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By P. S. Kotval, O. H. Nestor

Stacking fault energy measurements of nickel-base alloys have been mainly confined to binary and ternary systems. In this paper, the stacking fault energy has been measured by the rolling texture method in a series of ten alloys which comprise successive additions of Cr, Mo, Fe, and C to pure nickel, eventually resulting in an alloy of the composition of Hastelloy alloy X. The alloys studied here are single-phase, solid solutions with the exception of two alloys in which some undissolved particles of "primary" carbide have been retained. It is found that successive additions of chromium, molybdenum, and iron all lower the stacking fault energy, with iron having only a minor effect. The stacking fault energy is found to increase when carbon is added in solid solution. The results from the rolling texture measurements are correlated with thin foil observations of dislocation substructures in these alloys. In a recent paper' it was pointed out that the dislocation substructure of various superalloy matrices could be classified into three broad categories based on 'high', 'medium', and 'low' stacking fault energy. It has also been demonstrated2 that the dislocation substructure in each of these categories has a well defined role in the nucleation of strengthening precipitates which is different from the role played by the dislocation substructure in other categories. Thus, it becomes desirable to understand the influence of various solute elements on the stacking fault energy and hence on the dislocation substructure of the matrix, before any further development of superalloys by mi-crostructural predesign can be undertaken. Recently, Beeston and France have studied the influence of increasing solute additions on the stacking fault energy of a series of binary nickel-base alloys relevant to the Nimonic series using the rolling texture method, and have then estimated the effect of a given alloy addition in five commercial Nimonic alloys. However, comparison with stacking fault energy data from other investigations''5 suggests that the influence of a given solute element in a nickel-base binary system is not necessarily the same in a ternary or more complex superalloy system. Accordingly, the present work was undertaken to study the effect of successive addition of solute elements to pure nickel, the final composition being the nominal composition of Hastelloy X. The rolling texture method of stacking fault energy measurement was used since it can be used for the whole range of stacking fault energy values and does not have the disadvantage of, say, the Node method which is only applicable to low values of stacking fault energy. In addition, the rolling texture method provides a means of determining the stacking fault energy which is statistically more significant than that provided by other methods. EXPERIMENTAL TECHNIQUES Button heats of alloys of the compositions shown in Table I were prepared. Each button was remelted not less than four times. After a slight deformation (approximately 5 pct) all alloys were homogenized at 2200°F except alloys, H . I, and J. Alloys H and I were solution heat treated at 2150°F and alloy J at 2282OF. The buttons were cold worked by rolling, using "end-to-end" passes and intermediate anneals at the homogenization temperatures mentioned above. After each annealing treatment the samples were rapidly water quenched to avoid any precipitation. In alloys F and I, however, a few particles of "primary" carbides were retained even after the homogenization treatments at the temperatures mentioned above. Part of the solution heat treated material was cold worked to 0.04-in.-thick sheet and the penultimate reduction was -50 pct of deformation as recommended by Dillamore et al. All annealing was carried out in vacuo within sealed quartz capsules. Some of the material from each alloy was rolled down further to 0.004 in. strip for thin foil transmission electron microscopy specimens. Specimens of this strip were annealed at the homogenization temperature for 1 hr and then strained 7 pct by rolling at room temperature. Thin foils were prepared from the strip specimens by the 'window" technique using an Ethanol-Perchloric acid electrolyte at 32°F and a voltage of 22 v. Stainless steel cathodes were employed. All transmission electron microscopy was performed in a JEM-7 electron microscope using an accelerating voltage of 100 kv. Specimens from the 0.04 in. sheet which had been rolled -60 pct in the final pass were electropolished to remove the surface layers to a depth of approximately 0.002 in. Rolling texture pole figures for all the alloys were determined using a Schulz ring and nickel filtered CuKa radiation at 50 kv and 20 ma. The texture parameter Io/(lo + I,,) (where Io is the

Jan 1, 1970

By C. E. Sims, H. M. Banta, S. L. Hoyt

OVER the past few years, metallic arc welding has been extended to steels of the hardenable type. As compared with other methods of fabrication, production has been facilitated, service performance frequently has been improved, and the overall results have been so favorable that engineers and production men have placed great emphasis on this application of welding. The metallurgist and welding engineer, however, have been confronted with new problems, since, for one reason, the harder steels are more apt to crack and special precautions must be taken if sound joints are to be procured. Some of the factors that are pertinent to this problem are metallurgical, and it is an object of this contribution to discuss steel manufacture and processing methods and to show how their control assists in reducing the hazard of cracking. By this time, the problem of cracking specifically underbead cracking, or "hard cracks," or parent-metal cracks-is well known, and a large amount of work has been published, both here and abroad, dealing with this subject.1 The tests described are made under restraint to facilitate cracking and, are so designed that a semiquantitative, or at least a relative, value of the cracking tendency is obtained. As one would expect, it has been reported that the incidence of cracking increases with the carbon and alloy content of the steel. More specifically, the cracking tendency is said to increase with the hardenability of the steel, though it is now recognized that the situation is not so simple. The phenomenon itself, on the other hand, has been but loosely described, and it is in only broad generalizations that this cracking has been related to steel metallurgy. Thus, while it has been abundantly demonstrated that the hardenable aircraft steel, S.A.E. 4130, must tend to crack when arc-welded, the problem still remains of why two lots of the same composition and the same hardenability can vary so greatly in cracking tendency, or what it is about the prior metallurgical history that so strongly affects the results. Further, while the efficacy of preheating and postheating in preventing cracking is recognized and utilized in production, the mechanism by which they function seems not to have been described. These points are thought to be significant in the production of steel for fabrication by metallic arc welding and, hence, of interest to mill and process metallurgists and welding engineers. Consequently, it is another object of this paper to discuss the mechanism of parent-metal cracking, including the influence of the condition of the steel and the control of cracking by manipulation of the time-temperature conditions during and after welding. A new method of rating the cracking. tendency is also discussed, and the results of the test are related to steel composition and structure or prior history. The work to be described was the outgrowth of work in 1940-41 at Battelle Memorial Institute 0n two projects, one

