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By N. A. Gokcen, John Chipman

SILICON is the most commonly used deoxidizer and an important alloying element in steelmak-ing; hence a detailed study of this element in liquid iron containing oxygen is of considerable interest. The equilibrium between silicon and oxygen in liquid iron has been studied by a number of investigators but generally with inconclusive or incomplete results. The variation of the activity coefficients of silicon and oxygen with composition is entirely unknown. Published investigations deal with the reaction of dissolved oxygen with silicon in liquid iron and the results are expressed in terms of a deoxidation product. For consistency and convenience in comparison of the published information, the deoxidation product as referred to the following reaction is expressed in terms of the percentage by weight of silicon and oxygen in the melt in equilibrium with solid silica: SiO (s) = Si + 2 O; K'l = [% Si] [% 012 [I] Theoretical attempts to calculate the deoxidation constant for silicon in liquid iron from the free energies of various reactions yielded results which were invariably lower than the experimental values. Thus, the deoxidation "constants" calculated by McCance,1,2 Feild,3 Schenck, and Chipman were of the order of 10, which is below the experimental values by a factor of more than 10. Experiments of Herty and coworkers" in the laboratory and steel plant resulted in an average deoxidation constant of 0.82x10 ' at about 1600°C. The technique employed in their investigation was crude and the reported temperature was quite uncertain. The concentration of silicon was obtained by subtracting silicon in the inclusions from the total. Since at least some of the inclusions resulting from chilling must represent a fraction of the silicon in solution at high temperatures, such a subtraction is not justifiable. Results of Schenck4 for K'1 from acid open-hearth plant data yielded a value of 2.8x10-5, which was later revised as 1.24x10 at 1600°C. Similarly Schenck and Bruggemann7 obtained 1.76x10-5 at 1600OC. The discrepancies and errors involved in the acid open-hearth plant data as compared with the results of more reliable laboratory techniques were attributed by these authors to the lack of equilibrium and the impurities in liquid metal and slag, and are sufficiently discussed elsewhere." Korber and Oelsen" investigated the relation between dissolved oxygen and silicon in liquid iron covered with silica-saturated slags containing varying concentrations of MnO and FeO. The deoxidation products obtained by their method scatter considerably, and their chosen average values of 1.34x10, 3.6x10-5, and 10.6x10-5 1550°, 1600°, and 1650°C, respectively, represent the best experimental results which were available until quite recently. Darken's10 plant data from a steel bath agree approximately with their data at 1575° to 1625°C. Zapffe and Sims" investigated the reaction of H2O and H2 with liquid iron containing less than 1 pct Si and obtained deoxidation products varying by a factor of more than 20. Inadequate gas-metal contact and lack of stirring in the metal bath should require a longer period of time than the 1 to 5.5 hr which they allowed for the attainment of equilibrium. Furthermore, their oxygen analyses were incomplete and irregular and confined to a few unsatisfactory preliminary samples. Their results did indeed indicate that the activity coefficient of oxygen is decreased by the presence of silicon, although they made no such simple statement. They chose to attempt to account for their anomalous data by the unlikely hypothesis that SiO is dissolved in the melt. Hilty and Crafts" investigated the reaction of liquid iron with acid slags under an atmosphere of argon, making careful determinations of silicon and oxygen contents at several temperatures. Despite erroneous interpretation of the data at very low silicon concentrations, their data represent the most dependable information on this equilibrium that has been published. In the range 0.1 to 1.0 pct Si, their data yield the following values for the deoxidation product: 1.6x10-5, 3.0x10- ', and 5.3x10 at 1550°, 1600°, and 1650°C, respectively. The purpose of the work described herein was to study the equilibrium represented by eq 1 as well as the following reactions, all in the presence of solid silica: SiO2 (s) + 2H2 (g) = Si + 2H2O (g);

Jan 1, 1953

By R. A. Graham, R. W. Rohde

The effect of hydrostatic pressure upon the austenite start temperature of a commercial Fe-28.4 at. pct Ni-0.5 at. pct C alloy has been determined. For pressures to 20 kbar, the austenite start temperature decreased from its atmospheric pressure value of 380°C at the rate of about 4°C per kbar. These data are analyzed by two different thermodynamic approaches; first, considering the transformation as an isothermal process, and second, considering the transformation as an isentropic process. It was found that both these approaches fit the experimental data equally well. The effect of hydrostatic pressure upon the austenite start temperature is best described by considering the mechanical work done during the transformation as that work obtained by multiplying the applied pressure with the gross volume change of the transformation. It is widely recognized1 that strain has an important effect on the initiation of martensitic transformations.* For example, the martensite start tempera- *In this paper, use of the term martensitic transformation implies the reversal of martensite to austenite as wen as the formation of martensite from austenite. ture, M,, may be increased by plastic deformation. Similarly, plastic deformation is observed to lower the austenite start temperature, A,. The effect of uniaxial stress on the M, of iron-nickel alloys has been studied by Kulin, Cohen, and Averbach.2 They found that the martensite start temperature was significantly changed by stresses well within the elastic region. Moreover, the effect of tensile and compres-sive stresses differed. These effects were explained in terms of the interaction of the applied stress with both the dilational and shear components of the transformation strain. The magnitudes of the influence of uniaxial tension, compression and hydrostatic pressure on Ms were measured in 30 pct Ni 70 pct Fe by Pate1 and Cohen.3 Their thermodynamic calculations and similar calculations by Fisher and Turnbull4 predicted the experimental results when the transformation was assumed to occur isothermally at some fixed driving force. This driving force was assumed to be supplied by a combination of the chemical free energy difference between the austenitic and martensitic phases and the work performed during transformation by the applied stress. More recently, Russell and winchel15 reported the effect of rapidly applied shear stress on the reversal of martensite to austenite in iron-nickel-carbon alloys. They performed a thermodynamic analysis of this transformation based upon the assumption that the re- versal occurred adiabatically. They concluded that the applied shear stress did not significantly interact with the transformation strain and thus did not assist in inducing the reversal. Rather they concluded that the reversal was effected by localized strain heating which resulted from the gross local shear deformation of the experiment. In either the adiabatic or isothermal analysis it is necessary to compute the work performed by the interaction of the applied stress and the transformation strains. In the case of hydrostatic pressure this interaction has been treated by two different methods. In either case the applied pressure is assumed to remain constant during the transformation. In one treatment the applied pressure is assumed to interact directly with the dilatational strain associated with the formation of an individual martensite plate.3'4 This local strain has been measured at atmospheric pressure in iron-nickel alloys by Machlin and Cohen.6 In the above treatment this local strain is assumed invariant with temperature and pressure changes. In the other treatment the applied pressure is assumed to interact with the gross volume change of the transformation.7,8 The usefulness of this latter treatment has been demonstrated by Kaufman, Leyenaar, and Harvey7 who calculated the effects of pressure upon the martensite and austenite start temperatures of Fe-10 at. pct Ni and Fe-25 at. pct Ni alloys. Excellent agreement was obtained between their calculations and their experimental data on an Fe-9.5 at. pct Ni alloy. However, this treatment suffers from the fact that the data required to calculate the volume change of the transformation (i.e., the initial specific volumes, the thermal expansion and compressibility data for both the austenitic and martensitic phases) is, in general, not available for any material except pure iron. Thus the calculations of Kaufman et al.7 were necessarily performed by assuming that the volume change of the martensitic transformation in the iron-nickel alloys was that same volume change occurring during the a-? transformation in pure iron. While this approximation may suffice for very dilute alloys it is likely to be inaccurate in high nickel alloys. We have performed measurements of the effect of hydrostatic pressure to 20 kbar on the A, temperature of an Fe-28.4 at. pct Ni-0.5 at. pct C alloy. The composition is similar to the alloy used by Pate1 and Cohen3 to determine the effect of pressure upon the M, temperature. The present measurements permit calculation of the interaction between the applied pressure and the transformation strain. Additionally, measurements have been made which allow precise determination of the gross volume change of the transformation. The data allow direct comparison between the alternate hypotheses of the interaction between the applied pressure and a dilatational transformation strain characterized by either the formation

