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By C. R. Ramachandra, C. C. Patel

Flotation of chalcopyrite from a low grade ore was studied by using different xanthates and dixanthogens as collectors and by conditioning the flotation pulp with oxidizing gaseous systems. The improvement in the recovery of chalcopyrite with the oxidizing systems was mainly due to the auto-activation of chalcopyrite, leading to the enhanced adsorption of xanthates and dixanthogens even when the latter were formed by oxidation of the xanthates. With air or oxygen as oxidizing system, the amount of ethyl xanthate required at pH 8.5 was much less than normally employed in the industry, the recovery being 97-99% of chalcopyrite with a concentrate of 24% copper. With isopropyl and isoamyl xanthates, the recovery was improved but the grade was lowered. By conditioning the pulp with air-CO2 or oxygen-CO2 mixture, the recovery was lowered with ethyl xanthate only, mainly due to the depressant action of bicarbonate formed. Flotation with dixanthogens gave cleaner concentrates of chalcopyrite. because iron pyrite did not float with these collectors. Of the dixanthogens tried, diisopropyl dixanthogen gave the highest recovery, 99% chalcopyrite with a concentrate containing 28.5 to 30% copper, when the pulp at pH 8.5 was condi-tioned with air. Isopropyl dixanthogen was found to be the best collector of all the xanthates and dixanthogens studied. It was found that the adsorbed ethyl, isopropyl and isoamyl xanthates at chalcopyrite surface, activated by cupric sulfate, and at iron pyrite surface were oxidized to the corresponding dixanthogens to different degrees in presence of air, oxygen, air-carbon dioxide mixture and oxygen-carbon dioxide mixture.' The dixanthogen formed at chalcopyrite surface was, however, firmly adsorbed while that at iron pyrite surface was easily desorbed. Since dixanthogens formed have greater hydrophobicity than the corresponding xanthates, as revealed by contact angles,1,2 the former were expected to improve the flotation results. It was, therefore, thought desirable to find out the effect of air, oxygen, air-carbon dioxide mixture and oxygen-carbon dioxide mixture on the selective flotation of chalcopyrite when individual xanthates were used as collectors. The effect of direct use of dixanthogens as collectors on the selective flotation of chalcopyrite was also investigated. For these studies, low grade chalcopyrite ore, received from the Indian Copper Corp., Ltd. (I.C.C.), Ghatsila, Bihar, was employed. At the milling plant of the I.C.C., chalcopyrite is floated at a pH of 8.5, employing ethyl xanthate as collector. EXPERIMENTAL Materials Employed: XANTHATES - Ethyl, isopropyl and isoamyl xanthates of potassium were prepared in a pure state by the methods described earlier.4 A 0.03% aqueous solution of a xanthate was freshly prepared in air-free distilled water for flotation studies. DIXANTHOGENS - Diethyl, diisopropyl and diisoamyl dixanthogens were prepared by the oxidation of xanthates by iodine and purified by extraction with petroleum ether. A 0.03% alcoholic solution of dixanthogen was used as a collector. FROTHER - A 0.2% solution of pine oil in ethyl alcohol was employed. SODIUM CARBONATE - A 0.4 N solution was used. ORE - The composition of the -100 +200-mesh size (Tyler standard) powdered ore, as determined by standard methods of analysis, was: copper 1.57%, iron 10.29%, sulfur 3.15% and acid insolubles 66.10%. Iron present as chalcopyrite (CuFeS2) was 1.38%, while that present as iron pyrite (FeS,) was 1.37%. Both chalcopyrite and iron pyrite were found to be liberated from the gangue minerals and from each other, as observed under a microscope. Flotation of Chalcopyrite with Xanthates in Presence of Different Gaseous Systems: Fifteen grams of -100 + 200-mesh size ore were shaken well with 100 ml of distilled water. The pH of the pulp was adjusted to 8.5 by the addition of sodium carbonate solution. Then 0.1 ml of xanthate solution (0.005 lb xanthate per ton of the ore) was mixed with the pulp, and the latter together with washings transferred to a

Jan 1, 1963

By L. D. Jaffe, D. C. Buffum

EARLIER the authors and coworkers had pre sented data on isothermal temper embrittlement of an SAE 3140 steel?' In that work, however, attention was concentrated on embrittlement at 575°C and below. Preliminary measurements showed that embrittlement at 675 °C was much mote rapid than would be expected from the data at the lower temperatures: and indicated the need for further work above 575°C. The possibility of an upper nose in the time-temperature embrittlement diagram above 575°C was pointed out. Procedure The same heat of SAE 3140 steel was used as in the previous work, and experimental methods were the same. All blanks were austenitized 1 hr at 900°C and water, quenched. Groups were then tempered 1 hr at 675"C, water quenched, and given embrittling treatments for times of 5 min to 240 hr at 675" to 600 °C, as shown in Table I. Other groups, after the temper and quench, were embrittled 48 to 1440 hr at 550°C. Another series was tempered 240 hr at 675"C, quenched, and held for various times at 650" to 375°C (Table 11). For comparison, a series was tempered for these same times at 700°, 675", and 650°C (Table 111). The Ae, temperature for this heat of steel had been determined as approximately 690°C;1*a the treatments at 700°C were introduced to determine the effects of tempering in a range where a little austenite would fonn. A large number of other treatments, involving holding at two temperatures in succession after the temper, and occasionally slow cooling or heating, were also studied. These treatments are listed in the tables. Austenitizing was done in air; all subsequent treatments were in salt. Blanks were always water quenched after the last treatment. The heat-treated blanks were machined to standard V-notch Charpy specimens. A transition curve for each group was obtained by breaking the specimens at various temperatures on a 217 ft-lb impact machine having a striking velocity of 16.8 fps, and plotting the observed percentage of fibrous fracture vs the test temperature. From this plot, the transition temperature was taken as the lowest temperature at which the fracture appeared 100 pct fibrous; its accuracy is estimated as +5"C. Generally, hardness was measured on every bar. Since a decrease in hardness itself seems to decrease the transition temperature somewhat, a correctiona of —0.28"C per Bhn was applied to permit comparison of the transition temperatures on a uniform basis of 24 Rc. The transition temperature and hardness results are given in Tables I through VI. Where overlap of earlier work1.' occurred, data were combined.

