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By H. S. Drinker

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

By J. F. Enrietto

It is shown that one cannot approximate the broad internal friction Peaks observed in the Fe-Mn-N system as the sum of a number of subsidiary Peaks and still vetain the concept of a single relaxation time. Furthermore, the approximation of manganese affectirg only nearest neighbor interstitial sites is shown to be incompatible with the experimental results. The results of this experiment show that the influence of manganese in the ferrite lattice extends over 3 to 4 atomic distances. NUMEROUS experiments have shown that the presence of relatively small amounts of substitutional alloying elements can profoundly affect the profile of the damping spectrum in the vicinity of the Snoek peak in bee metals.1-7 Dijkstra and Sladekl observed this phenomenon in Fe-Mn-N alloys where it appeared as a broadening and shifting of the Snoek peak to higher temperatures. The authors interpreted this as being the result of a superposition of two damping maxima—the normal Snoek peak, located at its usual position and having a normal half-width, and an interaction peak having a normal half-width, but occurring at a somewhat higher temperature. They prclposed that the interaction peak was caused by nitrogen atoms jumping from Fe-Mn sites to Fe-Fe sites, the Fe-Mn interstitial sites having one manganese and one iron atom as nearest neighbors, and the Fe-Fe sites having two iron atoms as nearest neighbors. They further suggested that the Fe-Mn sites had a lower energy than the normal Fe-Fe sites because of the presence of the adjacent manganese atoms. From the relative heights of the two postulated damping peaks, the dif-

Jan 1, 1962

By A. G. Taylor, H. P. Ehrlinger, U. C. Tainton

FOR the last 15 years lead has been the standard material for anodes in electrolytic zinc production and it has been generally accepted that this lead should be as free as possible from impurities. Laistl says, "Next in importance to the purity of solution is the purity of the electrodes and tank lining material. The purest lead obtainable should be used for anodes." The same conclusion is expressed by Ingalls2 and by two German investigators who say that experiments have demonstrated that the purer the lead, the greater is its resistance to corrosion during electrolysis.3 Pure lead, however, while the best material hitherto available for anodes, is not entirely free from objection. The plain lead anode gradually disintegrates under electrolysis, so that the life of an anode, say 1/4 in. thick, is frequently not more than two years. Part of the lead from the anode finds its way to the cathode, lowering the purity of the deposited zinc and decreasing the hydrogen overvoltage. The rest of the lead that comes from the disintegration of the anode goes into the manganese dioxide which is precipitated in the cells, rendering unsaleable what might otherwise be a valuable by-product. Yet a further disadvantage of plain lead anodes comes from their tendency to bend or buckle during electrolysis, as a result of intercrystalline oxidation. This makes it necessary to use a fairly aide spacing between anode and cathode to avoid short circuits, and thus leads to a higher power consurnption than would otherwise be necessary. The power consumption is also affected by the high decomposition voltage at a lead peroxide surface.