Jan 1, 1970

By A. E. Lillstrom

IN the development of iron mines and production of iron ore from the Marquette range, drilling blast-holes is an important phase of the mining cycle. The ground drilled in ore production can be classified into two main categories, soft hematite and hard hematite or magnetite. Within these categories the material exhibits a wide range of penetrability by percussion drills. Development work encounters various types of rock. Slate and altered basic intrusives constitute the softer types commonly encountered. Harder materials are represented mainly by greywacke, quartzite, iron formation, and diorite. Prior to the first tungsten carbide trials in late 1947 and early 1948, hard-rock and ore drilling was done with steel jackbits starting at 21/4-in. diam. These were reconditioned by hot milling. Automatic or handcrank 31/2-in. drifters were employed, mounted on Jumbos, posts and arms, or tripods, depending upon the working place. With the exception of shaft sinking jobs where 55-lb sinker machines were and still are used with 1-in. quarter octagon steel, the other production and development mining utilized 11/4-in. round and Leyner-lugged steel. The following properties have been selected as typical examples wherein carbide bit applications have proved economical. The Mather mine "A" and "B" shafts and Cleveland-Cliffs Iron Co. mines are soft ore mines where insert bits are used in rock development only. The Greenwood mine, Inland Steel Co., Champion mine, North Range Mining Co., and Cliffs shaft mine, Cleveland-Cliffs Iron Co., are hard ore mines where all drilling is done with tungsten carbide bits. Mother Mine "A" Shaft In the Mather mine "A" shaft and other soft ore properties where only rock development work is done with the tungsten carbide bits, several types and makes of bits have been tried since early 1948. The greatest proportion of failures have been at the connection end, although the early trials with the 13 Series Carset 11/2-in. bit used in conjunction with 31/2 -in. automatic-feed drifters, showed an equal amount of shattered inserts. To combat this shattering, the 31/2 -in. drifters were replaced by 3-in. drifters, thus eliminating, for the most part, insert failures. However, the attachment end of the rod continued to be the main source of trouble. The greatest amount of failure was in the stud or at the upset section approximately 2 in. behind the drive shoulder of the rod. Heat treatment was changed several times as well as the composition of the alloy studs. Since this failed to correct the trouble, a decision was made to change to a heavier attachment section. Timken 11/2-in., type M, bits were then employed and showed an exceptional improvement. The rods are discarded when the thread contour shows sharpening or wear on the shoulder. It was also learned that the Timken insert did not show as rapid gage and cutting edge wear as did competitive makes, and footage per use increased by approximately 50 pct. Prior to the Timken trials the average life per bit at the Mather mine "A" shaft on 6-ft change chain-feed drifters was 500 ft, and the rod life at the connection end was 50 ft. The Timken bit with chrome-plated thread averaged 1200 ft, and rod life increased to as much as 500 ft. However, the life of the connection end was much better on shorter length drill rods or in places where machines with 34-in. change were used. The bit thread continued to be the point of ultimate failure with thread strippage, constituting the cause for discard of bits. In one of the new development headings, harder rock was encountered for approximately 800 ft, dropping the life per bit to a low of 90 ft with shank and thread life of rods dropping to approximately 125 ft average. The stripped bits were then welded to the rods, increasing the life per bit by 75 to 100 pct. The rod transportation for main level development was not a problem so intraset rods were tried. Intraset rods have tungsten carbide inserts set into the rods proper by the manufacturer and can be obtained with chisel or four point bits. This type of rod eliminates the need for any connection and the steel being a special alloy will show more feet drilled per rod. The first trial was made with eight rods, and final results averaged 350 ft per rod, six of the rods worked the life of the bit end, and two broke shanks at less than 50 ft. The preceding example showed a considerable improvement, so additional steel of the same type was purchased, but its use has been limited to main level drifting only, because of the handling problem involved in transportation of the complete rod to mine shops for resharpening. Further trials are being made on improving the life per detachable bit by chrome plating. To date, the chrome plating shows an improvement of approximately 100 pct. However, final results will not be known until the present long term trials have been completed. Mother Mine "B" Shaft In November 1947, tungsten carbide bits were first tried at the Mather mine "B" shaft. The use of 1%-in. Carset 13 Series bits, for drilling the 72-hole, 7-ft shaft round, decreased the drilling time from an average of 41/2 hr per round required with steel bits, to 2 hr with insert bits. The best drilling time for

Jan 1, 1952

By H. B. Pulsifer

.ABOut 1.5 varieties, or tnodifications, of the best rnagnesiurn available were prepared and subjected to etching tests, then examined for micro-structure. Of the 30-udd etching reagents that were tried, nearly half, mostly ammonium salts, etched the metal satisfactorily. The surfacing and etching of magnesium is shown to be a very simple and quick operation. The density and hardness of magnesium were determined. The most interesting new observations relate to the finding of the hexagonal etch figures, the crystal laminations, and the fact that the metal is plastic, cold, when restrained or quickly deformed. Under slowly applied pressule cold metal deforms slightly, then shears to fracture without plastic flow. The chief structural features of magnesium arc presented in two reduced photographs and 30 photomicrographs. Materials Tested There are only two sources in the United States from which new metal can be procured: the American Magnesium Corpn. and the Dow Chemical Co. No pronounced structural or property differences in the metals from these two companies were disclosed by the work of this investigation. The following materials were obtained from the manufacturers: 1. Massive crystals of distilled metal (American Magnesium Corpn.). 2. Rods of hot-extruded metal, ½-in. squares and 5/8-in. rounds (American Magnesium Corpn.). 3. Sheet magnesium, 0.005-in. thick (American Magnesium Corpn.). 4. Cast stick metal, 13/8-in. dia. (Dow Chemical Co.). 5. Magnesium-aluminum alloy, hot-rolled plate, ½-in. thick (American Magnesium Corpn.). From these materials, were prepared: 6. Sections from a furnace-cooled ingot made from clistilled crystal. 7. Cold-strained pieces from (6), (4) and (2). 8. Pieces cold-squeezed to fracture from (6), (4) and (2). 9. Hammer-struck pieces from (6), (4) and (2). 10. Steam-hammer struck pieces from (6), (4) and (2). 11. Sections from furnace-cooled ingot of the magnesium-aluminum alloy.

By L. Valentik

Many attempts have been made during the last 40 years to evaluate the performance of gravity separation equipement, that is, the effectiveness with which light and heavy particles are separated. The most comprehensive treatment of the subject was made by Cerchar at the 1st International Conference on Coal Preparation held in Paris in 1950. The methods suggested by the Conference were accepted and very widely used in the last two decades. This paper discusses an improved method of evaluation in the light of the now-accepted standard presentation. The float-and-sink analysis of the product is presented on a Gaussian distribution curve, resulting in an easier visualization of the inherent difficulties of separation. The ogives of the distribution curve me then plotted, giving a quantitative measure of the deviation from perfect separation as an error distance instead of an error area. Illustrations of the new method are given both for gravel and for coal preparation, but the content is valid and applicable to other types of minerals which are separated by gravity methods. Many attempts had been made during the last forty years to evaluate the performance of heavy-media separation (HMS) equipment, that is, the effectiveness with which floats and sinks are separated.'-' The most comprehensive treatment of the subject was made by Cerchar at the 1st International Conference on Coal Preparation held in Paris. 6 The primary aim was the thorough understanding of the mechanism of separation and the unified presentation of data on gravity separation so that the evaluation and comparison of washery performance could be made from all over the world. No strict overall standardization has been achieved, but after the conference a more or less uniform presentation of performance was accepted, which, during the last two decades, has been very widely used. In this paper, illustration of the old methods and an improved method of evaluation will be given. HEAVY-MEDIA SEPARATION (HMS) PERFORMANCE CRITERIA In the ideal HMS process, all material lower in density than the specific gravity of separation (SGS) would be recovered as floats and all material of higher density would appear as sinks. In order to evaluate the misplaced material, the washery products are tested at the density at which the washing unit is operated. The original type of plot1,7, 8 is shown in Fig 1; this was developed primarily for coal cleaning units. The curve for raw coal represents the cumulative percentages of sink material. The refuse curve is also plotted as a cumulative sink, the percentages being expressed in terms of raw coal. This diagrammatic representation of the results of washing units has the merit of easy visual observance of the degree of separation obtained. The error areas (cross-hatched) are a measure of the amount of misplaced material and therefore they can be used to characterize the quality of separation. The ideal and actual separating performance between floats and sinks can be best seen from the partition curve developed by Tromp,2 where the ordinate is the percentage recovery of the sinks, and the abscissa is the specific gravity (Fig. 2). It can be seen from the shape of the curve that as the SGS is approached, the proportion of material reporting to the improper product increases rapidly. In fact, the SGS can be defined as the density of the material in the feed that is distributed equally between float-and-sink products. When the upper half of the curve is inverted, a shape similar to that of a Gaussian error distribution curve is obtained and therefore the analysis of gravity separation may be carried out by using the law of probability. The shape of the curve in Fig. 2 is determined partly by the density composition of the feed, and partly by the sharpness with which the unit separates floats from the sinks.9, l0