Jan 1, 1945

By P. R. Sperry, M. L. Holzworth, P. A. Beck

The basic work of Z. Jeffries 1,2,3 has long ago established the main features of grain growth in the presence of a dispersed second phase. Working with sintered specimens of initially fine grained tungsten, to which various amounts of thoria had been added, Jeffries found that grain growth was inhibited, that is, practically prevented, up to a certain annealing temperature. When this temperature, which increased with the thoria content, was exceeded, extremely large grains developed abruptly from the fine grained matrix. such a a temperature" was later found by Gross-mann~3,14 in certain types of steel, while ~~i~ showed15 that in other steels grain growth was gradual, without inhibition and abrupt coarsening. Later work demonstrated the connection between coarsening and aluminum additions. The viewpoint that the direct cause of inhibition and coarsening in aluminum killed steels is a fine dispersion of aluminum oxide21 appears to be held quite generally, although some doubts have been voiced even very recently, 20 In steels with Ti additions the titanium carbide phase is considered responsible for the inhibition effects found.20 The coarsening temperature increases with the titanium content. That coarsening may occur, as a result of, certain heat treatments, even in low carboil rimmed sheet steel, was shown by Samuels. He also detected coarsening in a steel ingot of similar composition, after giving it the same heat treatment as that used by him for the sheet material. In this instance the identity of the inhibiting phase has not been determined. However, Tangerding found23 that even the few small carbide particles, which occur in carbonyl iron, have a very marked inhibitive effect. As a result of an oxidizing anneal at 850°C, large grains begin to form at the surface of the carbonyl iron specimen, and gradually grow inward as the subsurface oxidation of the carbide particles progresses and eliminates the obstruction. The carbide particles Can be removed also by annealing in hydrogen, with similar results. But if the oxidizing anneal of 10 hr at 850°C in air is followed between in hydrogen for 75 hr at the Same temperature, grain growth is more hibition usually the whole specimen is transformed into a single crystal (approx. 1.5 cm X 5 cm). The reversed procedure, that is, hydrogen annealing followed by oxidizing annealing leads to smaller grains. Tangerding arrived at the logical conclusion that there must have been in his specimens some additional obstruction to grain growth hibition from a minor impurity, other than carbon. It seems likely that this obstruction was caused by an oxide formed during the oxidizing anneal and gradually reduced rtain the subsequent hydrogen-annealing. Other instances of grain growth inhibition of varying severity, associated with a dispersed second phase, have been described by several authors. Notable examples: lead with 0.06 pet Cu or Ni,24 cartridge brass with 0.12 pct Cr,25 or with up to 0.15

Jan 1, 1949

By N. Kolev, M. Poibrenski, I. Nishkov

"The following paper examines the flotation of copper porphyry ores in Assarel concentrator with focus on the reasons for low molybdenum recovery in the final copper concentrate. The copper-porphyry ore of the Assarel deposit is characterized by its variable mineral composition and high clay content, which influence the process efficiency of their concentration. Particle size distribution, chemical and mineralogical analyses have been made on various flotation products showing the distribution of the main elements of interest - copper and molybdenum. It was found that molybdenum is not concentrated into the final copper concentrate as usual in the flotation of copper-molybdenum ores but into the tailings of the copper selective flotation (cleaner tailings). An attempt to explain these results has been made with proposals to improve molybdenum flotation. Low floatability of molybdenite is caused by the ultrafine clay minerals content in both the run-of-mine ore and copper flotation streams. In order to achieve selectivity of copper minerals against pyrite a significant amount of lime is used in the plant leading to higher pH conditions, which could be another possible explanation for the molybdenite depression. The results obtained after the laboratory experiments, showed that it is possible to restore molybdenite floatability using traditional reagents such as diesel, kerosene, etc., and introducing a new polymeric reagent as froth modifier.INTRODUCTIONOne of the main technical difficulties in the area of copper-molybdenum flotation is the selective recovery of both metals, especially the fine (<40 micron) mineral particles. The problem is even more serious when Cu-Mo ore treated contains a lot of natural fine particles, predominantly clays, micas and other gangue minerals. Peng and Zhao (2011) conclude that the mineral processing industry is well aware of the difficulty in treating ores in the presence of clay minerals. Tremolada et al (2010) found that clay minerals have been named as the main problem in ore beneficiation processes due to their small size, which cause a higher viscosity in grinding. In flotation, these clay minerals have natural hydrophobicity or float by entrainment.The impact of clay mineralogy on flotation performance has been reviewed in the literature and the various deleterious roles of clay minerals have been investigated (Bulatovic, Wyslouzil & Kant 1999; Jorjani et al, 2011; Farrokhpay, Bradshaw & Dunn, 2013). It is widely accepted that the actual flotation recovery problem appears to be caused by the properties of these fine particles: small mass (and/or momentum) – high interfacial free energy (Rubio et al., 2007; Farrokhpay, 2011). In most cases, the loss in recovery is found to be insignificant, but the detrimental effect on the concentrate grade is found to be more prominent. The main reasons for the flotation inefficiency are the low probability of bubble-particle collision and adhesion, formation of dense froths and as a result - low process kinetics. A great deal of fine valuable minerals is discarded as slime in flotation due to lack of effective recovery methods."

Jan 1, 2016

By M. Misra

Jan 1, 1992

By C. Cronquist

Since April, 1962, Shell Oil Co. has operated a peripheral line-drive waterflood of the 10.750-ft lower Tuscaloosa (Cretaceous) Denkman sand in the Little Creek field. Located in southwestern Mississippi in the lower Tuscaloosa trend, the field was discovered by Shell in Jan., 1958. Among the largest in the trend, the accumulation is controlled stratigraphically and occurs in an extensive series of alluvial-point bar deposits across a nose that dips gently to the south. The crude oil is 39" API gravity and was highly undersaturated at initial conditions. Early production behavior indicated a depletion drive with slight water influx; primary recovery wm estimated to be 25 million bbl, or 24.5 percent of stock-tank oil originally in place. A.s a result of a favorable mobility ratio and remarkably uniform rock properties, the volumetric sweep eficiency in the waterflood has exceeded 90 percent. Despite the high connate water saturation (56 percent), the displacement efficiency by water injection has been quite efficient, contrary to traditional concepts. Based on observed field flood-out performance, ultimate recovery is calculated to be 46 million bbl. As anticipated, the production rate began declining during 1964 due to a continu ing reducrion in the number of producing wells as flood fronts advanced across the field. Cumulative recovery as of Jan., 1968, was 44.8 million bbl; waterflood operations ~hould be complete by 1970. Introduction Most waterfloods described in reservoir engineering literature involve pattern operations, the most common of which is the five-spot pattern. Only occasiona&apos;lly are linear drives described, even though there probably are increasing numbers being put into operation. Increasing use of linear floods may reflect the accelerating trend towards early application of fluid-injection programs when linear drives are more efficient, as was the case at Little Creek. However. the Little Creek operation is unusual in several respects. It is one of a few successful waterfloods conducted in sands with connate water saturations approaching 60 percent, and it is among the deepest successful water-floods being operated anywhere. Development The Little Creek field was discovered in Jan., 1958, in the lower Tuscaloosa trend in southwestern Mississippi (Fig. 1). The discovery well, Shell-Lemann No. 1, was drilled on a seismic closure about 2 miles southeast of lower Tuscaloosa production at Sweetwater. The well was completed in the Denkman sand at 10,752 ft, and on initial potential it flowed 588 BOPD of 39" API crude on a %+-in. choke; psi was 730 and GOR was 442:l. Development proceeded rapidly on 40-acre spacing, and by the end of 1958 the field was producing approximately 9,100 B/D from 56 wells. Almost all this production came from the northern half of the field. The southern part was discovered in Nov., 1958, at a location almost 2 miles from previously established production. Subsequent drilling established an apparent conneotion about half a mile wide between the twd areas. Development of the field was practically complete by 1961 with the drilling of 155 producers and 35 dry holes. Some early wells were completed open hole but moslt are completed with casing cemented through the productive interval and perforated for production. Geology The lower Tuscaloosa structure in this area is a north-south nose with maximum dips of about 2" on the flanks. There is no discernible faulting at the producing horizon. Productive limits of the field are shown in Fig. 2, which is contoured on a marker just above the main productive zone. A common water level at 10,415 ft subsea was logged at both the northwest and southeast ends of the field. There is a gross oil column of nearly 100 ft (Fig. 2). Since there is less than 40 ft of closure, it is apparent that bhe accumubtion is controlled stratigraphically, being limited by a sand pinch-out across the creslt of the nose. The total produc-

