooperating with the other Founder Societies, is doing what it can, through the Engineering Manpower Commission, to see that engineers are not being wasted in the military establishment. So we passed this young man&apos;s letter along to T. A. Marshall Jr., executive secretary of that Commission. Mr. Marshall tells us that the normal procedure now being followed by the Army in handling Selective Service inductees is to have them first complete 16 weeks of basic training before assignment to specialized categories. During that time they are classified. If one has an engineering degree he should be classified as scientific and engineering personnel at the end of the basic training period. He should then be transferred to a technical detachment and assigned an MOS (Military Occupational Specialties) reflecting his education. If a young man is looking for a commission, it might be added that the current policy in the Army is to transfer to OCS (Officer Candidate School) applicants for commissions in the infantry only. The local unit commander can give information on this point, based upon the aptitude shown in tests. In view of the current shortage of engineers, it is believed that a young man should seriously consider remaining in enlisted status unless he can obtain a commission in a branch of the service that can make full use of his technical skills. His work may thus be more to his liking even if his uniform is not. New York or Atlantic City? In which place is it preferable to hold the annual meeting of a professional society like the AIME-New York or Atlantic City? That is, a winter meeting. The American Society of Mechanical Engineers has held its recent meetings at Atlantic City, with increasing attendance and satisfaction. The American Institute of Chemical Engineers chose Atlantic City for its recent December annual meeting. Only twice has the AIME ever met in this famous seashore resort: in 1898 and 1904. Chief advantages of going to Atlantic City over New York are that technical sessions are better attended, there being little else to do there in Winter; the cost to members is less, and hotel facilities are better adapted to a large meeting. As to cost, take a banquet, for instance. The Chalfonte-Haddon Hall, which caters especially to winter conventions, charges about $7 for a banquet dinner of standard quality. A similar dinner at the Statler in New York would cost about $8.50, and about $10.50 at the Waldorf-Astoria. (The charge for tickets is somewhat greater than this because of the cost of music, programs, flowers, and other incidentals.) Rooms, too, are somewhat less at Atlantic City, two in a room being accommodated for about $7 per day each, compared with $7 to $10 at a headquarters hotel in New York. On the other hand, AIME headquarters is put to some extra expense for staff travel, food and lodging, and members living in New York and environs are forced to pay a hotel bill. New York, of course, has more varied amusements to offer, plenty of big stores, and the glamour of a big city. The choice may be discussed at the forthcoming meeting in New York. In the meantime if any reader cares to express an opinion one way or the other we shall be glad to hear it. College Scholarships Most parents of children in the later stages of high school find it necessary to give some consideration to the cost of a college education. Although salaries have increased substantially it seems that the cost of the necessities of life, such as food, housing, and maintaining the family automobile, have gone up even more, so the problem of financing the children&apos;s education is as acute as ever. Many overlook the possibility of scholarships, or feel that only a student at or near the top of his class has a chance to get one. As a matter of fact, close to 150,000 scholarships and more than 15,000 fellowships, worth some $42,000,000, are available in 1200 American colleges and universities. Those who are interested should send 55¢ to the Superintendent of Documents, Government Printing Office, Washington 25, D. C. for a 248-page publication just issued by the U. S. Office of Education, entitled "Scholarships and Fellowships Available at Institutions of Higher Education."