Jan 1, 1970

By J. E. Berry

AII evaluation is made of the acoustic velocity log for measurement of formation porosity. Plots of field-observer1 velocities vs core-measured porosities of sandstones and limestotnes with inter intergranular porosity show that the velocity log gives a useful measure of porotity, in agreement with published data. A new method of plotting electrical and velocity log data makes it easier to recognize hydrocarbon-water coritnct~ from log darn and to make serni-quantitative estinlates of hydrocarbon saturation. Two examples of this interpretation technique are given. This paper also discusses the various factors that affect acoustic velocity and shows how corrections can be made for some of these factors. INTRODUCTION Although the continuous velocity log was developed primarily as an aid to seismic interpretation, it is finding widespread use for measurement of porosity. The purpose of this paper is to evaluate this use of the log. Results are presented for the velocity-porosity relationship in many sandstones and one limestone. Practical applications of porosity values derived from the velocity log to improve interpretation of electrical surveys are presented. Various parameters which affect velocity in porous media are considered. Porosity, composition, cementa-tion, pressure difference (overburden pressure minus fluid pressure), fluid saturation, and wellability are discussed. Some of the ideas are opinions based on general considerations of well logging problems while others are supported by published data and our own observations. Major conclusions from these studies arc given in the body of the paper and detailed discussions of some of the various factors affecting porosity are given in the Appendix. POROSITY Sandstones The primary factor affecting the velocity of sound in porous media is porosity. Wyllie, et al,1 have reported a time average equation for the relationship between velocity and porosity.* This has found considerable acceptance in the industry. Other relationships between velocity and porosity have been proposed, none of which are entirely satisfactory. It is doubtful that any one relationship will completely describe all situations. We have tried an empirical relationship based on correlating log velocities with core-measured porosities. Measured porosities were averaged over a depth interval equal to the spacing used in recording the velocity logs. The depth range for the information used was from 2,000 to 12,000 ft. Velocity measurements were normalized for the effects of depth and overburden pressure. A detailed discussion of the basis for this normalization is given in the Appendix. The change is usually insignificant and when working with individual formations it can be neglected. All velocity measurements were normalized to a depth of 10,000 ft. The velocity-porosity data for sandstones in Fig. 1 came from 20 wells scattered throughout the Mid-Continent and Gulf Coast areas. The scatter of the data points is attributed to three causes: 1. There is no theoretical basis for expecting an exact velocity vs porosity relationship and it is not possible to correct precisely for the effects of such factors as impurities, grain shape, manner of cementation, etc. Some of these factors are discussed in the Appendix. 2. Porosities measured by core analysis are not necessarily representative of what the velocity log "sees". 3. Inaccuracies appear in the velocity log reading. This is particularly true of velocities from one-receiver logging tools.*" A linear-time curve (time-average equation) is plotted in Fig. 1 along with the best straight line through the points. Velocities used for the linear-time curves were 19,500 ft/sec for the matrix velocity (zero porosity) and 5,500 ft/sec for fluid velocity. For clean sandstones with velocities greater than 10,000 ft/sec, there is little difference between the velocity-porosity and the time-porosity curves. Both curves are well within the scatter of data points. The data points in Fig. 1 were restricted primarily to clean sandstones, using the SP and/or gamma-ray logs as clues to the lithology. Where a lithologic de-

By Maurice Grieves

The great majority of shafts constructed today are still excavated by drilling and blasting, a method which changed very little in over 100 years until the introduction of the mechanical lashing unit and cactus grab by the South Africans, which enabled muck to be removed as fast as massive hoisting systems could handle it and resulted in very rapid rates of sinking. Record breaking month's performances were achieved at -Hartebeestfontein No. 4 shaft, October 1960-337.1 m; Western Reefs No. 4 shaft, October 1961-340.7 m; and Buffelsfontein eastern twin shaft, March 1962-381.2m. The method was very labor-intensive, requiring a crew of over 60 workers at the shaft bottom during the drill cycle. Safety precautions were strict, but in the drive to achieve rapid advance, cases of personal injury were still somewhat high because of the large number of people engaged in this potentially hostile environment. The South African method, as it came to be referred to throughout the rest of the world, was adopted in the United Kingdom in the late 1950s in a modified form with greatly reduced manpower and nonsimultaneous sinking and lining, which was insisted on by the British Mines Inspectorate. In that instance, it was successfully used to sink the 7.3-m-diam concrete-lined shafts at Kellingley to 770 m depth, with rates of advance of over 90 m/month achieved, a British record at that time. During the sinking of the 1.15-km-deep twin shafts at Boulby potash mine in the UK in 1970, the method was again used, but for the first time ever in Britain exemptions from the mining code permitted the use of crash beams, crash doors, jack catches, and semi-simultaneous sinking and lining techniques. New British shaft sinking records of over 120 m/month were achieved in both shafts. Similar equipment and techniques were used in the early 1970s to sink several deep shafts in Canada, notably Creighton #7 and the Con zinc mine at Yellowknife in the NW Territories. Today, this equipment is standard for deep shafts in the US and the rest of the world. However, with the tremendous escalation in mining labor costs, the impact of health and safety legislation, and environmental regulations, coupled with a very real shortage of miners willing to work in this exposed situation it was apparent that an alternative to the labor-intensive conventional method of shaft construction had to be found. Recognizing the trend is inevitable, one or two major German shaft sinking contracting firms began to take a fresh look at full face boring techniques applied to tunnels and raise bored shafts. The results were most encouraging. Tunnel drivage techniques using moles had developed considerably from Colonel Beaumont's original channel tunnel machine circa 1880 to the superbly engineered Priestley machine selected to cut the British side in 1975 and the double shielded Robbins Grandori borer on the French side of the English Channel. Full face tunnel machines were being successfully used to drive uphill in inclined shafts in Austria and Switzerland. At Mapprag in Switzerland, the Demag mole drove the first (intentional) vertical transition and curve, and then went on to successfully complete the 730-m-long penstock shaft at an inclination of 35°. In Austria, the Wirth mole drove the Kaprun Glacier ski-lift railcar tunnel at record breaking rates of 457 m/MONTH (best 30 m/d) through green schist at an inclination of 29° for a distance of 3.35 km while the Hydro tunnel at Sarrelli in Switzerland was being driven by the Robbin's mole at an inclination of 35°. Simultaneously, extremely promising results were obtained using large assemblies of cutter discs on raise borer heads, such as the 4.87-m-diam X 460-m-deep shaft raised by Teton for Jim Walter Resources Inc. Bearing in mind that most mine shafts in the future (unless in exceptionally competent rock) will require some form of lining, and the trend will be toward deeper shafts as the more easily accessible mineral deposits become exhausted, it was seen that normal raise boring had definite limitations in vertical accuracy, in the limitation imposed by the drill string on the available torque that could be applied to the cutter head, and in the risk of collapse of the unsupported shaft rock wall in friable or jointed and fissured ground, since it is not possible to apply any form of temporary support until the permanent lining is being installed. A further problem was the economics of installing a subsequent lining, necessitating setting up a headgear and hoisting arrangement approaching in size that required for conventional drill and blast sinking and lining. Because of the economics, German contractors opted for a phased transition from drill and blast to the full face, rodless, out of the solid shaft mole, by starting off with a down-the-hole shaft boring machine -without a drillstring-but using a pilot hole to get rid of the muck.