Jan 1, 1969

By C. Laird, C. E. Feltner

An investigation into the effect of a material&apos;s slip character on its high strain fatigue properties has been carried out using copper and Cu-7.5 wt pct A1 as representative wavy and planar slip mode materials, respectively. In copper, the constant term and exponent in the Coffin-Manson equation, ?EpNf1/2 = const, are unaffected by the temperature of testing or the history of the metal. On the other hand, in Cu-7.5 wl pcl Al, while the exponent is invariant, the constant is increased with respect to copper and increased further with decrease in temperature. The life in Cu-7.5 wt pct A1 is double that of copper at room temperature and about four times at 78°K. Fractoflaphic studies indicate that these diffevences in life are associated with the early stages of cracking. In copper, severe surface rumpling provides crack nuclei at pain boundaries independent of&apos; testing temperature or the history of the metal. In Cu-7.5 wt pct Al, the planar slip mode of deformation prevents grain boundary folding; crack nucleation is therefore mainly transgranular and the increase in life with respect to copper is associated with a prolonged linking-up stage of such transgranulay cracks. At the lower temperature the increased difficulty of cross slip inhibits multiplication of trans-granular cracks and retards the linking-up process, giving a further enhaancement in life. It is expected that such behavior will be found general in materials of wavy and planar slip mode, respectively. STAGE II crack propagation in high strain fatigue takes place by a repetitive process of crack tip blunting in the tension part of a fatigue cycle followed by resharpening of the crack in the compression part.1-3 This mechanism has come to be called the "plastic relaxation process".3 At very high strains (lives to failure less than 103 cycles), most of the life of a specimen is spent in crack propagation by the plastic relaxation process.4,5 Since the differences in the cyclic strain hardening behavior of metals and alloys are comparatively small in this strain range,6,7 the amount of blunting at a crack tip depends primarily on the applied strain amplitude. The rate of crack propagation in this life range is therefore nearly constant with respect to material and temperature of testing. In the longer life portion of the high strain fatigue range (l03 to 103 cycles to failure), a considerable fraction of the lives is spent in crack initiation and Stage I growth. Grain boundaries have frequently4,5,8 been found to be the sites for this growth which is also associated with a surface folding process.4&apos;5 Such folding behavior may also be general in any material having a grain structure and capable of plastic deformation. It has therefore been suggested475 that the universality of the Coff in-Manson law,9&apos;10 constant,* where ?Ep is the plastic strain range and Nf the number of cycles to failure, is reflected in the approximate invariance of both the crack initiation and propagation behavior. It happens that, in most investigations of the phenomena associated with failure,4,5,8 materials with a wavy slip mode have been used as vehicles for the studies, or else "normally" planar slip mode materials have been tested under conditions giving rise to wavy glide deformation, e.g., stainless steels cycled at high temperature.&apos;&apos; The purpose of the present investigation was to determine whether or not different cracking phenomena would occur when the slip mode of a material is planar and if the associated life to failure would be different than that of a wavy slip

Jan 1, 1968

By R. I. Jaffee, C. M. Craighead, G. A. Lenning

The effects of increasing hydrogen content, introducing a notch, and changing the strain rate on properties of titanium and one of its alloys were investigated over a range of testing temperatures from —196° to 200°C. Both high purity and commercial purity A-55 titanium were used as representative a materials, while a commercial Ti-8 pct Mn alloy was used for an a-P alloy. It was found possible to analyze the data, using the ductile-to-brittle transition temperature concept. Increasing hydrogen, the presence of a notch, and increasing the testing speed raised the transition temperature for the a materials. The presence of hydrogen and notches raised the transition temperature of the a-B alloy also. However, increasing the testing speed generally decreased the transition temperature of the a-p alloy. IN previous papers, the authors have shown that hydrogen is most detrimental to the notch-bend impact toughness of a-titanium&apos; and to the tensile ductility of a-B titanium alloys.&apos; These general effects have been confirmed by Kessler and co-workers," and by Kotfila and Erbin." Kessler showed that the slower strain rate in notch-rupture tests was a more sensitive measure of hydrogen embrittle-ment of a-B alloys than the tensile test. In the present paper the authors cover, with tensile and notch-bend tests, the effect of testing speed, notches, and testing temperature on hydrogen em-brittlement. High purity and commercial purity titanium, examples of u materials, and a commercial a-,f3 Ti-8 pct Mn alloy were investigated. The purpose of the investigation was to correlate the factors producing embrittlement. Materials The high purity titanium for this investigation was obtained by arc-melting crystal bar from Foote Mineral Co. into 1 Ib ingots. As-cast hardnesses for these ingots ranged from 60 to 90 Brinell. The arc-melted ingots were forged to % in. diam rods at 1470°F. These rods were swaged to & in. diam at room temperature prior to machining of the specimens for hydrogenation and testing. The commercial A-55 titanium and the Ti-8 pct Mn alloy were obtained from Rem-Cru Titanium Inc., as % in. diam rod. The A-55 titanium was hot swaged to 1/4 in. diam at 1300°F prior to machining of the specimens for hydrogenation and testing. The a-p C-11OM alloy was hydrogenated prior to hot swaging to 1/4 in. roc1 at 1300°F. Hydrogenation prior to fabrication permitted testing of the a-P alloy in the equilibrated and stabilized condition. The equilibration and stabilization treatment used was an air anneal for 1 hr at 1300°F, followed by furnace cooling to 1100°F and then air cooling. Hydrogenation was done in a Sieverts apparatus described previously.&apos; The hydrogenation treatment consisted of heating for 1/4 to 1/2 hr at 820 °C in purified hydrogen, slow cooling to 680"C, holding 1 hr for homogenization, and rapid cooling to room temperature. Intended hydrogen levels for the a-tita-nium materials were 0.05, 0.5, 1.0, 1.5, and 2.0 atomic pct—about 15 to 430 ppm. For the Ti-8 pct Mn alloy, the intended hydrogen levels were 0.05, 0.5, 1.0, 2.5, and 5.0 atomic pct—about 15 to 1100 ppm. The material for the 0.5 atomic pct level of hydrogen was vacuum annealed prior to hydrogenation to increase the accuracy of the addition. Vacuum annealing to remove hydrogen consisted of heating for 6 hr at 820°C, and resulted in a hydrogen content of about 10 to 15 ppm. When hydrogen contents were above 0.5 atomic pct—100 ppm—the hydrogen was added without prior vacuum annealing.