Jan 1, 1952

By L. A. Stewart

"SUMMARYThis paper describes most of the chrysotile-asbestos deposits of Arizona. Mining methods are discussed briefly and asbestos-mill flowsheets are incorporated. Arizona asbestos mines are the only sources on the American continents of naturally iron-free chrysotile spinning fiber that is so urgently needed for electric-cable coverings, especially on warships.Asbestos was first recorded and recognized in Arizona in 1872, and a minor amount was produced from a deposit in the Grand Canyon in 1900. The discovery of a deposit in the Salt River district in 1912 initiated intensive prospecting, and deposits soon were developed. The total Arizona production of all grades of asbestos through 1953 is estimated at 30,000 to 35,000 tons.The deposits of the Central Arizona region, which are discussed in this paper, are scattered over nearly 2,000 square miles. The annual output is relatively small, because the asbestos occurs in thin, discontinuous veins and only in areas where intrusions of diabase are adjacent to or crosscut certain favorable units of the pre-Cambrian Mescal limestone. This stratigraphic limitation, combined with other essential geologic conditions, tends to restrict the size of the deposits.Virtually all of the deposits are in rugged, mountainous country, and many of the mines are on steep canyon walls. In the average mine, production of 1 ton of commercial asbestos requires removal of 30 to 40 tons of waste rock. Mining and transportation costs consequently are high.Several deposits of chrysotile that is somewhat harsh and of moderate tensile strength have been known for years but were not worked because of the former exacting demands of industry for high-strength, soft asbestos. However, improve¬ments in asbestos fiberization and spinning technique, combined with the urgent need for asbestos, have made this so-called semisoft fiber desirable, and deposits of this type now are being exploited."

Jan 1, 1955

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

T. Rosenqvist—The most interesting point in this paper is the observed transfer of iron into the slag in the initial stage of the desulphurization process, after which the iron again is reduced to the metallic state. The authors interpret this observation as showing that the sulphur enters the slag as an iron-sulphur compound which subsequently is decomposed by the slag. The present writer has previously suggested the following equation for the desulphurization process: S + O2- ? S2- + O For equilibrium in the blast furnace the oxygen potential is defined by equilibrium with graphite and CO of 1 atm pressure: C + O ? CO [2] During the desulphurization process the reactions proceed in the direction of the arrows. If one assumes eq 2 to be significantly slower than eq 1, the transfer of sulphur into the slag, in accordance with eq 1, will build up a local oxygen potential at the metal-slag interface very much higher than that corresponding to the value defined by eq 2. This is possible because the equilibrium oxygen potential in eq 1 is high as long as the sulphur content in the slag is low. This oxygen potential will again be able to oxidize some iron: Fe + O ? Fe2+ + O2- and an increase in the iron content of the slag will be observed. Adding up eqs 1 and 3 one obtains: S + Fe ? S2- + Fe2+ The net effect is thus in harmony with the experimental observation but is obtained without assuming any close ties between the sulphur and iron atoms during the process. Furthermore, it follows from eqs 1 and 2 that when the sulphur content in the slag increases, and equilibrium with C and CO is finally approached, the local oxygen potential at the metal-slag interface will decrease, and the iron in the slag will be reduced back into its metallic state. C. E. Sims-—The data and conclusions presented in this paper are thoroughly convincing in establishing the mechanism of sulphur transfer from iron to slag as in a blast furnace. The evolution of gaseous CO in step 3 of the reactions given on p. 1112 makes the process