Jan 1, 1929

By Carville E. Sparks, Rollin Farmin

TRIAL installations of rock bolts, of the slit-rod-and-wedge type, were under way at several units of Day Mines, Inc., when Korean hostilities interrupted the already slow deliveries of steel bars to the Coeur d'Alene district. Factory-made bolts had not yet been put on the local market, so the program was halted for lack of supplies. Interest was revived by a visitor's description of wooden roof bolts. These were said to have been used briefly with apparent success in a coal mine, until apprehension voiced by the U. S. Bureau of Mines caused the practice to be suspended. To make wooden bolts for trial in ground support, Day Mines acquired a second-hand doweling machine equipped with two cutting heads, one to turn out the desired round rods of 2-in. diam, the other to turn out 1-in. rods to be used as powder-tamping sticks. This machine was installed in the all-weather sawmill of the Hercules mine unit at Burke, Idaho, where fabrication of the wooden bolts commenced early in 1951. Most of the mining in the Coeur d'Alene district is along steeply dipping veins in shaly quartzite and argillite of Algonkian age. Ground support commonly is required in zones where the rocks have been sheared, brecciated, and hydrothermally altered. Pressure from the sidewalls is more troublesome than weight overhead, but both increase with the size of the mine opening. Caving may come from a progressive sloughing of irregular rock fragments or from an exfoliation and buckling of the layered wall rocks. The disintegration is thought to develop from an initial elastic expansion of the rock toward the newly-created mine opening, followed by the dilation of many tiny partings in the rock by absorption of water. As the partings widen, masses of rock develop weight and become free to fall. The function of rock bolts is to prevent or retard widening of partings in the rock supported. Wooden Bolts, Wedges and Headboards Bolt assembly used by Day Mines consists of a bolt 4 or 6 ft long, two wedges 16 in. long, and a headboard 30 in. long, Fig. 1. All four pieces are made of local red (Douglas) fir, either green or well-soaked in the mill pond before it enters the sawmill. Bolts are fabricated from cants, 2 1/4 in. sq, cut from relatively straight-grained timber with a minimum of knots and trimmed to 4- and 6-ft lengths. The bolt then is turned in the doweling machine from 21/4 in. sq to 2 in. diam round, except for a 4-in. length at one end which is left full square to provide the striking head and the shoulder that holds the headboard in position for wedging. The foot end of the bolt is slit with a thin saw for a length of about 16 in., thereby making a slot to receive the wedge against which the bolt is driven for anchorage at the bottom of the rock hole. A similar slit, 12 in. long, is made in the opposite (head) end of the bolt to receive the second wedge, which crowds the headboard against the ground at the collar of the rock hole and puts the bolt in tension. The second slot is aligned 90" from the plane of the first slot to avoid Longitudinal splitting and is notched out slightly to allow easier insertion of the collar wedge after the bolt has been driven to bottom. To prevent splitting the headboard by spreading action of the head wedge, this slot is oriented at 90" to the grain of the headboard when the pieces are assembled, Fig. 2. The wedges are similar to standard mine wedges, but more slender; they are cut 1 7/8 in. wide and 1 in. thick at the heel and taper out in 16 in. of length. The headboard, or bearing plate, is not necessary for some types of ground but generally is desirable because it helps the bolt to support an area of loose, friable rock and reduces the tendency for the rock at the collar of the hole to split away from the wedged head by distributing the pressure over a wider rock surface. The headboard may be a 24-to 30-in. length of 3-in. plank, 8 to 12 in. wide, but a similar length of rounded sawmill slab serves equally well at 20 pct of the cost. A hole of 2-in. diam is bored or punched through the center of the headboard, either at 90" or at various high angles to its surface. The bolt is inserted to its shoulder through this hole, then driven into the rock hole. Bolts, wedges, and headboards are given a full timber preservative treatment to inhibit rot. Bundles of each are immersed in a warm saturated solution of Osmose salts in water for 48 hr, removed, dripped dry, and stored in a relatively humid underground depot to cure. Most wooden rock bolts used by Day Mines are 4 ft long. Holes to receive them, about 42 in deep and 2 1/8 in. in diam, are drilled into the rock' to be supported, nearly normal to the periphery of the mine opening. The type of drill used is dictated by convenience: stoper, jackleg, or jumbo-mounted drifter. Correct depth of the hole is assured by use of a measuring stick that has been cut to the proper distance from drill chuck to the ground at the collar of the hole when a standard length drill rod is at the bottom. The bolt is seated to the shoulder through the hole in the headboard, the foot-end wedge is placed in its slot, and the assembly is inserted into the rock hole. Then the bolt is driven until it is seated solidly on the wedge against the bottom of the rock hole. Driving may be by hand with a sledge, or

Jan 1, 1954

By M. Cohen, R. J. Teitel

There is considerable interest in beryllium because of its low density (1.84 g per cu cm), high modulus of elasticity (40 X 106 psi), high melting point (1280°C), and special nuclear characteristics. Moreover, it has fair resistance to corrosion and oxidation. However, the purest beryllium known to the authors is relatively brittle at room temperature, and cannot be fabricated readily. The search for improved properties leads naturally to investigations of beryllium alloys and the phase relationships upon which they are based. The present paper is concerned with the alloys of beryllium and iron, primarily from the standpoint of the equilibrium diagram. Iron was selected (1) because of its all-round importance in metallurgy, and (2) because the beryllium-iron system is known to be quite complex. The most recent diagram in the literature is that of Gaev and Sokolov,' and covers only the iron-rich end up to 16 wt pct beryllium. Beryllium markedly restricts the austenitic field, closing the gamma loop at about 0.4 wt pct (3 at. pct). On the other hand, beryllium is soluble in a-ferrite up to a maximum of 7.5 wt pct (34 at. pct) at the eutectic temperature of 1160"C. The eutectic composition lies at 10 wt pct beryllium (42 at. pct) and corresponds to a mixture of or and FeBez (called j3 in the present paper). The solid solubility of beryllium in the a phase decreases with decreasing temperature, and makes precipitation-hardening possible. These findings substantiate the earlier work of Oester-held,2 Wever and Mueller3 and Laissus.4 In addition to the FeBe2 (ß) mentioned above, Misch6 has reported on FeBes, and indicated evidence for another compound containing still more beryllium. A preliminary diagram by Gordon,5 not hitherto published, is given in Fig 1. Besides the phases previously named, there are shown a solid solution of extended range (e), a compound of limited solubility (t), and a high temperature phase (7). Beryllium is designated as 8. The t field contains the compound composition FeBea, and the f phase undoubtedly corresponds to the beryllium-rich compound found by Misch.6 The region around the { phase was proposed by R. J. Teitel based on Gordon's data. The details of the diagram are self-explanatory. It was the unusual complexity of this system that prompted the authors to undertake the present studies. By the same token, more emphasis was placed on the beryllium-rich than the iron-rich end of the diagram. Exprimental Details RAW MATERIALS The beryllium employed in this investigation was one of the purest grades available. Its chemical analysis after vacuum melting is given in Table 1. The metal assayed 99.4 wt pct beryllium. Vacuum melted electrolytic iron of purity shown in Table 1 was the source of iron.