Jan 1, 1957

By P. E. Baker

An experimental study of heat flow in steam flooding was carried out with steam displacing water in a plane-radial fluid-flow model. Temperature distributions in the model reservoir, overburden, and substratum were measured at frequent intervals during flooding by means of a fixed array of thermocouples. Heat losses to overburden and substratum and the division of heat between the steam and hot-water zones in the reservoir were calculated from the temperature data by performing numerical integrations over the appropriate volumes. Temperature profiles and heat distribution results are presented. Some results were (1) the fraction of injected heat that is lost to overburden and substratum, when expressed as a function of time, did not depend on injection rate; (2) a significant portion of the injected heat was contained in the hot-water zone ahead of and underlying the steam front in the reservoir. The division of heat between the steam and hot-water zones did depend on injection rate; and (3) considerable gravity override was observed. The experimental results are compared with those of several published theoretical studies of heat transfer in hot-fluid injection; approximate agreement on heat losses is found with some of the most idealized theories. A valid comparison cannot be made in regard to steam us hot-water zone; basic interpretation of the hot-water zone, in terms of a vertical steam front, is clouded by gravity override in the experiments. INTRODUCTION The experimental study reported here primarily concerns two problems of heat transfer in steam flooding that have economic and engineering importance. These are (1) the fraction of injected heat lost to the overburden and substratum, and (2) the division of the remaining heat between a zone containing steam and a zone of hot water at lower-than-boiling temperature. The volume of the steamed region, which has a direct effect on oil recovery, can be calculated readily from these results. Several valuable theoretical studies have been published on this problem, but no experimental tests of the theories have previously appeared. Tests were needed because the effects of the necessary simplifications in the theories could not be determined otherwise. The basic, untested theoretical problem is the effect of "decoupling" heat and mass transfer, and the resulting inability to include the effects of heat transfer and condensation on fluid flow. The theoretical problem solved so far has been a heat transfer problem in which fluid is present as a heat carrier and flows in a stipulated manner, either linear or plane-radial. The simplifications that have sometimes been made in regard to thermal conductivity by, for example, Marx and Langenheim,l Lauwerier2 and Thomas3 apparently do not have an important effect on heat loss calculations, although the results of the different methods disagree to a certain extent, as shown by Spillette4 and by Ramey.5 A recent paper by the author6 compared the results of experiments in hot-water flooding with several theoretical results and showed a rather poor agreement. It will be seen that the experimental heat loss results for steam flooding, however, show acceptable agreement. These comparisons are discussed in a later section. EXPERIMENTAL APPARATUS AND PROCEDURE The system studied was one in which a sand pack, initially water saturated and at a uniform temperature, was flooded with steam at a constant rate and temperature. The temperature distribution in the "reservoir" and adjacent "strata" was measured with an array of thermocouples, and heat losses were determined from these measurements. The sand pack contained no oil because the only concern of the experiments was heat flow, and it was important to have the flow pattern as simple

Jan 1, 1970

By C. D. Thompson, D. C. Violetta, C. A. Czako

The mechanics of the continuous sintering process are briefly reviewed and the application of this process for the beneficiation of principal minerals used in making cement is shown. Pilot plant tests using pyrobeneficiation techniques were conducted with mixtures of limestone and shale in various ratios. Results are illustrated and show the percentage removal of carbon, carbonates, sulfur, alkalies, and volatiles as a function of the limestone-shale ratio. Tests with varying quantities of solid fwls, added to mixtures containing limestone and shale in a 4 to 1 ratio, were made to attain optimum elimination of the volatile constituents. Experimental results show that 90 pct of these constituents were removed during the calcination sintem&apos;ng process. The advantages of using the beneficiated material for making cement are discussed with regard to prepam&apos;ng, handling, and clinkering operations. A sintering process for treating raw materials or kiln waste products which contain high amounts of alkaline materials is described. Tests were performed with a high alkaline waste dust mixed in a 1 to 1 ratio with raw cement meal. The data from these tests and others using different halides for aiding alkali removal are evaluated by the alkali content of the cement clinker. Test results showed that the Dwight-Lloyd process could be applied with facility in beneficiating raw materials for cement making. Substantial quantities of deleterious volutile constituents were removed by calcining and sintering techniques. The continuous sintering process has been applied for a number of years as a means of beneficiating minerals for subsequent conversion to merchantable products. The process was developed and first used for nonferrous metallurgical applicatfons. However, it has risen to prominence for ferrous metallurgical applications where last year in the U.S. and Canada nearly 40 million tons of iron ore were treated by grate firing processes. Through the years of process development many investigators have realized the versatility of the sintering process and have directed experiments toward the use of it for applications with nonmetallic minerals. These have included calcining limestone, carbonizing coal, expanding shale, and burning cement raw meal. Today commercial sintering plants produce cement clinker from mixtures of limestone and shale, lightweight aggregate from shale or clay, and calcined products from phosphate shale. The mechanics of the downdraft sintering process have been presented by many authors. In general practice the materials to be agglomated are mixed as moistened fines with a solid fuel and this mixture is spread uniformly on the grates of the sintering machine. The charge is then momentarily directed under an ignition furnace. After ignition a continuous downdraft of air promotes combustion of the fuel and through heat transfer particles become partially fused into a coherent mass. This is cooled from continued draft induced after combustion. Many investigators have shown that distinct zones are formed during the process of sintering iron ores. At the midway point of the sintering process these are prominent and are, respectively, from the top of the bed to the bottom, comprised of a cooling zone of sintered material, a combustion zone of sinter formation, a calcination zone, a drying zone, and a zone of unreacted materials. Completion of sintering occurs near the discharge end of the machine where the firing zone approaches either the grates or grate dressing material called hearth layer. Here the peak draft temperature usually can be noted in the windboxes. The zonal pattern has been illustrated with the use of graphs which show thermal gradients as they progress through the bed of materials during sintering. Fig. 1 shows, in simplified form, thermal gradients existing when clinkering cement on a sintering machine. The thermal gradients illustrated are for periods of initiation, midway and completion of the clinkering process. The occurrence of these zones has importance for the beneficiation of cement raw materials. A research program was performed at the Dwight-