virtually irreversible. Assuming that the process is similar in slag-metal systems other than in the blast furnace, it is readily seen why free CaO and re-ducing conditions so greatly favor desulphurization. On the other hand, the very effective desulphurization obtained in oxidizing slags when strongly basic, must be attributed to the relatively high stability of CaS as compared to FeS. The ease and simplicity with which the reactions of classic chemistry agree with the experimental data and explain the mechanism is noteworthy. The concept of molecules of FeS, soluble in both phases (metallic iron is not soluble in the slag), migrating from the iron to the slag and there reacting with CaO, which is soluble only in the slag phase, is clear and uncomplicated. This is likewise true for step 3. Those who would deny the existence of molecules or molecular-type combinations in liquid iron, must strain to provide a mechanism so lucid. In the absence of molecules, the Fe and S exhibit a remarkable collusion. L. S. Darken—The investigation and interpretation of rate phenomena in the range of steelmaking temperatures is a difficult task. Most of the laboratory investigations of steelmaking reactions have been concerned with equilibrium. Having determined the equilibrium, our attention naturally focuses next on the mechanism and rate of approach to equilibrium. The authors seem to have contributed substantially to our understanding of these factors for the case of sulphur transfer. I should like to ask the authors whether they consider that the sulphur transfer reaction is diffusion controlled as many high-temperature reactions seem to be. If so, it would seem reasonable to suppose that the slow diffusion step of the process is the transfer across a pseudo-static layer or film similar to that considered in heat flow problems. As the diffusivity and fluidity are smaller for the slag than for the metal, it may tentatively be assumed that the sulphur gradient exists in a thin layer in the slag adjacent to the slag-metal interface and that the metal and the main mass of slag are each maintained uniform by convection. On this basis the amount of sulphur transferred across unit area per unit time is D p (?S%)/100 ?1, where D is the diffusivity, p the density, (?S%) the difference in percent sulphur on the two sides of the layer, and ?l is the layer thickness. At the beginning of the experiment the main body of the slag and hence one side of the layer contains no sulphur; therefore (?S%) may be replaced by (S%), the sulphur content of the slag at the slag-metal interface, which in turn is equal to L[S%] where [S%] is the sulphur content of the metal and L is the distribution coefficient. The rate of transfer thus becomes DpL[S%]/100 ?l, which the authors designate K[S%]. Equating these two quantities and setting D = 10-6 cm2 per sec, p = 3 g per cm3, L = 40, and K = lo-+ g cm-2 sec-1, it is found that ?l, the film thickness, is about 0.01 cm—a value of the order of magnitude of that found in heat transfer problems in liquids. The uncertainty of the numerical values used leaves much to be desired, but at least it can be said that this calculation tends to support the proposed model involving diffusion through a film. Although this does not seem to affect the general argument, I should like to call attention to the fact that the diffusivity3 of sulphur in hot metal is found (on conversion of units) to be about 10-4 cm2 per sec rather than 104 cm2 per sec as stated by the authors. The three equations written by the authors to express the steps in the overall process of sulphur transfer may alternatively be written ionically as only two Fe + S = Fe++ + S-- Fe++ + O-- + C (graphite or metal) = CO (gas) + Fe where the underscore is used to designate the metallic phase; ionic species are slag constituents. After the authors have so neatly demonstrated that iron and sulphur transfer together (at least initially), this fact seems almost self evident; from eq 4 it is seen that if sulphur acquires a negative charge during transfer