Jan 1, 1950

Jan 1, 1902

By A. W. Bethune, L. M. Pidgeon

The equilibrium vapor pressure of zinc has been determined over the systems: ZnS(s) + Fe(s) = FeS(s) + Zn (vapor) and ZnS(s) + 2Cu(s) = Cu2S(s) + Zn (vapor) by reacting the components in an evacuated tube containing a thin copper fiber, and allowing equilibrium to be established with respect to the brass formed and the zinc in the vapor phase. The composition of the brass formed was determined and equilibrium vapor pressure values obtained from existing data. Between 850' and 1000°C, values for the first reaction ranged from 8 to 58 mm Hg; for the second reaction from 1.8 to 22 mm Hg. THE direct reduction of metallic sulphides may be indicated by an equation of the form MS + X = XS + M [1] where X is some suitable reducing agent. In the case where M is a metal of relatively high volatility, the use of vacuum will displace the equilibrium to the right at comparatively low temperatures in the face of the usually unfavorable thermochemistry. For this technique to be feasible, MS, X, and XS, must of course be nonvolatile compared to M. In this regard, zinc becomes an obvious metal both from the point of view of its volatility and the fact that most commercial ore deposits of this metal occur as the sulphide, and its direct reduction would avoid the expensive roasting process necessary prior to present day reduction methods. Two reducing agents are of interest, namely copper and iron, with the former having the added attraction of being relatively easily regenerated by oxidation in the copper converter. The possibility of these reactions has been considered before. Imbert1 patented a process whereby zinc sulphide was reduced by either copper or iron. The process as described took place at atmospheric pressure and, at the high temperatures required, the system was molten, resulting in the formation of a matte with the consequent reduction in the activities of the reactants. Peterson2 did further experimental work on these reactions and, although promising results were obtained on a small scale, attempts to increase the size of the furnace led to excessive "blue powder" formation. The early workers failed to realize the large reduction in temperature afforded by the use of vacuum and its desirable results. More recently, the direct reduction of zinc sulphide by iron has been the subject of a thermody-namic investigation by Kelley,3 and Gross and War-rington' have studied the kinetics of the reaction in the temperature range 900" to 1000°C, their experiments being conducted using vacuum and indicating that the reaction would proceed practically to completion in 1 to 3 hr, depending on the temperature. A qualitative examination of the system ZnS-Cu carried out at this University indicated that 98 pct of the theoretical zinc in the charge could be recovered in 1 hr at 1000°C when operating under reduced pressure. The purpose of this research program was to study the equilibrium of the two reactions. During the investigation, reference to the experimental work of Schenk5 was found but, due to lack of detail in the report, it was felt that it would be desirable to continue the equilibrium measurements. Selection of Method The measurement of vapor pressure of metals presents obvious problems resulting from the high temperatures which must be employed. In the present case, the production of the "reaction pressure" introduces a further difficulty since a chemical reaction must proceed before any metal vapor pressure is produced, resulting in depletion of the re-

Jan 1, 1954

Recognizing the fact that the problems of the committee named by the Board to prepare the "official ticket" for officers and Directors of the Institute are various and difficult; and desiring to assist this committee by setting forth some of the principles that should guide the committee and some of the qualifications that should be required of candidates, the following instructions are formulated and the Secretary is directed to publish them in MINING AND METALLURGY concurrently with the publication of the names of each newly appointed Nominating Committee and to send a copy to each member of the Committee