Jan 1, 1982

By H. Conrad, V. V. Damiano, G. J. London

A method is described for producing single crystals of high-purity beryllium, Be-4.37pct Cu, and Be-5.24 pct Ni. These crystals were prepared for testing in compression parallel to the [0001] by orienting and lapping to within ±3' of arc of the (0001). Microstrain testing apparatus is described along with c axis compression results for ingot purity beryllium, twelve-zone-pass material, and the above-mentioned alloys. Results show no measurable plasticity for the ingot purity material from -196" to 400°C, although some surface traces of (1122) slip was observed at 200°C and above. The twelve-zone-pass material shows substantial microstrain plasticity at 220°C with slip on (1122). Both alloys show significant plasticity at room temperature and above with slip also on (1122) planes. THE two slip systems which normally operate during the plastic deformation of beryllium in the vicinity of room temperature are:' basal slip (0001)(1120) and prism slip . Pyramidal slip with a vector inclined to the basal plane has been reported for elevated temperatures,'-a but occurs near room temperature only at very high stresses.~ A summary of the available data on the effect of temperature on the critical resolved shear stress for slip on these systems has been compiled by Conrad and Perlmutter.~ It has been postulated6'7 that one of the principal factors contributing to the brittleness of poly crystalline beryllium at temperatures below about 200°C is the difficulty of operating pyramidal slip with a vector inclined to the basal plane. Hence, detailed information on the operation of such a slip system is important to understanding the brittleness of beryllium. The operation of pyramidal slip with a vector inclined to the basal plane is best accomplished in beryllium by compressing single crystals in a direction parallel to the c axis. In such a test the resolved macroscopic shear strzss on the basal and prism planes is zero and (1012) twinning which is favored by tension along the c axis does not occur. Hence, in c axis compression of beryllium the normal deformation modes are inhibited and the operation of pyramidal slip with a vector inclined to the basal plane is favored. In the present investigation, c axis compression tests were performed on beryllium single crystal as a function of temperature (77" to 700°K), purity (commercial and twelve zone pass), and alloy content (4.37 wt pct Cu and 5.24 wt pct Ni). Presented here is a description of the test techniques employed and the gross mechanical behavior observed. A detailed analysis of the slip traces developed on the surfaces of the deformed specimens during these tests and the results of electron transmission studies of the deformed crystals are given in a separate paper.B PROCEDURE 1) Materials and Preparation. Single crystals about 1 in. diam were prepared of the following materials: commercial-purity beryllium, high-purity beryllium, and two beryllium alloys, one with 4.37 wt pct Cu and the other with 5.24 wt pct Ni. The commercial-purity single crystals were obtained by cutting specimens from large-grained ingot of Pechiney SR material, which is approximately 99.98 pct pure. The high-purity crystals were prepared by floating-zone refining (twelve passes) a rod (7 in. by 1 in, diam) of Pechiney SR grade cast and extruded beryllium. Although an absolute chemical analysis of the zone-refined material was not established, mass spectro-graphic analysis, emission spectrographic analysis, and y activation analysis indicated that it contained in atomic fractions about 5 to 10 ppm each of carbon and oxygen, 1 to 5 ppm each of nickel and iron, and about 1 to 2 ppm of copper, with the remaining residual impurities being less than 1 ppm. Further indication of the purity of this material is provided by the critical resolved shear stress for basal slip, which was approximately 300 psi. The starting material for the alloy single crystals was 1-in.-diam floating-zone-refined (six passes) rod of Pechiney SR grade beryllium. Two such rods were wrapped respectively with sufficient weight of wire of high-purity copper (99.999 pct) or nickel (99.999 pct) to yield a 5 wt pct alloy. A seventh floating-zone pass was then applied to each of the rods to accomplish the initial alloying and an eighth pass for homogenization. Analytical samples were taken from regions of the rod immediately adjacent to where the mechanical test specimens were cut; these indicated 4.37 wt pct Cu and 5.24 wt pct Ni. 2) Crystal Orientation. To avoid the occurrence of basal slip during c axis compression testing, it is necessary to load the crystals as nearly parallel to the c axis as possible. Preliminary c axis compression tests indicated that plastic flow and/or fracture occurred at stresses of the order of 300,000 psi; hence on the basis of a critical resolved shear stress for basal slip of 300 to 400 psi, the maximum crystal misorientation permitted is about 4 to 5' of arc. Since this accuracy cannot be obtained using the usual back-

Jan 1, 1969

By James K. Gluck, William E. Galloway, Gary E. Smith, John P. Morton, Christopher D. Henry

INTRODUCTION Surface mining and in situ leaching of uranium have the potential to alter ground-water quality around mines and leach sites. Of particular concern is the fate of uranium and its associated trace elements: molybdenum, arsenic, and selenium. We wish to under- stand the natural processes that control trace element concentrations in ground water and how these processes will influence dispersion of the elements from a mineralized zone, both naturally and during and after mining or restoration. For example, it is commonly recognized that the trace elements are soluble in oxidizing ground water but are insoluble, and can be precipitated, in reducing ground water. Thus oxidizing, metal-bearing water leaving a deposit could re- enter reduced ground, causing the water to be re- reduced and the trace elements to be, reprecipitated. In a sense, this is recreating the original mineralization process. To accomplish the above goals, we have (1) examined the theoretical controls of concentrations based on the available geochemical and thermodynamic data, (2) determined the major ion composition and oxidation-reduction status of Oakville waters because of the influence of these factors on trace element solubility, and (3) determined trace element concentrations and distribution in Oakville ground water. The last approach is used to evaluate how well actual behavior follows predicted behavior. This report focuses on two elements, uranium and molybdenum, because they exemplify the results obtained. The report also is restricted to a regional study of Oakville ground water. Results of more de- tailed study in and around major uranium districts in the Oakville and much of the raw data that support the conclusions in this report are presented in Galloway, Henry and Smith (1980). This report is part of that larger study, which concerned the depositional systems, hydrology, and geochemistry of the Oakville. The U.S. Environmental Protection Agency funded the study, under grant numbers R-805357-01 and R-805357-02. Theoretical controls were determined by reviewing the available literature on aqueous chemistry and behavior of uranium and molybdenum. To aid in under- standing water chemistry, Oakville water analyses were run through a modified version of the computer model WATEQF (Plumer, Jones, and Truesdell, 1976). WATEQF calculates speciation of dissolved ions and determines saturation with respect to a variety of minerals. In the discussion below, ion activity products (IAP) are compared with the equilibrium constant (KT) for various reactions and mineral products. Values of log IAP/KT near zero indicate that the water is in equilibrium with a mineral. Values less than -1 indicate considerable undersaturation and values greater than +1 indicate oversaturation. Galloway, Henry, and Smith (1980) give a more complete discussion of the application of this approach to Oakville water chemistry. Eh-pH diagrams have been constructed or adapted from the literature to predict what form -- dissolved ion or stable mineral species -- uranium and molybdenum assume under various conditions. Construction of the diagrams has followed procedures described by Garrels and Christ (1965). This approach is particularly appropriate because the solubility of the elements is Eh-dependent, and Eh varies greatly within the Oakville aquifer. A number of assumptions or approximations are inherent in the use of Eh-pH diagrams and chemical models such as WATEQF and in the interpretation of water chemistry in general. Both Eh-pH diagrams and chemical modeling rely entirely upon available thermo- dynamic data, including free energies of formation and dissociation constants for various reactions. These values are known to varying degrees of accuracy. Most major ions and minerals are relatively well control- led; however, data for trace metals are much poorer. Thermodynamic data are not available for some minerals, and for other minerals, two or more divergent values exist. By necessity, we have relied on the judgment of others to evaluate thermodynamic data. Calculations by WATEQF and constructions of Eh-pH diagrams are based on an assumption of equi1ibrium. Equilibrium may not be comnon in low-temperature aqueous environments; at best, ground-water composi tion may be in a state of dynamic equilibrium, continuously changing due to changes in environmental conditions. Eh-pH diagrams show what phases are stable at equilibrium under given conditions; they do not prove that the phases actually exist. Many minerals persist or form metastably under conditions outside their equilibrium stability field. The kinetics of reactions, which cannot be evaluated here, are important in determining what phases occur. Kinetics may be less of a problem for ground water that travels and evolves slowly through a semihomogeneous matrix than for many other natural systems. Eh-pH diagrams show equilibrium fields only of phases included. They do not indicate anything about stability relative to phases not included in the diagram. WATEQF, obviously, cannot calculate the degree of saturation of a mineral not included in the program or for which the appropriate ions were not analyzed. Thus, a mineral that was not considered may be the most stable phase under a given set of conditions and may control the solubility of a trace element. Also, this study is limited exclusively to in- organic compounds. Organic material is known to be an

Jan 1, 1980

By J. H. Polhems, R. B. Brackin, D. B. Grove

THE heavy media separation process plays an outstanding role in the concentration of 4000 tons of zinc ore per day at the Mascot mill of the American Zinc Co. of Tennessee. Of the total tonnage, 72 pct is treated in the heavy media separation plant to reject 56 pct of the ore as a coarse tailing, which has a ready market. Concentrates from this separation are beneficiated further by jigging and flotation. Approximately 25 pct of the total zinc concentrate production is made in the jig mill. Jig tailings are ground and pumped to the flotation circuit where the balance of the production is made. Fig. 1 shows a generalized flowsheet of the mill. The Mascot ore is a lead-free, honey-colored sphalerite in dolomitic limestone, with lesser amounts of chert and some pyrite. A mineralogical analysis is given in Table I. After 10 years of successful operation with galena medium and treatment of nearly 10,000,000 tons of ore, a decision to convert to ferrosilicon was made early in 1948 because of the increasing price of galena and consequent high operating costs. The conversion was made on Nov. 6, 1948, and the results obtained since that time have shown remarkable improvement over those made with galena. The Table I. Mineralogical Analysis of Mill Feed, Pct Calcium carbonate 49.5 Magnesium carbonate 35.2 Iron oxide and aluminum oxide 1.5 Zinc sulphide 4.5 Insoluble 9.3 100.0 Table II. Comparative Data, Galena and Ferrosilicon Ferro- Diner-Gelenaa siliconb ence Operating costs per ton milled, ct. 21.21 9.12 12.09 Medium consumption per ton milled, lb 0.80c 0.15 0.65 Reagent consumption per ton milled, lb 0.45 0.02 0.43 Tailing assay, pct Zn 0.310 0.297 0.013 Concentrate. oct Zn 12.08 10.33 1.75 Heavy medla ieparatlon recovery. pct 89.38 90.22 0.84 Mill feed rate, tons per hr 153 166 13 Heavy mesa separation feed rate. tons per hr 100 10 0 Tons milled per heavy media separation man shift 350 620 270 Mill feed to coarse tailings, pct 51.0 56.7 5.7 Lost mill time, pct 5.6 5.0 0.6 Power consumption, kw-hr per ton 2.06 1.92 0.14 a 1947. " First 6 months of 1950. c Net consumption after deducting credit for reclaimed waste galena. Consumption of new galena was 1.320 lb per ton milled. For entire life of galena operation, a credit of 40 pct of the value of the new galena added was realized from the sale of waste galena. comparisons given in this report cover the first 6 months of 1950 as representing the ferrosilicon operation, and the year 1947 as representing the galena operation. This was the last full year in which galena was used exclusively and is representative of the best work done during the 10 years of operation with this medium. After only 2 years' operating experience, with ferrosilicon and treatment of 1,807,585 tons many advantages have been revealed and are summarized in Table 11. Development Prior to the introduction of the heavy media process, all the mill feed was crushed through 5/8 in. and treated by jigging. A finished tailing assaying 0.66 pct Zn was made on rougher bull jigs, and cleaner jig tailings were ground for treatment by flotation. The first test work on the sink-and-float method of mineral beneficiation was carried out at Mascot in 1935, using a 3-ft cone and galena medium for batch tests. The following year a 6-ft cone was installed for pilot-plant work. This unit became a part of the mill circuit on March 1, 1936, and handled a gradually increasing tonnage in the next 2 years as the process developed to the point where it could treat all the + 3/8-in. material in the mill feed. Coarse jigging was then discontinued on March 1, 1939, and all coarse tailings have been made by the heavy media separation plant since that time. Feed Preparation: The original feed preparation plant consisted of a drag washer followed by two 4x10-ft Allis-Chalmers washing screens. A surge bin and two additional 5x12-ft AC washing screens were added in 1943. Use of primary and secondary washing screens was found essential to provide the cleanest possible feed for the cone and thereby avoid excessive contamination of the galena medium. Improved washing was obtained by replacing the drag washer with a 7x20-ft Allis-Chalmers scrubber, shown in Fig. 2, which has been in service since May 1944. Throughout the life of the galena operation, delivery of extremely muddy ore to the mill overloaded the medium cleaning system, and it frequently was necessary to cut off the feed and clean the medium for several hours until its normal viscosity had been re-established. The cleaning circuit