Jan 1, 1951

By F. B. Micheli

Introduction and Historical Background Although both sluices and devices using a simple flowing film are among the oldest known methods of concentration, they are still extensively employed in a variety of situations where they are not only efficient but also have a low operating cost. More particularly, they are valuable for making a bulk concentrate of high specific gravity minerals from alluvials and beach sand and in upgrading low value ores. The antiquity of some of the basic concepts can be seen in the fact that they are found in Agricola&apos;s De Re Metallica which was published at Basic in 1546, while in 1602, Carew114 gave a fairly clear account of tin ore dressing in Cornwall in which grass turves were used to catch the cassiterite. This is perhaps one of the forerun¬ners of the blanket strake. The planilla of Mexico, the lanchute of Malaysia and the old Cornish box huddle are also descended from these early strakes. In fact, a sluice with holes in the bottom for concentrate discharge was shown by Agricola, not unlike the principle used in the plane table and the lamflo sluice. The convex huddle for the treatment of sandy material appeared about 1848 and the concave form eight to ten years later. For the treatment of "slime" sizes, wooden dead frames which were manually worked were used in Cornwall for many years but gave way to the automatic water-operated type in 1860, and even a mechanically tilted frame was described by Henderson115 in 1857. Theoretical Considerations The behavior of solid particles in suspension depends to a great extent upon the pulp density and the size of the suspended particles. In a fairly dilute suspension, such as that normally used when dealing with small particle sizes, the behavior of particles in a flowing film results from two effects. These are the lateral displacement, which is determined by the time taken for each particle to penetrate the flowing film and reach the solid surface, and the resistance offered by each particle to further displacement after it has reached this surface. The initial penetration through the flowing film depends on the size and specific gravity of the particle and the thickness and viscosity of the film. As a result the smaller particles will migrate further before their movement is retarded relative to larger particles of the same specific gravity. The behavior on reaching the solid surface de¬pends on whether there is a single particle layer or, as is more often the case, a multiple layer or thin bed of material which is sufficiently dilated, permits the penetration of higher specific gravity grains. Thrust of Flowing Films. When a thin film of liquid flows over a plane solid surface, the layer next to the surface remains at rest but the velocity of the film increases with the distance from the surface and becomes a maximum near, but not quite at, the free surface. Therefore, a particle in suspension in such a film is acted upon by a greater force near the upper part of the film than at the lower part, resulting in an overturning effect. After a particle reaches the separating surface or an accumulated densely packed bed of other particles, the liquid flow causes it to move downstream by rolling, sliding, or by a movement involving alternating suspension and deposi¬tion (saltation). In rolling and sliding, which are brought about by a substantially noneddying stream, the large submerged particles are acted upon to the greatest extent and they move more rapidly than smaller ones, notwithstanding their greater mass. When two particles of the same size but of different specific gravity are considered, the higher density one moves more slowly by reason of its greater mass. As a result the particles tend to become arranged in the manner shown in Fig. 29. If any particle is so large as to stand above the water surface, the transporting force on such a particle, although the maximum available, will have less effect than that on a submerged one. Such large grains are therefore carried a lesser distance. Fluid Velocity in a Flowing Film It can be shown that if y is the distance within the flowing film from the interface, the fluid velocity v&apos; at this point is given by IIn this equation p&apos; is the fluid density, g the acceleration due to gravity, a the viscosity, a the angle made by the film to the horizontal, and 0 the film thickness (in cgs units). Furthermore, the volume Q of fluid per unit time and unit width is related to the film thickness and can be calculated, since which upon integration gives Consideration of the velocity v&apos; (Eq. l) and the terminal falling velocity µ derived from Stokes law enables the horizontal travel dz in time dt to be calculated as follows: Again integration and substitution of Eq. 3 gives In this equation, z is the distance travelled before a particle of density p at the top of the fluid film has settled on the concentrating surface. From these expressions it can be seen that the depth of the flowing film varies as the cube root of the volume of fluid and inversely as the cube root of the sine of the plane surface inclination. Similarly the distance a particle can travel before reaching the surface is directly related to the quantity of fluid and its viscosity. The viscosity of pulp is greatly increased by the presence of near colloidal particles and consequently the penetration of the somewhat larger grains through the flowing film is retarded, resulting in them being carried further across the concentrator deck. Obviously this effect can be reduced by prior removal of some or all of the ultra¬fine particles. The detrimental effect of kaolin on the separation of fine ferrosilicon from quartz has also been demonstrated by Johnston,116 the relationship between viscosity and efficiency of separation being shown in Fig. 30. The pH value also has a very marked influence on the viscosity and the use of acid mine water, for example, can increase viscosity materially and thereby result in lower efficiency. Dilation of Multi-Particle Systems Except during washing (when a single particle layer may exist) concentration usually takes place in a bed of particles which is partially dilated or stirred by eddy currents. The extreme case is seen in various forms of the sluice where the flow of water and the presence of riffles combine to create a fluid bed in which reverse classification takes