Jan 1, 1943

By Harold R. Bradford

EXPANSION initiated during and after the war has placed industrial plants in new areas and increased reduction and manufacturing facilities in communities already established. With added expansion interest has grown in possible area contamination from waste products. Evolved flue waste contains many chemical contaminants that may or may not cause concern; however, atmospheric pollution has always been a problem to industrial communities. For some time effects attributed to small quantities of compounds of fluorine in various waste gases have been studied by industrial, medical, agricultural, and associated interests.1'2 Aside from an article by Churchill, Rowley, and Martin in 1948," little has been published on the subject of fluorine in coal except by foreign investigators. The first indication of fluorine compounds in coal was noted in England in 1934 by R. Lessing,4,5 who found that the presence of fluorine in Midland and West County coals was definitely established by etching of glass and by severe corrosion of porcelain fillings in a gas works scrubber. Etching of glass in a direct contact annealing kiln initiated research, and in 1944 four papers were published by H. E. Crossley,6-9 giving results of investigations of fluorine in coal of England by the Fuel Research Board, Department of Scientific and Industrial Research. These investigations studied methods of determining fluorine, the quantities present in various coal deposits, the geographical distribution of fluorine in British coals, occurrence of fluorine in minerals associated with coal, amounts of fluorine compared to other constituents of coal, the nature of fluorine compounds, and industrial significance of fluorine in coal. It was shown that British coals contain from 5 to 200 ppm of fluorine, which is probably present as fluorapatite [Ca10F2(PO4)6], often written as [3Ca3P2O8.CaF2]. The fluorapatite is intimately associated with shale in some coals and can often be separated from these coals along with the shales. H. E. Crossley8 also reported no correlation between the amounts of chlorine, sulfur, or ash and the amount of fluorine in different coals. He found, however, a relationship between the content of fluorine and phosporus in many British coals. Thus it should be possible to anticipate likely fluorine-rich coals from existing data giving phosphorus content. Behavior of the fluorine compounds of coal was compared with that of 12 fluoride-phosphate minerals in sink-float separations and solvent ex-

Jan 1, 1958

P. H. ARGALL, Magdalena, N. Mex. (written discussion *).-The Chilean mill has generally been considered a fine-crushing machine and as such has had many advocates; it has also. had quite a number of denouncers; hence, the action of the Portland engineers in using the Chilean mill for the preparatory work usually accomplished in, rolls or rock breakers and adding a ball mill to do the fine grinding is novel, though of doubtful utility in the long run, even though a fair efficiency is obtained for the combination machines, as measured by Chilean mill practice. In order to follow the development of Chilean mill practice on Cripple Creek ores, I might say that two 6-ft., Chilean mills were placed in operation in Stratton's Independence mill in April, 1908, treating ore crushed by rolls to pass Y4 in. screen aperture. Screening was, however, found to be an unnecessary refinement, as was also the use of plows in the mills; both screening and plowing were abandoned before the Portland engineers finished the experiments in stamp milling upon which they were then engaged. Two additional 6-ft. Chilean mills were added in the spring of 1911, three of them giving the required 10,000 tons per month capacity, one being always in reserve. The Portland company began using Chilean mills in June, 1910, at its Victor mill, and the same company purchased the Stratton's Independence mill in June, 1915, and remodeled it; it is now known as the Independence mill. A description of Stratton's Independence milling practice, is given in the November, 191.1, issue of Mining Majazine (London). These Chilean mills were used as fine grinders and that they accomplished their purpose is shown by the screen analyses, in which 62 per cent. of the pulp passed 150-mesh screens. To make coarse crushers of these Chilean mills, the Portland engineers have increased the speed, substituted steel in various parts for cast iron, provided new and enlarged thrust bearings to take care of the centrifugal force incident to higher speed, and have, thereby, greatly increased the cost of the machines. They have also reduced the width of the tires and dies from 8 in., as used in the original mill, to 7 in., a reduction of 12 ½ per cent. in the crushing surfaces, to say nothing of the loss of abrasion or attrition incident to the larger surface. At the same time, they have apparently increased the weight of the rollers by using tires 5 7/8 in. thick as against tires varying between 3 and 4 in. on the original mills, thereby increasing

Jan 4, 1920

By George Rice

THE subject of violent bumps in coal mines has been again brought to attention by a recent succession of such occurrences in the coal mines of the Cumberland field of eastern Kentucky and southern Virginia, which have caused fatalities and injuries to underground employees, a report on which is being issued by the United States Bureau of Mines1. The A.I.M.E. has had two valuable papers on bumps in the No. 2 Mine, Springhill, Nova Scotia, each followed by general discussion on bumps and their causes2. This paper gives a brief review of suggested causes and remedial measures given in English, Canadian and United States reports on bumps in coal mines, and then presents suggestions arising from the writer's recent investigation of the bumps in the Cumberland mines, made at the request of the mine operators, to determine the causes and remedies. These bumps, which occur under the simplest natural conditions, provide better opportunity for analysis of causes and effects than afforded under more complex conditions. The simple conditions in this field consist of relatively thin and level coal beds and uniform roof strata free' from folding and faulting.