Jan 1, 1952

By E. R. Borcherdt

IT is notable that the first large-scale mine operation equipped entirely with detachable bits was the Badger State mine of the Anaconda Copper Mining Co. in Butte, Montana, just 30 years ago. This mine in 1922 was producing approximately 1200 tons of ore per day. Much of the data presented in C. L. Berrien's article' describing the development and installation of the Hawkesworth detachable drill bit were obtained from these operations. As in any pioneering effort, no precedent existed and many difficult problems required solution, so that the changeover to detachable bits at all Butte hill mines was not completed for 6 years. There was widespread disbelief as to the probable efficiency of the new installation. Some attempts were made in 1931 by the owners of the Hawkesworth patents to interest Ontario gold mine operators in the bit. These efforts were not successful, but they undoubtedly stimulated thinking which resulted in the invention and patenting of several well-known Canadian detachable bits, one of which is now a widely used throwaway bit. The success of the Butte installation also led to the development of the threaded type of bit connections by several well-known manufacturers, and in 1935 these bits were introduced to the mining industry on a national scale. The original Hawkesworth bit was not provided with a water hole but, depended upon water passing through the clearance opening between the tongue in the bit and the groove in the rod to flush cuttings from the drill hole, see Fig. 1. In December 1935 it was found that this method of introducing drilling water to the bit face resulted in high dust counts. To correct this a water hole was drilled on the central axis of the bit, passing through the tongue. Unfortunately, quenching water would rise through the small water hole, spot-hardening the tongue to cause breakage, never completely eliminated. In the fall of 1936 large-scale tests indicated that savings would be effected by use of a threaded type of bit, which was therefore adopted as standard for all Butte mines. This type of bit was used until 1947, when it was superseded by a one-use slip-on type. Since the first use of the Hawkesworth bit every detachable bit of importance has been investigated, and where advantages which might reduce costs or increase efficiency were indicated, substantial tests of the bit were carried on in the Butte mines. When tests demonstrated the advisability of changing from one kind of detachable bit to another the change was made at one level or in one area each day until the new rod and bit equipment was used throughout the mine. This involved a minimum of cost and disruption of drilling. Intelligent selection of a detachable bit to obtain optimum results requires careful consideration to achieve a balance between the three principal types of equipment used in the drilling process: 1—drill bits, 2—drill steel, and 3—drilling machines. Optimum results imply maximum output and minimum cost per unit of output. Since every rock type differs in drillability and it is generally impractical to provide equipment for more than one or two types of rock which may occur in one operation, selection of equipment must encompass average drilling conditions. However, on exceptional occasions several widely differing conditions may make it mandatory to provide equipment best suited to each condition. The choice of rock-drilling equipment is a most controversial subject and one that is further complicated by unreliable and frequently misleading performance claims. Small operators without the means for making accurate evaluations of equipment frequently suffer from these over-enthusiastic claims. It is apparent from experience in rock drilling throughout the world that rock drillability is not alike in any two places, and that selection of proper equipment can only be made after conducting thorough trials of various types of equipment. Some recent drilling tests in tactite and hornstone at the Darwin, California mine of the Anaconda Co. present some interesting clues on rock drillability. Microscopic examination of thin sections of these rocks reveals that mineral composition and rock texture are equally important in governing drillability. The Darwin hornstone is at times so abrasive that the carbide bit cutting edges become flattened to 3/32 in. in 2 to 4 ft of drilling, and some carbide bits were dulled to this point after 9 to 10 in. of drilling. This wear was determined to be the proper point for resharpening to eliminate carbide insert breakage or breakage of the steel rod when drilling with 1½ to 1?-in. bits, with a drifter of 2 3/4-in. diam and 90 to 100 psi air pressure, see Supplement A. Before considering the merits of various bit designs it may be well to review the mechanics of drilling rock with percussion drills. A sharp bit cuts by penetration and chipping. The amount of penetration governs the amount of chipping and depends upon the contact area of the cutting edge, the foot-

Jan 1, 1954

By AIME AIME

PROBABLY no man has been of greater service to the Institute and has kept more in the background than Henry Krumb. A Vice-President continuously) for the last eleven years, apparently neither his picture nor a biographical sketch ever have adorned these pages and were he forewarned in the present instance he would order us to "forget it." He is a Columbia School of Mines man, Class of '98. He worked underground at Rossland, B. C. for a time, then for a year and a half a. chief engineer of the famous Camp Bird at Ouray, Colo. For three wars he was examining engineer for the Guggenheims and since 1901 has been an independent consulting engineer with experience throughout the Americas.

Jan 1, 1939

By J. W. Smith

Our continuous search for underground productivity improvements has been brought about by the diminishing ore grades in existing underground mines. The need for more efficient mining methods is a result of the economic problems facing our industry today, and this has caused us to evaluate underground haulage methods which have traditionally been the "bottleneck" in the flow of material from the ore in the natural state to the surface processing facility of any underground mining operation. Small improvements in the face haulage systems have yielded much greater benefits as they relate to overall mine productivity so it's only natural that we are all concerned with the best method of moving ore from the face to the main line haulage. In a recent paper titled "Underground Haulage Trucks - Gaining Momentum Worldwide", Richard A. Thomas concludes that the use of trucks to haul ores in underground mines is on the increase spurred by the convergence of a number of technology advances and economic realities. Perhaps the most important stimulus for the growth of trackless haulage is the high degree of haulage flexibility in underground operations. On the economic side, the demand for higher productivity from underground mines has resulted in larger physical dimensions of haulage roads, that is, higher backs and wider drifts to provide more room for high capacity haulage units. In the process of determining the most effective type of equipment for haulage, the power source must be a major consideration. For the purpose of this paper, we will limit the comparison to rubber-tired trackless haulage vehicles and not try to make a comparison between rubber-tired haulage, continuous haulage systems and rail-mounted haulage. Cost is perhaps the only really measurable factor when making a comparison between electric and diesel haulage. You will find that some costs will be very well defined in absolute terms. In other areas of comparison, cost can be fairly well estimated, and yet in still others, the costs are totally arbitrary. Let's take a look at some of the cost considerations. (Figure 1) first of all, is the initial cost of the equipment. This capital cost quite often is a determining factor in the type of haulage vehicle to be selected, yet this initial cost is perhaps the most insignificant of all costs when evaluating an operation over the long term. Of much greater concern, is the cost of maintenance. This cost will often run three times the original capital investment during the life of a single piece of haulage equipment. This factor can include rebuild to extend the life of the original capital investment, but certainly includes the labor and materials necessary, plus the inventory to keep the equipment in good repair. Perhaps one cost which is now playing an even greater role in the rubber-tired haulage operation, is the cost of fuel. Conoco has recently come up with some rough estimates which indicate that diesel fuel will cost an average of three times the equivalent kilowatt output in direct electric power. Diesel fuel is almost twice the cost of stored electric power. (This of course relates to the efficiencies of charging and recovery of power from lead acid storage cells.) These particular figures of course will vary from one area to another but I think that there is enough significance here to certainly warrant the further study of fuel costs for each particular area or mine. Another cost is breakdown expense. This must be treated differently from maintenance costs because a potentially larger expense is involved, more than just parts and labor. Now we have to deal with the cost of lost production time, which can have a much greater overall effect. Mine plan economics are another cost consideration where we can't make a comparison without looking at specifics. Here you must look at the movement of power centers vs. the flexibility and freedom of movement of vehicles. The determination must be made as to what types of equipment will fit into any predetermined mine plan and if a change in the planned roadway dimensions for the mine plan itself would be more economical so that more efficient type of equipment could be utilized. Finally, two of the most important aspects to be considered with potential ramifications far beyond what we have mentioned previously, is the cost of health and safety, which is really the cost of meeting current and future government regulations, reasonable or otherwise. And of course, when making any consideration here it is impossible to come up with anything more than an educated guess on the cost of meeting the new regulations. Now let's take a look at some of the advantages of diesel vehicles as well as advantages offered by electric vehicles, both battery and cable powered versions (Figure 2). Much of the data used in this comparison is based on experience with three vehicles manufactured by Jeffrey Mining Machinery Division, Dresser Industries. Jeffrey manufactures all three types, each with approximately a 15-ton capacity, even though few of these Jeffrey vehicles are used in uranium mining operations. Much of our experience comes from the 4114 diesel powered RAMCAR which is a 4-wheel drive, articulated steering,vehicle powered by a Caterpillar 3306NA engine and using a powershift transmission. This will be compared with the performance of the Jeffrey 404H battery powered RAMCAR with articulated steering which utilizes a separate 35 HP DC drive motor on each of two wheels with solid-state speed controls, and the final comparison will be made on the Jeffrey 4015 cable-reel shuttle car which is powered by two 60 HP constant