Jan 1, 1961

By S. Pearson, L. Rotherham

Measurements have been made of the variation of internal friction with temperature for spectroscopically pure silver, and for o series of solid solutions of silver with cadmium, indium, and tin, using a Ke-type torsion pendulum apparatus. Some experiments have also been made to investigate the effect of nonmetallic impurity on grain boundary relaxation in silver. The effect of the alloying elements is to increase the grain boundary viscosity, and to raise the activation energy for grain boundary relaxation from 22,000 cal per mol for pure silver to 43,000 cal per mol for the solid solutions; the same value being obtained, within the limits of experimental error, for all the alloying elements and solute concentrations investigated. Results of the experiments show exactly the same trend as those obtained previously for a similar series of copper solid solutions. They are in agreement with the general theory of grain boundary relaxation developed by Zener and Ke, but do not seem to be in agreement with either of the mechanisms so far put forward to explain grain boundary slip. EXPERIMENTS described in this paper were made as part of an investigation into the effect of alloying elements in substitutional solid solution on grain boundary viscosity, measured by the internal friction method developed by Ke.&apos; The apparatus and experimental procedure have already been described.&apos; Preliminary work was done with spectro-scopically pure silver. Experimental Work Materials Used—All the materials were supplied in the form of % mm diam wire by Messrs. Johnson Matthey and Co. Ltd. The silver and all the alloying elements used were spectroscopically pure. Composition of the alloys is given in Table I. They were all melted and chill cast under the same conditions as the copper alloys. The ingots were rolled to 1/4 in. sq rod with intermediate annealing at 600°C in an atmosphere of cracked ammonia when necessary. These rods were then drawn to wire by standard procedure with intermediate bright anneals at suitable steps in the drawing operations. Specimens from the annealed wires were sectioned longitudinally. mechanically polished, and examined under the microscope. They were then etched and reexamined. Grain sizes were reasonably uniform for all wires, and no traces of a second constituent were observed. The wires were weighed before and after testing so as to measure any loss of the alloying element due to evaporation. The loss was less than the accuracy of measurement—< 0.1 pct of the total weight— for the Ag-In and Ag-Sn alloys, but appreciable loss of weight occurred in the case of the Ag-Cd alloys. The percentage loss is given in Table 11. X-ray diffraction photographs were taken of all thc wires after test to determine whether there was any marked preferred orientation. Results of the examination are given in Table 111. Experiments on Spectroscopically Pure Silver-— Experiments were made on specimens in which the grain diameter was small relative to that of the wire, and on one specimen in which the grains were grown by the usual strain-anneal method to extend across the full diameter of the wire. These large grained specimens will be referred to as single crystal specimens. The wires were annealed for 11/2 hr at 700°C, and experiments were made at two frequencies of vibration, 1.5 and 0.4 vibrations per sec. The variation of internal friction with temperature is shown in Fig. 1. Experiments on a second small grain specimen gave approximately the same results as for the first wire. It will be seen from Fig. 1 that the peak in the curve for the small grained wire, which is presum-

Jan 1, 1957

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.&apos;-&apos; 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 D. R. Uhlmann, T. P. Seward, K. A. Jackson, J. D. Hunt

microscopic ohservations on alloys of organic trzaterials show that dendrite arms can melt off under normal conditiorzs of growth. This occurs because of the interactiorz of&apos; heat and matter fluxes during dendritic growth. The re melting occurs where the den-drite arms are joined onto the main deizclrite sterrz. Many detached crystals can be produced in this fashion. It is postulated that this remelting phenornenon occurs in metal alloys, giving rise to the equi-axed region in castings. Experiments peyformed orr castings of transparent materials support remelting as a primary source of detached crystals. Ob~ervations an equiaxed tones in metal castings are discussed in terms of this and other mechanisms. It is concluded that the partial remelting of dendrites is on important mechanisnl for producing the equiaxed regions in castings. THE freezing of a casting can be divided into three regions: a chill zone formed near the mold wall, a columnar region, and a third region, at the center of the casting, which is called the equiaxed zone. The origin of the grains that comprise the equiaxed region is the subject of this paper. We will first outline the mechanisms which have been proposed to account for the origin of the equiaxed region and discuss some recent experiments which cannot be explained by these theories. We will then present some microscopic observations of dendritic growth which have led us to conclude that the equiaxed zone in castings can arise from partial remelting of dendrites. This proposal is supported by observations on transparent castings. Finally, we will discuss several observations on metal castings which could not be explained previously, and which can be accounted for by the present proposal. Previous Mechanisms. Winegard and Chalmers3 pointed out that the molten central region of an ingot could be constitutionally supercooled, because of the solute layer at the growing interface. In this supercooled region, nucleation on particles could occur, giving rise to the equiaxed crystals. Another mechanism has also been suggested,4y5 that the nuclei originate in the chill zone then float or are carried by convection to the center of the casting, where they grow to produce equiaxed crystals. In this case, all the crystals nucleate immediately on casting, so this mechanism will be termed "big-bang" nucleation. Walker&apos;s Experiments. J. L. Walker has made observations on the solidification of nickel and Ni-Cu alloys which cannot be accounted for by either of the above mechanisms. In studying the grain size resulting from freezing of nickel at various undercoolings, walker6 found that at large undercoolings the grain size was small, due to mechanical nucleation, as discussed by walker7 and Horvay.&apos; At small undercoolings, the grain size was large, and the growth dendritic. In Ni-Cu alloys, however, for initial undercoolings starting from just below the melting point and extending for a hundred or so degrees of undercooling, fine grains were also observed. Samples were undercooled 200 Centigrade degrees without nucleating, and were then heated 130 Centigrade degrees so that the undercooling was 70 Centigrade degrees, and nucleated. The samples had a small grain size, indicating multiple nucleation. There were no heterogeneous nuclei in the liquid which could operate at the temperature of growth. Constitutional supercooling could not have been greater than the supercooling of the initial melt, since the temperature rose as the freezing continued. walker -as also shown that a Ni-Cu alloy which can be undercooled to below its solidus will produce typical columnar and equiaxed regions when cast into a cold mold. This alloy could not reach a temperature below its solidus by constitutional supercooling. It could not, therefore, have nucleated heterogeneously. Neither the big-bang mechanism nor the Winegard-Chalmers mechanism can account for the fine grain

Jan 1, 1967

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 C. S. Matthews, M. Prats, R. I. Jewett, J. D. Baker

By mathematical analysis it was found that injectivity history of a uniform five-spot pattern can be calculated by rather simple formulas. These calculated injectivities were found to agree rather well with injec-tivities measured in a potentiometric analog. With this as a basis, a simple method was finally developed which allows prediction not only of rate of injection but also of rate and kind of production at the production well, for a uniform five-spot. Using this method formulas for the eflective injectivity and production behavior of a waterflooded reservoir having a wide range in permeability can be calculated by considering that the reservoir consists of several layers, each having a uniform but different permeability. Rather good agreement is shown by a comparison of the observed production history of a field in the Illinois basin and that calculated by the methods just described. INTRODUCTION The importance of being able to predict the injection rates for a water flood is well known, for upon injection rate depends the life of the flood, the size of pumping and treating facilities, and the rate of oil recovery. One method of determining rates of injection is through a pilot flood. However, it appears that even if a pilot flood is resorted to, a theoretical method for calculating injection rate will be valuable in extending field results to times after inter- ference and to locations of other five-spots. It is also important to be able to predict the production history of a flood. Such prediction must generally be done by theoretical means, since in general a small pilot flood will not furnish much quantitative information in this regard because of the distortion after oil-bank interference. Methods for predicting the behavior of five-spot water floods have been proposed by Yuster and Cal-houn&apos; and by Hurst.? However, these methods do not consider the effect of the mobility of the different fluids in the reservoir. Other investigators3,4,5,6 have determined the effect of the water-to-oil mobility ratio on the production history by means of different experimental techniques. Most of the published work applies to homogeneous sand bodies, does not provide information for determining the injection requirements of a flood, and seldom considers the rate of build-up of the oil bank as it fills up the partially depleted reservoir.* The present work was undertaken because of the lack of predictive method which takes into account mobility ratio and the rate of buildup of oil bank resulting from the void space in a partially depleted reservoir. As water is injected into the reservoir, which contains oil, water, and gas in macroscopically homogeneous saturation, it forms several banks ahead of the injection well. In each bank or region there is assumed to be only one mobile phase—water in the water bank, oil in the oil bank and gas in the gas region. The saturation in each region and a plan view of the banks are shown in Figs. 1 and 4, respectively. FACTORS AFFECTING INJECTION AND PRODUCTION RATES Results in this report are determined as functions of the three parameters, F, M, and M, , and of a fourth parameter, r,/L** which relates the size of the well to the flood spacing. The parameter F is the displacement factor defined by Hurst in Ref. 2. It determines the rate of build-up of oil bank as it fills up the partially depleted reservoir. When gravity effects are neglected, the four parameters just given are sufficient to describe the waterflood behavior for horizontal and homogeneous reservoirs having incompressible liquids. Definition of InjectIvity The dinlensionless injectivity (or injection rate) used in this report is defined by