Jan 1, 1935

By Richard H. Hopkins, R. W. Kraft

A new technique facilitating the rapid determination of the three-dimensional morphology of phases suspended in an opaque matrix is described. The vapidity of the technique is based upon continuous cinephotomicrographic recording of specimen microstructure as the specimen surface is removed under controlled conditions in an electrolytic cell. The advantages of the technique for determining spatial morphologies in opaque materials are illustrated by application of this method to the study of the configuration of faults in a specimen of directionally solidified CuAl2-Al eutectic alloy. A spatial model of the eutectic fault network deduced in this manner has substantiated the facts that faults in this alloy tend to lie parallel to one another in a direction close to that of the specimen growth axis and that the majority of faults tend to bridge lamellae rather than run parallel to them as growth proceeds. For the specimen studied fault density decreased as growth proceeded. ALTHOUGH metallographic observation is a powerful tool for investigating the correlation between structure and properties in metallic alloys, ordinary metallographic methods suffer the fundamental limitation that any conclusions concerning the three-dimensional structure of these opaque materials must rely upon inference from two-dimensional microsections. For this reason it is frequently impossible to make valid judgments regarding the true spatial distribution of the micro-constituents in specimens under investigation. Several techniques have been devised to circumvent the basic limitation of planar microsec-tioning and employed with varying success. These include construction of three-dimensional models of alloy microstructure from sequences of photomicrographs,1-3 quantitative deduction of spatial features based upon statistical studies of data from random and/or selected plane microsections,4-6 and selective dissolution of one phase in a two-phase alloy to reveal the topography of the phase remaining.' However these techniques also have certain obvious disadvantages. Construction of three-dimensional models is a very time-con- suming procedure due to the tedious polish-etch-photograph cycle needed to obtain the necessary photomicrographs to portray the spatial microstructure. Statistical methods by their very nature give only average values of the quantities being studied and cannot therefore give a detailed physical picture of the microstructure to be observed. Finally, the technique of selective dissolution is only applicable in systems where the phases are of vastly different chemical reactivity; it is of little use in the study of single-phase alloys. Since the determination of the three-dimensional structure of opaque materials is hindered by two facts—the materials are opaque to radiation in the visible spectrum and construction of spatial models of microstructure is quite laborious—two solutions to the problem suggest themselves: 1) use radiation of energy great enough to render the alloy transparent, i.e., some microradio-graphic method, or, 2) devise a rapid means to section and record alloy microstructure. Stereomicroradiography has indeed been used to study internal arrangements of phases in several materials8,9 but the method, in general, is limited to very thin specimens (due to X-ray absorption) and simple microstructures. It would seem then that development of a rapid, simple, and generally applicable technique for studying the details of alloy microstructures in three dimensions can best be undertaken by following the second approach above. Using this idea a new technique for rapid evaluation of spatial microstructures was developed based upon cinephotomicrographic recording of the micro-structure of a specimen undergoing controlled electrolytic dissolution. It was further felt that application of this technique to the careful study of the spatial configuration of faults in a CuAl2-Al eutectic alloy would yield information that could be correlated with the extensive data3,6,10-12 already compiled concerning the solidification and crystallography of this eutectic. The practicality and usefulness of similar methods for studying three-dimensional features in alloys has recently been illustrated by the study of dislocation loops in zinc13 and spheroidized structure in steel.14