Jan 1, 1985

By M. C. Young

The Ozark Lead Company operating facilities are located in Reynolds County at the south end of the "New Lead Belt" of southeast Missouri. Development of this wholly owned subsidiary of Kennecott Copper Corporation started in May 1964, with a six foot diameter by 1250 foot deep steel lined shaft. A 20-foot diameter, 1480 foot deep, concrete lined production shaft was collared in March 1968. Production from the facilities began that month and has reached a present average of approximately 6000 tons per day. This paper describes the fully automatic system used to hoist the lead-zinc ore from a depth of 1420 feet to the surface crushing and concentrating facilities, as well as the man and materials hoisting system which allows the lowering of large equipment into the mine. MINING Galena, sphalerite and some copper minerals are found in several mineralized zones of the relatively flat lying dolomite of the "Bonneterre formation". The mineable ore zones vary from 10 feet to 110 feet thick, from 200 feet to 1200 feet wide, and several thousands of feet on strike. Two-boom diesel-hydraulic jumbos equipped with rotary percussion drills are used for both development and production. The jumbos are used principally to open the room and pillar stopes, and a track mounted down hole drill is used to bench in areas where the ore thickness warrants. Explosives used are ANFO, dynamite, and water-jell which are detonated by electric blasting caps. Broken ore (and waste) is hauled by rubber tired transports to chuted transfer raises. Approximately every 500 feet, these raises connect the orebody to the 15 foot by 28 foot train haulage level below. Two twenty ton diesel-hydraulic locomotives haul 160 tons per trip to storage pockets which have a total storage of 1500 tons of material. From these pockets the material is fed over a 6 foot by 20 foot mechanical vibrating feeder which includes one 4 foot by 6 foot grizzly section to a 48 inch by 60 inch jaw crusher. The grizzly and crusher products are transferred horizontally 45 feet on a 48 inch conveyor belt, to one of two 750 ton storage pockets. The material can flow unimpeded into one pocket or be deflected into the second pocket by a movable rock filled hopper. LOADING POCKET The material in each storage pocket is controlled by six lengths of ship anchor chain. The upper end of the 20 foot chain is anchored behind a brow in the pocket, while the lower ends are joined by a header. The lower ends of the chains are raised by two electric powered marine barge winches using 3/4 inch cables. The crusher operator controls the raising and lowering of these chains from a console in the crusher station. The upper and lower travel of the chains is controlled by cams and limit switches on the winch drive. The winches are located within 5 feet of the shaft, and shaft water continually causes problems with the electric components. A malfunction in the upper travel switch will cause the winch to pull the chains into the cable sheave (breaking the cable or sheave), while the lower switch causes the winch to unwind too much rope. Serious consideration is being given to changing the electric motors to pneumatic, since pneumatic motors will eliminate the need of any electrics at the winches. The winches can be "bottomed out" against the sheave without damage to the cable or sheave when raising the chains, and can be allowed to "free wheel" to lower the chains. A surge pocket below the two storage pockets feeds material to two

Jan 1, 1975

By D. Cronan

Both the S.W. Pacific and Caribbean island arcs are host to submarine hydrothermal mineral deposits, the former more so than the latter. In the S.W. Pacific, the whole gamut of such deposits occurs on the sea floor, sulphides, silicates, Fe and Mn oxides, whereas during the limited studies to date in the Caribbean only hydrothermal Mn oxides have been found to occur on the sea floor there in any abundance. Comparisons between the respective Mn oxides are thus the most fruitful at the present limited stage of our knowledge of Caribbean submarine hydrothermalism. Southwest Pacific hydrothermal ferromanganese oxides and their possible relationships to underlying sulphide mineralisation or bedrock anomalies (the geochemical window concept) has been a fruitful field of research in recent years. Certain element enrichments in crusts may be indicative of underlying mineralisation (Rogers et al. 2001). However, more recent studies on buried hydrothermal crusts in the S.W. Pacific (Cronan et al. 2002) have shown there to have been postdepositional diagenetic uptake of elements in them, and thus these elements cannot be used with any confidence in geochemical exploration for buried sulphide deposits unless sediment involvement can be eliminated.

Jan 1, 2002

To provide accurate navigation and control of continuous mining systems.

Jan 1, 1996

Improve the dust-control effectiveness of the Rotoclone1 dry-dust collection systems used on surface coal mine drills.