Jan 1, 1982

By A. E. Lee

THE Blackwell Zinc Co., Inc., a subsidiary of the American Metal Co., Ltd., operates a horizontal retort zinc smelter at Blackwell, Okla. The plant has 14 furnace blocks of 800 retorts each, fired with natural gas on a 48 hr cycle. Over 13,000 tons of zinc-bearing material, chiefly sulphide flotation concentrates, are treated monthly to produce slab zinc and high lead-cadmium fume. In 1942 a program of rebuilding and modernizing the smelter was started. By 1947 furnace smelting capacity had been increased to a point where roasting and sintering facilities were inadequate, and it was necessary to purchase oxidized materials to supplement sinter production. The seven 210 ft Ropp roasters and three 42 in. x 44 ft Dwight-Lloyd machines then in use had been in service at least 20 years and were in need of major rebuilding. Thus it was entirely practical to consider all new equipment and a change of method rather than rebuilding and repairing obsolete units. A study of the problem indicated that roasting as such could be eliminated and roasting and sintering accomplished in one step by a modification of the Robson process,' which had been used since the early 1930's by the National Smelting Co., Ltd., at their plants at Avonmouth, England, and Swansea Vale, South Wales. Francis P. Sinn, General Manager, Zinc Smelting Operations, The American Metal Co., Ltd., who was familiar with the practice in England, suggested the use of one large machine for the entire operation from concentrate to sinter. One step sintering appeared to best meet Blackwell's plant requirements and indicated substantial savings in labor, gas, coal, and repair costs. Choice of Machine Size The sinter machine size was set at 12x168 ft for a rated capacity of 540 tons per day. This tonnage, produced on a five day week, would meet the seven day requirements of the 14 furnace blocks. The one large machine was quoted at a lower cost than two or more 6 ft wide machines of similar total capacity. Further, the larger machine could be housed in a smaller structure and only one set of equipment for charge preparation and delivery and for disposal of sinter cake was needed. One machine on a five day week made possible a concentration of the skilled operating personnel and required less men than a plant including two or more machines and related equipment circuits. Fewer units of equipment meant less maintenance, and the two down days weekly allowed ample time to repair and, if necessary, to make up lost production. Experience had indicated better sintering quality and rates with larger masses of material, not only on wider machines, but also in deeper beds. The ratio of windbox perimeter to area for the 12x168 ft machine is 0.179, compared to 0.353 for a 6x102 ft machine and 0.617 for a 42 in. x 44 ft machine. This meant less air leakage with resulting fan power savings and less spoilage of charge along the pallet sides. Performance Initial operation of the new sinter plant was made in November 1951 and regular production attained late in December. The average product sinter output during 1952 and the first half of 1953 has been 18.2 tons per hr. The average for one month has been as high as 22.4 tons per hr. Considerable experimenting with varied operating conditions accounts in part for the below capacity — 24 tons per hr — average output, and work to further improve production rate continues. A typical sinter analyses is 66.0 pct Zn, 0.3 pct Pb, 0.1 pct Cd, 0.3 pct S, 8.0 pct Fe, 2.0 pct SiO,, 0.8 pct CaO, and 0.2 pct MgO. Use of this material has made possible increases in furnace burden and improved furnace operation over the former practice using sinter made from Ropp roasted concentrates. Better lead and cadmium elimination in sintering has permitted the furnace production of slab zinc lower in lead and cadmium. Anticipated economies of operation have largely been gained. The sinter plant is operated by seven men per 8 hr shift — one head operator, three equipment operators and three sweepers — plus one oiler on day shift only. While it has been necessary at times to operate seven days a week to produce the required sinter tonnage, the five day work week usually has been adequate. Consumption of natural gas for sinter bed ignition is 200,000 to 300,000 cu ft per day. Green Ore Sintering Practice The 30 to 31 pct sulphur content of the —200 mesh zinc concentrates is the fuel used to sinter the charge, no coal addition being required. In the feed to the machine, sufficient concentrates are added to crushed return sinter fines containing 0.3 to 0.5 pct sulphur to produce a charge averaging 5.0 to 6.5 pct sulphur. Since the return sinter used in Blackwell's practice is varied from — 1/2 to — 1/8 in., the actual sintering mixture of fine sinter and concentrates is somewhat higher in sulphur. The coarser sinter particles are too large to resinter and merely aid porosity in the sinter bed. The ratio of concentrates to return sinter in the charge ranges from about 1:4 to 1:5.5. Variations are based on the appearance of pried up bed sections, bed exit gas temperature trends, windbox suctions, and return sinter size. Sufficient sulphur must be used to obtain fritting of the charge into a soft sinter cake and to aid in the elimination of lead and cadmium. Excessive feed sulphur will result in partial slagging of the cake impairing porosity and prolonging sintering time.