Jan 1, 1965

By R. L. Slobod, B. H. Caudle

In the past, the main emphasis in attacking the problem of the recovery of oil has been on the determination of fluid flow characteristics and residual oil saturation in the part of the reservoir which is contacted by a displacing phase. It is recognized that the determination of the fraction of the reservoir contacted in secondary recovery operations is also of great importance in predicting the ultimate recovery to be expected. A method is described in which radiographic techniques are used to determine these areal sweepout factors for any type of well spacing. Data showing the relationship of viscosity ratio and relative permeability to areal sweepout efficiency are presented for the five spot and the direct line drive well spacing patterns. Further applications of the method to studies of the displacement efficiency in the area contacted are mentioned. INTRODUCTION The problem of increasing oil recovery from typical reservoirs is generally attacked by trying to devise some means for reducing the residual oil in that part of the reservoir which is contacted. The fraction of the reservoir contacted in the displacement process. however, is also of major importance but has been given less consideration in the past. one reason for this neglect is the meager information available on the portion of the reservoir contacted and the magnitude of changes in this factor as operating conditions are varied. Until recently, the only means available for determining this quantity has involved the use of the electrolytic or poten-tiometric modelsl in which one is limited to working with a mobility ratio of unity. while means for calculating the volume of a reservoir contacted do exist. the calculations are very tedious for systems with mobility ratio different from unity. Recently several such determinations have been reported,213 and an experimental method involving a step-wise use of potentiometric models has been described.' A new method which has been developed for studying the areal distribution of phases in a porous medium is discussed below. This tool can be' used to determine the effect of operating variables such as viscosity, relative permeability, presence of gas, mobility ratio, rate and stratification on the fraction of the reservoir contacted by the displacing medium. This development based on the use of X-rays gives a pictorial representation of the location of the phases. Further, the amount of each phase at each point of interest can be determined quantitatively. X-RAY METHOD FOR AREAL STUDIES In this method the location of various phases distributed Over a test area are observed simultaneously by obtaining a photographic record of the transmission of X-rays through the test plate. This procedure is an adaptation of the radiographic techniques which have 10% been used in the medical and industrial fields. Just as the medical profession uses a chest plate to show spatial interrelationships, the location of an X-ray absorbing fluid and a nonabsorbing fluid in a porous medium can be determined by obtaining a photographic record of transmitted X-rays. Moreover, as changes occur in such a System, these changes can be observed by taking successive photographs. Previously, X-ray methods have been used in fluid flow studies to determine quantitatively the saturation at a given point in the system. The intensity of the transmitted X-rays was measured by an ionization chamber, so that a point by point survey of saturation could be made. In comparison with this earlier quantitative procedure, the photographic procedure which simultaneously observes the saturation at many

Jan 1, 1952

By Paul Weir

BLEACHING clays have been used extensively in the oil-refining industries for a number of years. Their use in water purification is relatively recent and less extensive. They are frequently classified as (1) those naturally active and (2) those active after artificial activation with acid. Fuller's earths fall within the general group of bleaching clays and gradually the commercial term "fuller's earth" has been applied to natural clays originating, for the most part, in the south-eastern section of the United States. Bentonitic clays are usually classed as activable and are prevalent in the southwestern section of the United States.1 Bleaching clays vary considerably in chemical and physical properties throughout the country, or even in different parts of one deposit (Table 1). TABLE 1.-Physical and Chemical Composition of Different Types of Clays Composition, Per Cent Constituents and Properties Bentonite Fuller's Mud Clay (Wyoming) Earth (P Peak)' (Alluvial) Silica (SiO2) 60.0 69.6 61.67 Alumina (Al203) 19.0 14.3 20.34 Ferric oxide (Fe203) 2.8 3.90 7.70 Titanium oxide (Ti02) 0.13 0.70 1.01 Lime (CaO) 0.5 0.80 0.00 Magnesia (MgO) 2.0 1.00 0.24 Soda (Na2O) 2.4 0.40 0.31 Ignition loss 4.7 6.77 8.51 Moisture content 8.0 10.0 20.0 Isoelectric point, pH 9.2 9.2 6.5 Average swelling volume 15.0 0.0 0.0 Weight per cubic foot, lb 55.0 40.0 65.0 Material passing 200-mesh sieve, per cent 90.0 85.0 50.0