By J. J. Taber

The turbulent flow drag coefficients, or friction factors, have been experimentally determined for the cut-tings normally encountered in drilling operations. The gas law and average drag coefficients for characteristically flat particles (limestones and Shales) and for angular to sub-rounded particles (sandstones) were used to extend Newton&apos;s equation to a more useful form. The resulting equations are expressions for the slip velocity of a particle as a function of the particle size and shape, the bottom-hole injection Pressure, and the Bottom-hole temperature. The velocities necessary to lift actual rock cuttings as observed in the laboratory were compared with velocities predicted from the derived equations. Results indicate that there is no single correct circulating velocity and that the commonly used 3,000 ft/min linear velocity may he sufficient to lift only very mall particles. The advantage, in terms of horsepower requirements, of cycling high pressure air to obtain lower compression ratios was shown. INTRODUCTION A promising departure from standard rotary hydraulic drilling is air or gas drilling. In many instances the rate of penetration and bit life increases resulting from this method have been very substantial. Further the application of air drilling in areas of lost circulation or water sensitive pay zones has been highly successful. Purpose For the drilling contractor, the practical question remains, "How much air pressure and volume should I have?&apos; lt is necessary to have a reasonable knowledge of the air velocities and pressures required in air drilling operations for the proper design and most advantageous use of surface equipment and for adequate removal of cuttings from the borehole. Eased on experience, most operators agree that at least 3,000 ft/min linear velocity is necessary for satisfactory hole cleaning.&apos; It was the purpose of this investigation to evaluate the drag coefficients, or friction factors, for the cuttings normally encountered in drilling operations, i.e., sandstones, limestones and shales, and to utilize the values obtained to develop an expression for the minimum velocities and pressures (hence the volumes) necessary to lift these cuttings. THEORY Several investigators have done extensive work on the ability of drilling mud to lift bit cuttings. They conclude that the ability of a drilling mud to lift cuttings is dependent on the density difference between the rock being drilled and the drilling mud, the cutting size and shape, the mud flow constants, and the flow state (laminar, transitional or turbulent) of the mud. However, drilling muds are not true fluids but have a variation of viscosity with velocity in laminar flow and a constant apparent viscosity in turbulent flow. A survey of the literature revealed that no experimental work had been done to determine the velocities necessary to lift actual rock cuttings using a compressible fluid such as air or gas. In 1953, Nicolson6 stated that in air drilling fluid mechanics the particles could be assumed spherical in shape and that ; constant drag coefficient of 0.5 could be used due to the highly turbulent flow of the circulating air. By equating the turbulent resistance on the particle to the gravitational force on the particle, he obtained the expression, ^ = 2.67^y ,.......(1) where V, is slip velocity of spherical particle, ft/sec; D is particle diameter, in.; p, is particle density, 1b/ft8; and p, is fluid density, lb/ft3. However, as stated earlier, most operators seem to agree that adequate hole cleaning can be obtained by circulating a sufficient volume of air to give a linear velocity of 3,000 ft/rnin in the annular space. Basic Equation The velocity with which a solid particle freely falls through a fluid will increase until the accelerating force, gravity, is equal to the resisting forces, buoyancy due to the fluid displaced by the particle and friction due to the relative motion of the particle through the fluid. When the accelerating and resisting forces become equal,

By J. S. Kirkaldy, D. H. Weichert, G. R. Purdy

Two-phase diffusion couples have been used to simulate the growth of proeutectoid ferrite in ternary Fe-C-Mn austenites. It has been shown, theoretically and expermentally, that the results fall into two classes: a high-super saturation class in which the major role of manganese additions is to influence the boundary conditions for carbon diffusion; and a low-supersaturation class in which the partition of manganese is required. In the latter case, a drastic inhibition of the reaction rate is to be expected. As an aid to the kinetic analysis, a portion of the iron-rich comer of the Fe-C-Mn constitution diagvam has been determined. The conclusions drawn from this work have a direct bearing on the kinetics of the grain boundary ferrite precipitation reaction and thus yield insight into the manner in which alloying elements can influence the rate of austenite-decomposition reactions. THE role of alloying elements in the austenite-decomposition reactions has provoked considerable interest, since it is apparent that an understanding of this phenomenon will yield insight into the harden-ability problem. In the present investigation, attention is focused upon the proeutectoid-ferrite transformation in the system Fe-C-Mn, and upon the manner in which manganese additions affect the growth rate of ferrite. When hypoeutectoid austenite is supersaturated, the ferrite first formed usually exhibits two distinct types of morphology. At low super saturations, grain boundary layers grow into the austenite grains with an approximately planar interface, and usually possess at least one incoherent phase boundary. At high supersaturations, Widmanstiitten plates appear, which bear an orientation relationship&apos; to the parent austenite. In this case, there is a chance of partial lattice matching along the sides of the plates and an attendant semicoherent interface structure. Purdy and Kirkaldy2 have used a two-phase binary Fe-C diffusion couple to demonstrate that carbon volume diffusion controls the incoherent reaction, while Rouze and rube,&apos; using thermionic-emission microscopy, have shown that the thickening of Widmanstatten ferrite plates is appreciably slower than expected for carbon-diffusion control. This suggests that the plates are bounded by partially coherent, low-mobility interfaces. The redistribution of alloying elements during transformation has been the object of studies by Aaronson4 and owmman,5 who have shown that no alloy partition occurs during ferrite precipitation in the systems Fe-C-Cr and Fe-C-Mo, respectively. More recently, Aaronson et al.6 have shown that some manganese partition occurs at low super-saturations in the system Fe-C-Mn, but that no partition occurs at higher super saturations. Pickle-simer et al.7 observed that no manganese partition occurred during the early stages of the pearlite transformation in a manganese steel, and concluded that volume diffusion of the alloying element is not the rate-deter mining factor. The diffusion coefficients pertinent to this investigation have been measured by Wells, Batz, and Mehl8 (carbon in austenite), R. P. smith9 (carbon in ferrite), Wells and Mehl 10 (manganese in austenite), and Kirkaldy and Purdy11 (diffusion of carbon on a manganese gradient in austenite). Kurdjumov 12 has shown that the diffusivity of substitutional elements in ternary austenites up-quenched from martensite may be greatly enhanced, due, apparently, to the presence of a persistent defect structure. The recent work of Krauss 19 showing high densities of tangled dislocations in reversed Fe-Ni martensite is in agreement with this conception. THEORETICAL Diffusion in ternary or higher-order systems may be described with the aid of Onsager&apos;s extenbion of Fick&apos;s first law,14 in which the flux of each of the n components is assumed to be a linear function of all concentration gradients:

Jan 1, 1964

By C. E. Lesher

THIS paper reports a study of the reactivity of 950°F and 1650°F cokes as measured by relative rates of reduction of iron oxides at temperatures up to 2200°F. Previous work cited shows general acceptance of the theory that reduction by carbon is a gaseous reaction, and that kind and character of carbon as well as particle size have measurable effect on the velocity of reaction. As will be shown, the data obtained in this study confirm those conclusions. The work was not designed to examine iron oxide reduction equilibrium, but if reaction velocity be defined as the speed with which "a reaction tends to approach conditions of equilibrium," the data here presented may be considered as a study of reaction rates, and the relative degree of reduction to metallic iron as the measure of reactivity. Three standardized combinations of Lake Superior brown iron ore with carbon were tested by similar procedures. One combination was a mechanical mixture of carefully sized high-temperature coke (1650°F) with the ore. The second was a mechanical mixture of the ore with Disco* obtained by carbonizing the identical coal at 950 °F. The third was an agglomerate prepared by carbonizing the coal and ore at 950°F, premixed in proportions to give as nearly as possible the same relative amounts of carbon and ore as the mechanical mixtures. This agglomerate, obtained by heating the finely divided ore (through 30 mesh) with coking coal through the plastic temperature range so as to form solid aggregates, gives a product in which the oxide particles are impregnated with, and intimately bound together with low-temperature coke. The agglomerate-—ore-Disco—was most active in oxide reduction; the mechanical mixtures of Disco and ore next in order, with coke the least reactive. General Discussion: Carbon exists in many forms and it is well known that the form or nature of the carbon used in reduction of oxides is related to the critical temperature of reduction. Sugar carbon, charcoal, and lampblack are forms of carbon that will reduce oxides at lower temperatures than high-temperature coke, and coke will, in turn, give a lower critical reduction temperature than graphite. There have been many investigations of this characteristic of carbons. Johnson&apos; reported a difference of 130°F (70°C) in the critical reduction temperature of zinc oxide as between charcoal 1891 °F (1033°C) and Acheson graphite turnings 2021°F (1105°C) with zinc oxide. Bodenstein2 using charcoal and coke, found a difference of 138°F (77°C) comparing an experimental figure of 2066°F (1130°C) for coke and 1928°F (1053°C) for charcoal, in the reduction of zinc oxide. He concluded that this is very marked and observed that the "type of carbon merely raises or lowers the temperature at which rapid reaction takes place." Comparing the effectiveness of types of carbon in reduction of zinc oxide, it was found that a "brown coal coke" gave 97 pct zinc elimination at 1832°F (1000°C), as compared with 48 pct with "hard coal coke."&apos; A wide range of metallic oxides was studied by Tammann and Sworykin,4 who found that the temperature at which decomposition of oxides begins depends on the nature of the carbon used. Carbon in the form of graphite, lampblack, and sugar carbon was investigated. Sugar charcoal will reduce Fe2O3 to Fe3O4 at 842°F (450°C) as compared with 1112°F (600°C) for coke, according to Meyer." Direct reduction of iron oxides by charcoal begins at 1382°F (750°C), but "first becomes intense" at 1652°F (900°C), whereas with coke, direct reduction begins at 1742°F (950°C), and "first becomes appreciable" at 2012°F (1100°C).6 he total reduction of the sample under certain conditions when heated in a current of CO with charcoal was about 100 pct for limonite and about 77 pct for magnetite. Using coke under the same conditions, the respective percentages were 75 and 47. In a study of processes for sponge iron7 by the Bureau of Mines, the conclusion was reached that a low-temperature char from noncoking subbituminous coal is the most satisfactory solid reducing agent. In a critical study of zinc smelting from a theoretical viewpoint Maier8 concluded that the reduction is by CO, that the reaction between ZnO and CO is intrinsically more rapid than the subsequent reduction of CO2 by C, which is limited by diffusion rates, which in part effectively limits the smelting process. Maier said that the operation is improved with the activity of the reducing carbon. An active carbon, he said, is one maintaining a low CO, content in the retort. Reactivity of Carbon: One form of carbon is more potent in reducing oxides than another. A carbon that reacts faster than another at a given temperature is said to be more reactive. Reactivity is measured by several methods, using carbon dioxide, air, or steam as reactants.9 ebastian and Mayers" have developed a method for the determination of absolute reaction rates between coke and oxygen by a study of ignition points under certain conditions. These and other investigators have established the relative reactivity of types of carbon. Lignite, charcoal, bituminous coal, cokes in the ascending order