Jan 1, 1954

By A. J. Mortlock

Cobalt and nickel have been diffused at tracer concentrations in gold at several temperatures in the range from approximately 700° to 950°C. The diffusion penetration profiles were determined by a serial sectioning technique in which the gold is first anodized and then the anodic layer is dissolved in acid. In this ulay sections as thin as 250A could be removed reproduci-bly. In all cases, the region close to the specimen surface was characterized by irregular behavior in the sense that the logarithm of concentration was not linear in the square of the penetration distance. In sotne cases, there zuas an indication of the operation of very slow dijfusion in this region, while in others the apparent diffusion coejj'icient was negative. Possible reasons for this anomalous behavior are briefly discussed. In recent years it has been found that the region close to the surface of a metal can sometimes exhibit anomalously slow diffusion characteristics relative to the interior of the metal. One of the best examples of this fact is the work of Styris and omizuka,' who showed that the apparent diffusion coefficient for zinc in the region withi: about 1 p of the free surface of copper was about ,,,, that at deeper penetrations. This result is particularly interesting, because it is free from the possibly complicating effects of low solubility of the diffusing tracer in the solvent metal. In the case of diffusion under conditions of low solubilitjr, interpretaticn of the results in terms of lattice diffusion is difficult because of the enhanced short-circuiting produced by segregation to dislocations.2'3 Measurements by Duhl et 1. suggest that cobalt diffusing in gold may also show a near-surface effect of this type. Once again the solubility is high, so that this result could be of great interest. However, the technique used for analyzing the diffusion penetration zones by Duhl, viz. the counting of residual gamma activity in the specimen following sectioning, appears to have indicated a near-surface effect in a parallel experiment on the self-diffusion of gold reported at the same time. The latter result is known to be spurious, since Kidson5 has demonstrated that self-diffusion in gold does not show this effect. Duhl et 01. also reported some measurements on the diffusion of nickel in gold, but failed to give any data for the near-surface region. As the solubility of nickel in gold is high, such data would also be of special interest. We, therefore, decided to conduct another set of experiments on the diffusion of nickel and cobalt in gold, using a sectioning technique that allows the individual sections to be assayed for solute content and thus gives direct determinations of penetration profiles. Also, by sectioning with an anodizing/stripping tech- nique, very thin layers can be removed and the region close to the surface studied in detail. MATERIALS The gold specimens were supplied as single crystal disks $ in. in diam by a in. high by Monocrystals Co. of Cleveland, Ohio. The gold itself was of spectro-scopic purity, i.e., better than 99.99 pct pure. METHOD Specimen Preparation. One flat end face of each gold crystal was spark planed with a Servomet spark erosion machine set for minimum spark energy. Following this treatment the crystals were preannealed for 2 to 4 days at temperatures of either 400" or 700°C. The three crystals preannealed at 700°C showed signs of recrystallization. The spark-planed end face of each crystal was then coated with the appropriate amount of 63i or 60 radioactive tracer. This deposit was laid down in a simple plating bath containing the as-supplied solution of the radioactive isotope as well as sufficient ammonium oxalate to saturate the solution. Some ammonium oxalate remained undissolved on the floor of the bath for this purpose. During plating further additions of ammonium oxalate were sometimes required to allow the plating to continue satisfactorily, perhaps due to passivation of the undissolved oxalate already present. The thickness of the deposited layer was determined by comparison of the apparent surface activity of the plated specimen with that of a similar specimen having a weighable deposit of the isotope on its end face. Correction for self-absorption of the radiation was made in this calculation. Annealing. The deposited crystals were annealed in a hydrogen atmosphere in sealed silica tubes. During this heat treatment they were supported, active face down, on optically flat silica plates. The temperature was measured with calibrated Pt vs Pt-10 pct Rh thermocouples, and the tabulated values can be taken to be correct to Z°C. All the crystals showed evidence of recrystallization following these heat treatments, suggesting that initially they may not have been good single crystals or had suffered strain during delivery. Concentration Profile Analysis. After annealing, the crystals were sectioned by the anodizing-stripping technique.6 The anodizing involved suspension of the specimen with its cylindrical axis horiz6ntal by a gold wire in a 200-ml beaker containing 1 M Hg304. A cathode in the form of a strip of gold sheet, 2 in. wide and positioned to be in contact with the curved side of the beaker, completely encircled the specimen. An anodizing current of 30 ma, corresponding to a current density of 5 ma per sq cm on the surface of the specimen, was passed for times ranging from 5 to 150 min depending on the thickness of gold to be removed; the solution was stirred continuously during this process. Following this treatment, the specimen

Jan 1, 1969

By R. D. Lord

The search for the best practical means of storing sulfide bearing tailings, where there is no residual excess of carbonate material is discussed in this paper• Usually the sulfide content decomposes, with the aid of bacterial action, and the resulting sulfuric acid escapes, along with any heavy-metal solutes, through embankments that are usually porous to some degree• The problem is typified in the tailings of the uranium operations of Elliot Lake, Ont., where mining started some 20 years ago• The approach to tailings disposal paralleled the practice for other hydrometallurgical plants treating gold and base-metal ores• Impoundment areas were designed to retain solids, and a clear and neutral overflow was considered satisfactory practice• Now experience has shown that these areas, some of which have been idle for over a dozen years, release acids in seepage and overflows to an unacceptable degree• To protect natural water courses, neutralizing plants are operated wherever required• Lime slurry is fed continuously into the tailings outflows in a quantity sufficient to raise the pH to 8•5 and precipitate heavy metals that may be in solution• The objection to this procedure is that the plants will require servicing indefinitely, unless a better remedy is found• The problem differs only slightly from that common to base-metal concentrators in that here the ore has been leached with sulfuric acid for the recovery of uranium• Any native content of calcareous material has been digested, and only that added for final neutralization is available to maintain a pH unfavorable to bacterial activity• Chemical oxidation slowly lowers the pH and when this reaches a level of 4•5 or less, bacteria become active and greatly accelerate the formation of acid. The bacterial process is probably at least ten times as fast as the chemical oxidation• Location and Processing The operations referred to, uranium and one copper mine, are located at approximately 46°N and 82°W longitude• This is typical Canadian Shield country, a land of lakes, deeply glaciated and rocky, with sparse soil which supports mixed forest cover• Drainage is to Lake Huron, 25 miles to the south• Average temperature is 45°F, ranging from -40° to +95°F• Annual precipitation is 38 in•, about half of which is snow• The ore is Precambrian, quartz-pebble conglomerate, with mineralization in the matrix• From 5 to 10% pyrite is present• All known means of pre-concentration have been tested, but a bulk sulfuric acid leach has proved the most efficient. Tailings have from the outset been neutralized before release• Current practice is to add ground limestone to bring the pH to 4•5, and then lime to raise the value to 10•5• Environmental regulations have recently been increased and the foregoing meets the new standards• Separate measures are taken to precipitate radium• Remedial Measures Since the outstanding environmental problem is the oxidation of pyrite by bacterial action, the solution is to contain the products, or arrest the process• Given the ambient temperature, favorable half of the time, four items are essential to the activity• 1) Pyrite• 2) Moisture pH < 4•5. 3) Oxygen• 4) Bacteria• Removing any one of these out of the range of tolerance will bring the reactions under control• A variety of proposals considered, and a number tested for the arrest of the process, are: (a) render embankments impermeable, (b) provide an impermeable cover, (c) cover with an oxygen absorbing layer, (d) provide a vegetative cover, (e) flood the site, (f) remove pyrite from current tailings, (g) add excess limestone to current tailings, (h) poison the bacteria• Bank Seal-On existing impoundment areas, where the embankments are several thousand yards in length, it is believed that any program of injecting sealants can have small chance of success• However, a moisture barrier is an indicated specification for future construction, and this can be highly expensive• Surface Seal-Depending on the configuration of the deposit, the downward travel of water should be prevented, and oxygen excluded• Burying a plastic membrane just below the surface has been considered, as has the application of a liquid sealant that would penetrate the surface. The objection to these remedies is the excessive cost of dealing with large areas and the expectation of only temporary benefit as a result• Frost penetration is over 4 ft, and frost action breaks up asphalt paving and all but heavy concrete in a few years• Organic Layer-An oxygen-absorbing layer, such as bark fines from paper mills has been proposed as a surface treatment• Cultivated into the tailings such material might be expected to arrest subsurface oxidation for some years• Estimates are 100 tons per acre of bark fines, or 35 tons per acre of sawdust, and these enormous quantities do not so far give assurance of providing a long-term remedy• Vegatative Cover-Several obvious benefits would result from a good growth of grass or other vegetation on abandoned tailings• While restoring the natural green of the tract the growth would prevent wind-blown dust and reduce erosion• Subsurface oxidation should be reduced, as well as the upward movement of ground moisture as occurs in dry weather. To this end, considerable research and field testing has been carried out to arrive at a formula - a prescription which will provide a self-sustaining growth on the tailings surface, or at least one that would survive with reasonable maintenance attention. Many test plots have been run with different combinations of surface treatment and seed mixtures. Generally, by addition and close cultivation of limestone, lime, and fertilizers, technical success has been demonstrated• Plants with a high tolerance for acid soil seem the more hardy, and a pH above 3 is indicated so that nutrients can be absorbed• Recommendations are for 12 to 15 tons of