Jan 1, 1939

Jan 1, 1947

Jan 1, 1947

By C. T. Liu, J. Gurland

The deformation behavior in tension of spheroidized carbon steels was studied at room temperature as a function of carbon content, 0.065 to 1.46 wt Pct, and carbide particle size, 0.88 to 2.77 p. It was found that the Hall-Petch strength-grain size relation is directly applicable to the yield and flow stresses of the two lower-carbon steels , 0.065 and 0.30 pct C. The strength data for the medium- and high-carbon steels, 0.55 to 1.46 pct C, also satisfied the Hall-Petch relation, provided that these data are based upon the particle spacing. Beyond 4 pct strain, the flow stress data of all the steels studied could be represented by the same Hall-Petch relation with dinerent spacings for grain boundary and particle strengthening. The behavior of the higher-carbon steels was consistent with the postulated formation of a dislocation cell network during processing and initial deformation (up to 4 pct strain). The cell size was assumed to be equal to the planar particle spacing. The true stress at the ultimate tensile strength was also found to be a function of the particle spacing. At a given temperature and strain rate, the yield and flow stresses of carbon steels depend on the type and dimensions of the microstructure. Starting with the work of Gensamer et al. in 1942,' experimental studies on pearlitic and spheroidized carbon steels revealed that the strength of steels is a function of two main parameters: the ferrite grain size2'3 and the carbide particle spacing;1'4'5 on this basis, two different strengthening mechanisms have been developed to apply to steels of low and high carbon contents, respectively. In polycrystalline iron and mild steels the grain boundaries are regarded as the major structural barriers to slip. The relation between strength and grain size is generally represented by the Hall-Petch equation which is based on a linear proportionality between strength and the inverse square root of the average grain size.2'3y677 However, Gensamer et al.' and Roberts et related the yield strength of medium -and high-carbon steels to the carbide particle spacing alone, and they found a linear relation between the logarithm of the mean free path in the ferrite and the yield strength in both spheroidized and pearlitic steels. By means of the electron microscope, Turkalo and LOW' extended the study to finer structures; they concluded that the logarithmic relation is not valid for the entire range of microstructures unless grain boundaries are also included in the measurement of the mean free path. For the specific case of spheroidized steels, Ansell and aenel' found that the yield strength data,4'5 when plotted as a function of mean free path, fit the Hall-Petch equation; however, T'ysong found that the same data fit the 0rowanl0 relation if a planar inter-particle spacing is used. Recently Kossowsky and ~rown" studied the strength of prestrained spheroidized steels, 0.48 and 0.95 pct C, and concluded that the strength due to the carbide dispersions varies linearly with the reciprocal of the square root of the mean free path between carbide particles and dislocation networks. Such networks were first observed by Turkalo." The conclusion common to all these studies is that the available slip distance in the ferrite is the most important variable in determining strendh. Previous work on carbon steels is restricted to limited composition and strain ranges. The mechanism which governs the flow properties is not clearly understood, and, in particular, little is known about the composition dependence of the transition between grain boundary strengthening and particle hardening. The purpose of the present work is to investigate the strengthening mechanism in spheroidized steels over a wide range of carbon content, 0.065 to 1.46 wt pct, and plastic strain, yielding to necking. The spheroidized structure was chosen because of its relative simplicity and the relative ease of control and measurement of the structural parameters. The experimental work is limited to tensile testing at room temperature at constant extension rate. The effects of the carbide particles on the fracture behavior of spheroidized steels are discussed elsewhere.13 EXPERIMENTAL PROCEDURE Eight different grades of vacuum-cast carbon steels were supplied in the form of forged and rolled plate by the Applied Research Laboratory of the U.S. Steel Corp. The compositions furnished with these steels are given in Table I; the carbon content ranges from 0.065 to 1.46 wt pct, or from 1.0 to 22.3 vol pct of carbide. The steel plates were cut transversely into rods a little larger than the test specimens, 1 in. gage length, i in. diam. The rods were austenitized in air (enriched with CO by a consumable carbon-rich muffle) at 50° C above theA, orA., temperature for 2 hr and then quenched in oil with vigorous stirring. The as-quenched rods were tempered in two stages in order to obtain the desired distributions and sizes of carbide particles. The rods were first tempered at 460° C for 10 hr and then at 700" C for periods ranging from 4 hr to 3 days, in vacuum. After final machining, all specimens were vacuum-annealed again at 650°C for 1 hr in order to relieve residual stresses. The tension tests were carried out in two steps. The initial part of the load-strain curve, up to about 2 pct strain, was determined on a Riehle testing machine with an extensometer of small strain range, 4 pct strain, in order to obtain the yield and initial flow piopertiesi As soon as the first part of the test was finished, the specimen was placed in an Instron testing machine equipped with a strain gage extensometer with a maximum strain range of 50 pct. The load-strain curve to fracture was