Jan 1, 1994

Provide ear protection which allows normal, unaltered hearing at low noise levels, yet protects the wearer from excessive noise. The Approach: A commercial earmuff with integral speakers has been adapted with a special amplifier so that low sound levels are not affected, but higher noise levels automatically reduced. How It Works: The discriminating earmuff developed by the Bureau of Mines has two modes of operation. In the first mode, where input sound levels are less than or equal to 83 decibels, the wearer will perceive noise at its actual level. For inputs in excess of 83 decibels, the second mode of operation begins and the sound perceived by the wearer is progressively attenuated, or reduced, as the input level is increased. A recognizable difference still exists between two adjacent sound levels enabling the wearer to differentiate between them.

Jan 1, 1975

Reduce the dust exposure of workers in large processing buildings at mineral processing operations, using a total mill ventilation system.

Jan 1, 1994

Joy Mining Machinery has developed a new generation of ultra-heavy shields for Anglo Coal Australia?s Moranbah North Mine in the Bowen Basin region of Queensland. These are the largest two-leg longwall shields yet produced by Joy. The longwall at Moranbah North posed a number of challenges, among them a thick, inclined, sometimes undulating seam with increasing depths. The result of this was the need for larger, heavier, more powerful roof supports. Knowing that future longwalls would be going deeper, mine management determined that shield yield ratings of at least 1,750 t (1,930 st) was what would be required.

Jan 1, 2011

The 1980&apos;s has seen a revolution in gold mine developments in Australasia. We are witnessing a near ten fold increase in all phases of gold mine development ranging from exploration through to production. Ounces produced have yet to mirror this activity but have increased from a low point of 600,000 in 1980 to 1,800,000 ounces in 1985 in Australia alone. In 1980 the region had less than 10 gold mines, yet 14 new mines started production in 1985 and as many as 28 might start up in the next 18 months. Developing gold mines in the 1980&apos;s in Australasia has not been the easiest of jobs, yet the great number of new developments suggest that we have entered a new era that will rank in history along with the 1850&apos;s, 1890&apos;s and 1930&apos;s as high points.Within- our Australasian region two very distinctive styles of gold deposits are being discovered and developed. On the Australian mainland the new mines are in geologically old rocks and are mostly the low grade haloes surrounding worked out high grade deposits. Most are limited life open pit operations depending on recent innovative carbon-in-pulp or heap leach technology. In the Pacific Islands region the style of discoveries are geologically recent, large bulk-low-grade deposits often geologically analogous to porphyry copper deposits. These deposits are typically 10 to 20 times larger than those being proved up on the Australian mainland.

Jan 1, 1986

By Gus Van Weert

Capital and operating costs for a 10 tpd NITROX PROCESS® demonstration plant are being established. The flowsheet is based on Dickenson Mines' Red Lake arsenopyrite concentrate. Support for the work comes from a number of mining companies, engineering firms and the Northern Development fund of the Ontario Government.

Jan 1, 1998

By Janet Briggs, Michael Vertes, J. F. Shirley, R. E. Cutherbertson, Alexander Sutulov, J. D. Vincent