Jan 1, 1977

By M. S. Azun

M.S. Azun I have many objections to the content of the author&apos;s paper. Before discussing it, however, I would like to repeat the property of semivariogram function. Second order stationary properties of regionalized variables (ReV&apos;s) such as semivariogram function ?(h) are perfectly known in geostatistics. Also, the kriging equations in the language of mathematical statistics using second order stationary properties are well understood. However, the way to use the sample (estimated) semivariogram function in any one of the kriging procedures is vague. The sample semivariogram function is given as follows: [1 N-hy*(h) = 2(N h) i21 {Xi-Xi+h}Z, h=0, 1, N-1] where N is the total number of samples, Xi is the sample value at the i - th location, X i+h is the sample value at the i +h - th location, and h is the distance among the samples. An estimation variance of sample semivariogram function of first lag is smaller than that of higher order lag. The theoretical semivariogram function reaches the variance of samples asymptotically. But this is not easily observable because of the larger variation involved in the estimate of semivariogram function. In general, an estimation procedure is done for h = 0, 1, 2,…., up to the greatest integer less than N/2, even though sample semivariogram function can be computable through N-1. After estimating semivariogram function, the critical question of how to model sample semivariogram function arises. As seen in the above equation, sample semivariogram function is discrete and can be smoothed by the model being selected. Therefore, modeling of sample semivariogram function is the most important step in geostatistics. It not only smoothes a discrete function but also affects the results of the kriging procedure. When the only aim is to model the semivariogram function, which is the basic point of the author&apos;s paper, one can employ any fitting techniques, such as curve fitting, or any ar¬bitrary functions, which are called submodels in the paper. The term "arbitrary function" is used rather than "submodel" because there is no basic understanding of developing them. The author suggests that the sum of those submodels can also be used for the modeling of sample semivariogram function. The combination of any arbitrary functions brings many problems instead of giving an insight of the domain structure considered. The author used two arbitrary functions and the nugget effect in response to sample semivariogram function (Fig. 10). For the same example, he stated that the parameters involved in the mixed arbitrary function model can be accepted when the discrepancy between sample semivariogram function and the model is small visually. For verifying the fitting behavior of any selected model, one should not be contented with the visual satisfactory. Some statistical measure such as goodness of fit has to be used. The author&apos;s practice is no more than an exercise in curve fitting without any fundamental understanding or conceptualization of the underlying physical mechanism. Furthermore, the selection of any model is not an easy task if the purpose is the search for the "best" response to the observed second order properties of ReV&apos;s. I suggest that the Markovian model (Azun, 1983), on the basis of a theoretical understanding of underlying mechanism, which gives more information about the occurrence of regionalized variables, is used to respond all properties of ReV&apos;s. There are a lot of problems for modeling of onedimensional sample semivariogram function. Thus, it is not appropriate to go to higher order dimensional sample semivariogram function modeling. In the meantime, I would recommend that one can connect the values of standardized sample semivariogram function rather than simple values of semivariogram function in the two-dimensional estimation. The standardized values can be computed in dividing the semivariogram function value by the number of sample pairs involved in each lag regardless of the directions. In conclusion, geostatistics is an interdisciplinary area in mining that uses the principles of mathematical statistics. Thus, it should not violate any probabilistic and statistical rules. When Matheron was developing the theory of geostatistical study in the early years of geostatistics, many mining people had a reservation accepting the geostatistical tools. However, this does not mean that we, the geostatisticians, might try to convince those people using some "strange" tools or rules as some authors implied (Baafi and Kim, 1984). Instead, we have to develop and explain the geostatistical tools staying only in the framework of statistical concepts and properties. ? References Azun, M.S., 1983, "Stochastic Process Modeling of Spatially Distributed Geostatistical Data," Columbia University, Ph.D. Thesis. Baafi, E.Y., and Kim, Y.C., 1984, "Discussion - Comparison of Different Ore Reserve Estimation Methods Using Conditional Simulation," Mining Engineering, Vol. 36, No. 3, p. 280. Reply by J.M. Rendu The interactive method proposed by Rendu allows practitioners to develop semivariogram models that take into account not only the numerical information obtained by sampling, but also highly significant additional information that often cannot be quantified. The geology of the deposit - including hypotheses concerning its genesis, sampling methods, assaying methods, and mathematical methods used to calculate the semivariograms - all have an influence on the numerical results obtained and on how these results should be interpreted. If all the information concerning the spatial distribution of values in a mineral deposit was contained in the sample values, it could be argued that statistical techniques alone would produce optimum models. However, this is rarely, if ever, the case. Methods that allow the user to take into account his experience and his geologic understanding of the deposit should not be rejected for the sake of theoretical statistical purity. ?

Jan 1, 1986

By Gilbert Soler

Producers of alloy steels recognize the importance of chemical composition in relation to the hot-working properties and the typical surface defects found in their product. Each analysis of steel has its own peculiar characteristics. Under conditions of standard mill practice each analysis is susceptible to certain types of defects. Mill practice must be varied to obtain the best combination of surface and internal quality in the product. Chemical composition influences the cast structure and crystallization characteristics of the ingot. It also determines the rate of heating and cooling, the plastic hot-working range, and the phase structure of the steel at various temperatures. as well as the tendency toward scale formation and decarburization. This paper endeavors to emphasize the manner in which chemical composition affects the various properties of steel, and to indicate the relative importance of these factors in relation to the hot-working properties and surface characteristics of killed steel. The influence of chemical composition may be outlined as follows: I. Effect on the cast structure of steel, including: A. State of deoxidation, and type of inclusions. B. Gas content of steel. C. Freezing point and melting point of steel. D. Crystallization characteristics and segregation. 11. Effect on the hot-working properties and surface characteristics of steel, including: A. Plastic hot-working range. B. Phase structures at hot-working temperatures. C. Rate of heating. D. Cooling characteristics. E. Scale formation. F. Surface decarburization. EFferect of Chemical Composition on Cast Structure of Steel State of Deoxidation and Type of Inclu-sions.—The state of deoxidation is limited by the final chemical analysis desired in the finished product, and is controlled primarily by carbon, manganese, silicon, and aluminum, and to a lesser degree by chromium, titanium, vanadium, or other deoxidizing elements. The degree of deoxidation affects the density of the cast structure and broadly classifies the steel as killed, semikilled, or rimming. This in turn manifests itself in surface characteristics. The manner and extent of deoxidation also controls the amount, type, and distribution of nonmetallic inclusions formed. The equilibrium of manganese, silicon and aluminum with the slags and pouring refractories with which the metal comes in contact is important, especially in regard to inclusions of fire-clay origin. Some surface Seams and hot-working difficulties can be traced to nonmetallic inclusions.

Jan 1, 1941

By Gilbert Soler

Producers of alloy steels recognize the importance of chemical composition in relation to the hot-working properties and the typical surface defects found in their product. Each analysis of steel has its own peculiar characteristics. Under conditions of standard mill practice each analysis is susceptible to certain types of defects. Mill practice must be varied to obtain the best combination of surface and internal quality in the product. Chemical composition influences the cast structure and crystallization characteristics of the ingot. It also determines the rate of heating and cooling, the plastic hot-working range, and the phase structure of the steel at various temperatures. as well as the tendency toward scale formation and decarburization. This paper endeavors to emphasize the manner in which chemical composition affects the various properties of steel, and to indicate the relative importance of these factors in relation to the hot-working properties and surface characteristics of killed steel. The influence of chemical composition may be outlined as follows: I. Effect on the cast structure of steel, including: A. State of deoxidation, and type of inclusions. B. Gas content of steel. C. Freezing point and melting point of steel. D. Crystallization characteristics and segregation. 11. Effect on the hot-working properties and surface characteristics of steel, including: A. Plastic hot-working range. B. Phase structures at hot-working temperatures. C. Rate of heating. D. Cooling characteristics. E. Scale formation. F. Surface decarburization. EFferect of Chemical Composition on Cast Structure of Steel State of Deoxidation and Type of Inclu-sions.—The state of deoxidation is limited by the final chemical analysis desired in the finished product, and is controlled primarily by carbon, manganese, silicon, and aluminum, and to a lesser degree by chromium, titanium, vanadium, or other deoxidizing elements. The degree of deoxidation affects the density of the cast structure and broadly classifies the steel as killed, semikilled, or rimming. This in turn manifests itself in surface characteristics. The manner and extent of deoxidation also controls the amount, type, and distribution of nonmetallic inclusions formed. The equilibrium of manganese, silicon and aluminum with the slags and pouring refractories with which the metal comes in contact is important, especially in regard to inclusions of fire-clay origin. Some surface Seams and hot-working difficulties can be traced to nonmetallic inclusions.

Jan 1, 1941

By D. Harker, M. J. Murphy

THE mechanism by which age-hardening takes place is still not completely understood. The principal theories range from the extreme of "precipitation-hardening" to that of "order-hardening," with many intermediate gradations. In the hope of obtaining new data on which to base a choice among the many theories, the authors have made electron micrographs of formvar replicas taken from metallic specimens at various stages of age-hardening. The results so obtained on an alloy of 2 per cent beryllium in copper and one of 20 per cent molybdenum in iron will be described in the following pages, as well as the technique used in preparing the replicas. PREPARATION OF FORMVAR REPLICAS OF METALLIC SPECIMENS This section is concerned with the techniques that have proved successful in preparing formvar replicas for studying the microstructure of metals with the electron microscope. Because of its simplicity and accuracy, the formvar replica1 method developed in the General Electric Research Laboratory has been used exclusively, since it is much simpler than the silica replica method of Heidenreich and Peck2 and produces as good results. The formvar method to be described requires a technique acquired only by practice, and no foolproof set of rules can be given. It is a rapid method and a replica can be obtained in less than five minutes under favorable conditions. This does not mean that perfect replicas can be produced every five minutes, but rather that one can well afford to take several strippings* in order to produce a replica of high quality. The metallographic technique required for replica work-very fine polishing and etching with very dilute solutions-is described here. It can be said, in general, that almost any surface that has been carefully prepared for a photomicrograph at 1000 or more diameters can be used to obtain a replica for the electron microscope. PREPARATION OF METALLOGRAPHIC SPECIMENS The preparation of the metallographic specimens is somewhat more exacting than that usually given to samples to be used for inspection under the light microscope. After much experimentation, it was concluded that the best method consists in polishing the metal through to the coarse cloth, then etching and repolishing on final cloth until the layer of distorted metal is removed. After this procedure has been completed, a dilute etch-to approximately the depth necessary for a micrograph at 1000 diameters-usually produces a sur-

Jan 1, 1945