Jan 1, 1951

By Francis S. Kendorski

INTRODUCTION Caving offers the lowest cost per ton of any large-scale mining method, but its successful application demands an ore body that conforms to several rigid requirements. The deposit must be of wide areal extent, massive and not spotty in ore values, and insensitive to ore dilution. It must also be a rock mass that breaks up readily. There are only three active caving operations in the US-Climax, Henderson, and San Manuel-but caving methods have recently taken on new importance as deeper lower grade mineral occurrences and ore bodies are found. These deposits are too deep for surface min¬ing methods, and too low grade to support any type of underground mining except a bulk method such as caving. Announced discoveries or indications that may be amenable to caving include: Climax&apos;s Mt. Emmons molybdenum discovery in Colorado; Molycorp&apos;s Goat Hill molybdenum prospect in New Mexico; the Phelps Dodge molybdenum deposit in Beaver County, UT; Arizona copper occurrences such as Asarco&apos;s Sacaton, Hanna&apos;s Casa Grande, Noranda&apos;s Lakeshore, and Ken¬necott&apos;s Safford; Anaconda&apos;s suspected deep copper de¬posit in Butte, MT; Anaconda&apos;s Carr Fork, UT, deposit; and perhaps others. CAVABILITY&apos;S ROLE IN FEASIBILITY STUDIES Caving is a system of underground mining which removes support from underneath an ore body. As a result, the rock mass fractures, fails, and flows vertically downward by gravity to be collected in previously ex¬cavated funnels. Types of ore that have been mined by caving include molybdenum, copper, iron, nickel, as¬bestos, and diamonds (Julin and Tobie, 1973). It is primarily a large-scale method, with production rates of more than 45 300 t/d (50,000 stpd) having been achieved. However, the initial capital investment before return is very high, often in the hundreds of millions of dollars. The cavability of an ore body or mineral occurrence is a critical item in the feasibility study of a proposed mine, not only from the point of overall minability, but from the point of impact on other costs such as blasting, loading, hauling, crushing, and grade recovered. Aside from the often-asked question of, "Will it or will it not cave?" the real questions are, "Can we afford to make it cave, carry the rock away, and extract the mineral?" The last is not a topic of this chapter, but the first two are. The cavability of an ore deposit or mineral occur¬rence is based on many things, but clearly, if a large enough area is undermined, any rock mass will cave. The result could be a violent collapse as occurred at Urad, CO (Kendrick, 1970), or perhaps the rock mass will cave beyond the ore boundary. Another unfavor¬able result could be ore blocks that are too large for the equipment and orepasses to handle without considerable secondary blasting. Weak rock with numerous fractures may produce a very fine ore when it caves, resulting in dilution and ground control problems. DETERMINING STRUCTURAL DOMAINS It has long been recognized that the geologic nature of an ore body is important to cavability (King, 1946). Such items as weak rock material, intensity of fractur¬ing, and severity of faulting all contribute to the success of a caving operation, and information regarding these is required as a minimum for the cavability determination. In practice, the rock mass-defined as the blocks of intact rock together with the intervening fractures, joints, faults, bedding planes, and other discontinuities-that contains the ore body, as well as the surrounding and overlying rock, must be examined in a systematic and detailed fashion. Surficial geology maps must be pre¬pared, exploration holes drilled, and core logged for en¬gineering information. The fracturing of the rock mass must be studied to ascertain the three-dimensional dis¬tribution of fractures and their characteristics, and faults must be located and described. The strength and other mechanical properties of the rock material, the fracture surfaces, and the fault filling materials must be tested and reported for later use by designers and planners. With this basic information and an understanding of the geologic setting, the rock mass can be divided into one or more structural domains which tend to behave similarly in response to engineering activities (Robertson and Piteau, 1970). One must keep in mind that the determination of the structural domains goes beyond the geologic units present. Several lithologic units may be lumped together, while a single lithologic unit can be divided into multiple domains. Major faults often form their own domain, and the direction of engineering ac¬tivity-for example, cave advance to the north rather than the south-may alter rock mass behavior, resulting in different domains. As an example of the detailed rock fracture map¬ping required for such studies, the structural domain determinations at the Climax mine (Kendorski, 1973) are shown in Fig. 1. The circles are Schmidt equal area projections of the three-dimensional attitudes of frac¬tures (Ramsay, 1967) mapped in detail at various rock exposures. The attitude of fractures is important to the cavability determination since it dictates the directional behavior of the rock mass as it fails, and determines the effectiveness of arching, keying, and rock block inter¬locking. Low-angle fractures must be present to allow movement of the rock in the vertical direction during undermining (Mahtab and Dixon, 1976); if low-angle planes of weakness are absent, the rock mass may arch with a keystone effect, rather than moving vertically downward.

Jan 1, 1982

By Bernard F. Anderson

INTRODUCTION In this chapter, the term "down-the-hole drill" (DTH drill) is used as a generic name that encompasses the various trade names and other references such as "downhole drill," "in-the-hole drill," etc. This chapter is limited to a description of DTH drills used in stoping large underground ore bodies. DTH drills differ from conventional drills by virtue of the placement of the drill in the drill string. The DTH drill follows immediately behind the bit into the hole, rather than remaining on the feed as with ordinary drifters. Thus, no energy is dissipated through the steel or couplings, and the penetration rate is nearly constant, regardless of the depth of the hole. Since the drill must operate on compressed air and tolerates only small amounts of water, cuttings are flushed either by air with water-mist injection or by standard mine air with a dust collector at the collar. HISTORICAL DEVELOPMENT Mine managers have long known the economies enjoyed by quarry and open-pit operators in producing large quantities of ore. The savings are due primarily to the availability of massive equipment, capable of drilling large blastholes to reduce the amount of drilling, increase the fragmentation, reduce secondary blasting, and im¬prove the flow of the product. In an attempt to reduce underground mining costs, various methods are used for long-hole drilling, includ¬ing standard pneumatic percussion drifters and diamond drills. These systems have their shortcomings; percus¬sion drills are limited to small hole sizes and they ex¬perience excessive deviation and significant loss of energy with increased depth. The diamond drills provide deeper and straighter holes, but only at high cost. Both systems suffer from high noise levels, low penetration rates, and poor explosives distribution, among other problems. When the mining companies approached the drill manufacturers for a compact and portable large-hole jumbo for underground use, they specified not more than 1 % deviation on 60 m (200 ft) of vertical hole and a penetration rate of 15 m/h (50 fph). On Dec. 23, 1960, a test unit was placed in service in Montana and met the performance criteria. Though lacking the so¬phisticated features available today, the economies of surface blasting were brought underground. Unfortunately, the first system did not gain immedi¬ate acceptance in the industry. Among the factors con¬tributing to its demise were resistance to change, the need to alter development methods for the ore bodies, and a lack of flexibility in moving the rig from setup to setup and from level to level. In 1972, the mining industry again challenged the drill manufacturers to provide a workable jumbo that would combine compactness, ease of maintenance, relia¬bility, and efficiency, all on a self-propelled chassis. The manufacturers responded by providing improved jumbos, which have been accepted with enthusiasm throughout the mining industry. Today&apos;s DTH jumbos are capable of drilling from 100 to 200 mm (4 to 8 in.) diam holes that can be reamed to even larger diameters. The holes can be drilled to depths of 150 m (500 ft), depending upon ground conditions and the capability of the jumbo to retrieve the steel and drill. Fig. 1 illustrates a typical DTH jumbo. APPLICATIONS The uses to which DTH drill jumbos have been put are quite numerous, with new uses being found regularly. For convenience, these uses may be classified as primary blastholes and nonblasting holes. Primary Blastholes The original purpose for the development of the DTH jumbo was for drilling primary blastholes that could be mined by open-stope methods. Prior to the advent of the DTH jumbo, extensive development was required before production drilling could begin. Sub¬levels were required to allow access for column-and-arm stopers or ring/fan jumbos, to the extent necessary based on the effective penetration of the chosen machine. With the DTH jumbo, the mine engineer is able to reduce preproduction time and development costs. How¬ever, the most significant saving results from an im¬proved cost per ton of broken ore in the production phase. To utilize a DTH system, only a top heading and drawpoints are necessary. The top heading can be the width of the ore body with a 3.7-m (12-ft) back. A drop-raise pattern is drilled and shot to begin the stoping operation, providing a free face for subsequent blasting. A typical layout is illustrated in Fig. 2. The advantages of this system include: 1) Drilling and blasting are independent operations, and blasting can be performed at a rate congruous with the mine&apos;s ton-per-day capacity. 2) The development layout is simplified. 3) Good explosive distribution is achieved, provid¬ing more uniform fragmentation. 4) Environmental conditions for operators are im¬proved, including improved safety with all work directed downward (not overhead), lower noise levels, little fog, and a reduced dust count. 5) Improved production per manshift. 6) Simplified and easier operator work cycles. 7) Reduced cost per ton of product. 8) Fewer holes lost due to ground shifts. Nonblasting Holes With the introduction of the compact DTH jumbos, other practical uses became apparent, including the drilling of: 1) Holes for sand fill, from level to level and from level to stope. 2) Drain and dewatering holes. 3) Power and communications cable holes.

Jan 1, 1982