Molybdenum, the 42nd element in the periodic table, is now a very important metal in industry of the United States and the world. Chemically, it is a very complex element having valences of 0, +2, +3, +4, +5, and +6, and can exist as a mixture of these valences in many compounds. It was discovered and separated by Scheele in 1778. At the present time, molybdenite is the only important commercial source of molybdenum. Molybdenum occurs in some 12-14 different recognized minerals, but practically all of the production is molybde¬nite. As late as 1915, it was considered a rare metal. Germany made extensive use of it during the first world war, which resulted in the establishment of a molybdenum industry in the United States at Cli¬max, Colo., in 1917. It was not until the mid-1920s when its introduc¬tion to steelmaking developed for the US automobile industry that the metal gained important acceptance. The first notable example was the Willis-Saint Claire automobile introduced in 1921 by C. Harold Willis, formerly chief metallurgist for Ford Motor Co. Molybdenum has also been produced in the US from wulfenite (PbMoO4) from Arizona and Nevada, and a small amount of molyb¬denite was produced up to the early 1970s as a coproduct from two molybdenum-bismuth mines in Quebec, Canada. The first byproduct production of molybdenite from a copper con¬centrating operation was achieved at Cananea, Sonora, Mexico, and was followed shortly by the Utah division of Kennecott Copper Co. in the mid-1930s. This section of the Handbook is divided into four main subdivi¬sions: Chapter 2, which deals with the concentration and recovery of molybdenite from primary molybdenum ores; Chapter 3, which deals with the recovery and separation of molybdenite as a byproduct from copper production; Chapter 4, which deals with the conversion practices employed for converting molybdenite to the oxide or metal form; and Chapter 5, which covers the uses of molybdenum in steel¬making, chemicals, pigments, lubrication, catalysts, etc. The Free World molybdenum production, for the past eight years from all sources, is shown in Table 1. 2. Primary Molybdenite Ores General In 1978 there were four primary molybdenite recovery plants in operation: Climax, Endako, Quests, and Henderson. For over half a century the Climax mine has been by far the largest and most important source of molybdenum in the world. During the early 1970s a few primary molybdenite mines in British Columbia were shut down due to adverse economical conditions. The largest of these was British Columbia Molybdenum, formerly owned by Kennecott Copper Corp. The Climax plant and operations are described in detail and the opera¬tions of the other plants are summarized following a general discussion of recovering molybdenite from primarily molybdenite ores. Recovery of molybdenite as a byproduct will be taken up in Chapter 3. Geology Molybdenite production from mines in which it is the sole or preponderantly important mineral is confined to ore bodies in granites or igneous-type rocks or pegmatite dikes. More than one state of mineralization is sometimes present as at Climax, York, Hardy, and Questa. In the case of Climax, important byproduct recovery of tung¬sten, tin, pyrite, and monazite is also attained. The molybdenite is commonly disseminated very finely in the siliceous granitic gangue but does occur in stringers and blebs in silicified mineral solution channels. Treatment of Primary Molybdenum Ores All primary molybdenum ores are low in molybdenite, the content ranging from less than 0.20% MoS2 to a high of about 0.40% MoS2. A high resistance to grinding is a common ore characteristic combined with a habit of very fine dissemination of molybdenite in the siliceous portion of the ore. Because of low MoS2 content, high hardness, and fine dissemination of molybdenite in primary ores, it has been necessary to develop a special mill treatment which yields an accepta¬ble recovery at a reasonably low milling cost. Essentially, the treatment consists of intensive rougher flotation at a relatively coarse grind with production of a low grade rougher froth containing a substantial percentage of the molybdenite in the form of low grade middling particles, followed with multistage regrinding to liberate the locked molybdenite and several stages of flotation cleaning to produce a finished grade concentrate. Rougher flotation on pulp ground to ap¬proximately 35 mesh is common practice, even if complete liberation of molybdenite from siliceous gangue may require regrinding to 200 mesh. The flotation of middling particles with a very low content of molybdenite in the rougher flotation step has been made possible by the application of a combination of reagents first developed at the Climax mill and used with some modification in the treatment of all primary molybdenum ores. The reagents commonly employed and their functions in flotation are:

Jan 1, 1985

By E. Miller, J. M. Eldridge, K. L. Komarek

Aclivilies of magnesium in liquid Alg-Si alloys have been delermined between 5 and 60 at. pcl Si, close to the melling point of Mg2Si, by an improved isopieslic melhod. Silicon specinrens, held in alumina crucibles and graplrile conlainevs of special design, were healed in a letrlpevalure gvadient and equilibrated with mag-nesilcrrl rapor in a closed lilanium system. The ther-madynamic Junctions were calculated and compared with the thermodyuamic properties of the other three mg- Gvoup IVB systems. Lattice paramelers of three Mg2X compounds were measured. The bonding in the Mg2X compounds is largely covalent with small and uarying amounts of metallic and ionic conlvibutions. The Mg-Si phase diagram1 has one congruent melting compound, Mg2Si, of essentially stoichiometric composition, two eutectics, and very limited terminal solid solubilities. Little information is available on the thermodynamic properties of this system. The free energy of formation of Mg2Si has been determined by the Knudsen cell technique2 in the range 572" to 680oC, by the transportation method3 between 858" and 950oC, and by the electromotive-force method4 in the range 400o to 600°C. Kubaschewski and villa5 and caulfield6 have measured the heat of formation of Mg2Si. An electromotive-force study of magnesium-rich liquid alloys was recently published by Sryvalin el al.7 The present investigation was undertaken to complete a general survey of the thermodynamic properties of the homologous series of Mg-Group IVB systems, i.e., Mg-Pb,a9,Mg-Sn,10,11 mg-Ge,12and Mg-Si. An isopiestic technique, previously developed for similar measurements on liquid Mg-sn11 and Mg-Ge alloys,12 was modified for the Mg-Si system. Specimens of the nonvolatile component, silicon, were contained in dense alumina crucibles placed inside covered graphite crucibles which were heated in a temperature gradient in an evacuated and sealed titanium reaction tube and equilibrated with magnesium vapor of known vapor pressure. The alumina crucibles prevented contact between the highly corrosive liquid Mg-Si alloys and graphite. The graphite cruci- bles effectively preserved the high-temperature equilibrium composition of the liquid alloys containing highly volatile magnesium on termination of the experiments during the quench to room temperature. EXPERIMENTAL PROCEDURE Silicon of semiconductor-grade purity (E. I. du Pont de Nemours and Co., Brevard, N.C.) and 99.99+ pct Mg (Dominion Magnesium Ltd., Toronto, Canada) were used. Graphite crucibles with press-fitted lids were machined from high-density (1.92 g per cu cm) rods (Basic Carbon Corp., Sanborn, N.Y.) which had a maximum ash content of less than 0.04 pct. The alumina crucibles had a purity of 99.7+ pct (Triangle RR grade, Morganite, Inc., Long Island City, N.Y.). In preliminary runs the liquid alloys were contained in graphite crucibles following the exact procedure developed for the Mg-Ge system.&apos;2 These runs failed due to appreciable reaction between the molten Mg-Si alloys and graphite, and the results have been discarded. The procedure was then modified and the Mg-Si alloys were subsequently held in alumina crucibles. For most of the runs alumina crucibles of known weight and approximately 6.3 mm ID, 12.5 mm height, 1.0 mm wall thickness were loaded with weighed amounts of silicon and encapsuled in tightly covered weighed graphite crucibles 5/16 in. ID, 2 in. helght, 3/32 in. wall thickness). The graphite crucibles were machined from rods which were 85 pct of the theoretical density. These crucibles were therefore sufficiently porous so as to permit magnesium vapor to effuse through the silicon under the experimental conditions of approximately 970O to 1220°C and 1 day equilibration time. However, negligible magnesium was lost from the crucible during the quench due to the slow effusion rate through the pores of the graphite. The inner alumina crucible prevented the liquid alloys from contacting the graphite, and the very tightly fitting graphite crucible lids served to retain any magnesium vaporizing from the alloys inside the crucibles during the quenching step.12 The loaded silicon-alumina-graphite cells were positioned, one above another, on a 16-in.-long titanium thermocouple well and tied securely to the titanium tube with thin molybdenum wires held in grooves around the circumference of the graphite crucibles. A thin (0.005-in.) molybdenum strip prevented contact between the graphite crucibles and the titanium. This assembly was lowered into a titanium reaction tube (la in. ID, 16 in. long, $ in. wall thickness) closed on one end which contained a 11/2-in.-long cylinder of magnesium at the bottom. The inner titanium thermocouple well was positioned eccentrically in the large tube because of the eccentric mounting of the cells on the well. Appropriate modifications of the titanium cap"&apos;12 were made to join the inner and outer titanium

Jan 1, 1968