Jan 1, 1969

By J. E. Dutrizac, R. J. C. MacDonald, T. R. lngraham

When sintered disks of synthetic chalcopyrite (CuFeS2) were leached in acidified aqueous solutions of ferric sulfate, the following reaction stoichiometry was obtained: CuFeS2 + 2Fe2(SO4)3 = CuSO4 + 5FeSO4 + 2S Over the temperature range from 50º to 94ºC, the reaction displayed parabolic kinetics. The parabolic rate constant for the dissolution of copper is given by the equation: log.k(mg2/cm4-hr)= 11.850 - 3780/T The activation energy for the dissolution process is 17 ± 3 kcal per mole. The parabolic kinetics have been attributed to the progressive thickening of a sulfur film on the surface of the chalcopyrite. When the leaching solutions contain less than 0.01 molar Fe+3 , the Fe concentration influences the rate of leaching, probably through a mechanism involving the diffusion of ferric sulfate through the sulfur layer. At higher Fe+3 concentrations, the rate control in the leaching. reaction has been attributed to the diffusion of ferrous sulfate through the sulfur. The rate of reaction is insensitive to changes in acid concentration and in disk rotation speed. ThE reaction of acidic ferric sulfate solutions with various sulfide minerals is of practical interest for both bacterial and heap leaching. This leaching medium is generally used with low-grade ores that cannot be treated profitably by conventional means. In both bacterial leaching1-3 and heap leaching, the active agent for sulfide dissolution is ferric sulfate. Although the reactions of ferric sulfate with chalcocite, covellite, and bornite have been investigated,4*7 the kinetics of leaching chalcopyrite with ferric sulfate have not been thoroughly studied. This paper reports a study of that reaction. EXPERIMENTAL Reagent-grade sulfur was purified by the method of Bacon and FanelliB and then it was vacuum-distilled to remove any soluble magnesium salts that had been introduced during the purification procedure.9 From stoichiometric quantities of the purified sulfur and hydrogen-reduced electrolytic copper sheet (99.90 pct Cu), CuS was synthesized at 450°C in a vacuum-sealed, pyrex vessel. About 24 hr was required for the completion of the reaction. A similar procedure involving hydrogen-reduced iron wire (99.90 pct Fe) was used to synthesize FeS1.002. A 2-furnace arrangement was required. The iron was heated to 800°C while the sulfur was maintained at about 400°C. Although the reaction to consume the sulfur was rapid, the material required additional heating (1 week) in a sealed silica ampoule at 800°C before it was homogenized. X-ray powder diffraction analysis confirmed that the copper sulfide was covellite and that the iron sulfide was troilite. The composition of the iron mineral was confirmed by wet chemical analysis. The two sulfides were ground to minus 100 mesh, weighed in equimolar amounts, mixed thoroughly, and pressed into pellets at 80,000 psi. The pellets were vacuum-sealed in pyrex ampoules and then sintered for 3 days at 550°C after an initial heating at 450°C for a few hours. The pellets were then cooled, polished with 3/0 emery paper, rinsed in acetone, and stored. The material had the characteristic brassy color of chalcopyrite and was shown by X-ray diffraction to be CuFeS2. Microscopic examination of the polished surfaces revealed small inclusions of pyrite (approximately 0.5 vol pct) as the only impurity. The presence of small amounts of a second iron compound will not alter the amount of dissolved copper but might increase the amount of ferrous ion slightly. It was calculated that dissolution of all of the pyrite and 100 mg of Cu (a typical value) would change the expected ferrous concentration by only 4 pct. Microscopic examination of a pellet after leaching revealed that the pyrite was not preferentially solubilized; only those pyrite grains at the surface were attacked. Hence, the pyrite is unlikely to alter the rate of copper dissolution. The chalcopyrite disks were about 1.7 mm thick and 27 mm in diam. They were about 80 pct of theoretical density, and for this reason their true reaction area was somewhat larger than the 5.8 sq cm area presented by the polished face. The disks were cemented to lucite cylinders in such a way that only the polished face was exposed. The disks were then leached by methods previously described.6,7 RESULTS AND DISCUSSION Stoichiometry and Kinetics. The initial experiments were directed to the problem of resolving the stoichiometry of the leaching reaction. Disks of CuFeS2 were leached at 80°C for various periods of time in acidified ferric sulfate solutions that were protected from oxidation by a cover of flowing nitrogen. When the disks had been partly leached, they were removed, their soluble salts were washed out, and then they were treated with CS2 in a Soxhlet extraction apparatus. The ratio of elemental sulfur to dissolved copper thus obtained was approximately 2 to 1. After the extraction of elemental sulfur from the pellet, the residue consisted of unreacted chalcopyrite only. For runs in which an appreciable amount of copper was dissolved, the ratio of ferrous ion to cupric ion in the solution was

Jan 1, 1970

A. L. FIELD, Cleveland, Ohio (written discussion*.-)In equation 2, B is used to denote the ratio of bases (lime plus magnesia ) to acids (alumina plus silica) it being stated that this ratio gives more concordant results than the ratio either of bases to silica, or of lime to silica. I find, however, that, by making use of the ratio between true CaO (CaO by analysis minus sulfur calculated to CaS) to acids, equation 2 may he modified to agree with the observed facts within an average error of 0.0041 per cent. sulfur instead of 0.0045 per cent. The revised equation is s = 0.072 + 0.01835 - 0.00061 (t - 1400) - 0.0403B where B is the ratio of true CaO to Al2O3 Plus Si02, and the other symbols have the same significance as before. I would not appear to stress too much this slightly better agreement of the revised equation with the observed sulfur. Yet I believe that the use of this ratio of true CaO to acids is to be preferred, if for no other reason than the fact that it is more in harmony with the previous findings of the Bureau of Mines with regard to the effect of magnesia and other impurities on slag viscosity.1

Jan 12, 1919

By W. A. Barrows, Carl Zapffe

Jan 1, 1913