Jan 1, 1958

The Library is a cooperative activity of the American Institute of Mining and Metallurgical Engineers, the American Society of Civil Engineers, the American Institute of Electrical Engineers and the American Society of Mechanical Engineers. It is administered as a public reference library of engineering and the allied sciences. It contains 133,000 volumes and pamphlets, 6700 maps and receives currently most of the important periodicals in its field. In order to place the resources of the Library at the disposal of those unable to visit it in person, the Library is prepared to furnish lists of references to engineering subjects, copies or translations of articles, and similar assistance. Charges sufficient to cover the cost of this work are made.

Jan 1, 1933

Denver, the Mile High, Queen City of the Plains, your host for AIME's 99th Annual Meeting has many exciting and picturesque places for you to visit during your stay. A host of restaurants cater to every taste; supper clubs and theaters provide nighttime entertainment; there are sports facilities and a plethora of shops for the ladies. When it comes to dining out, there are many excellent restaurants within walking distance of the Denver Hilton --among them, three in Brooks Towers: the Gitana, the Boston Half Shell serving the finest seafood, and Marvelous Mary's. In the Larimer Square area there in the Bratskellar with a true Germanic atmosphere and Sperte's Laffite offering a typical New Orleans setting.

Jan 1, 1970

By P. Chiotti, J. T. Mason

Ther?nal, X-ray, metallographic, and vapor pressure data were obtained to establish the phase diagram and standard free energy, enthalpy, and entropy of formation for the compounds in the Sw-Zn system. Four compounds, SmZn, SmZn2 , SmZn4.s, and SmZn8.5, melt congruently at 960°, 94Z°, 908°, and 940°C, respectively. The cornpounds SlnZns, Sm3Znll, and SnzZn7.3 undergo peritectic decomposition at 855", 870°, and 890C, respectively. Another compound of uncertain stoichiometry, SmZn11, undergoes peritectic decomposition at 760°C. Four entectics were observed with the following compositions in weight percent zinc and eutectic tenzperatures in degrees Centigrade: 12 pct, 680°C; 36 pct, 890°C; 58 pct, 850°C; and 72 pct, 900°C. An allotropic transformation and a composition range were observed for the SmZnz compound. The transfor)nation varies from 905" to 865°C as the zinc content increases from 16.0 to 48.5 wt pct, respectively. The free energy of formation of the compounds at 50PC varies between -15.9 kcal per mole for SmZn to -51.1 kcal per mole for SmZn,.,. Corresponding enthalpies vary between -19.2 to -78.3 kcal per mole. The ther-modynamic properties for the liquid alloys are described by the relations: A search of the literature revealed very little information on the Sm-Zn system. Chao et al.&apos; as well as Iandelli and palenzonai have reported the structure of SmZn to be cubic B2 type and Kuz&apos;ma et al3. have reported the structure of -sm2zn17 to be of the Th2Ni17 type. The purpose of this work was to establish the phase diagram of this system, to determine the zinc vapor pressure over the solid two-phase regions of the SYstem, and to calculate the thermodynamic properties of the compounds. MATERIALS AND EXPERIMENTAL PROCEDURES The metals used in this investigation were Bunker Hill slab zinc 99.99 wt pct pure and Ames Laboratory samarium. Analysis of the samarium by chemical, spectrographic, and vacuum-fusion methods gave the following average impurities in ppm: Nd, <200; Eu, <100; Gd, <100; Y, <50;Ca, 225; Ta, 400; Mg, 10; Cu, ~50; 0, 175; H, 20; and N, 15. The elements Fe, Si, Cr, Ni, Al, and W were not detected. The samarium was received as sponge metal and was kept under argon except when being cut with shears and when being weighed. Tantalum was found to be a suitable container for alloys with zinc contents up to the Sm2Znl, stoichio-metry. At higher zinc contents the grain boundaries of the tantalum containers were penetrated by the alloy and the containers failed during prolonged annealing. About 25 g of massive zinc and samarium sponge were sealed in tantalum crucibles equipped with thermocouple wells. These crucibles were in turn sealed in stainless-steel jackets. All closures were made by arc welding under an argon atmosphere. The samples were equilibrated in an oscillating furnace and in some cases were given various heat treatments in a soaking furnace. After appropriate heat treatment the steel jackets were removed and the alloy subjected to differential thermal analysis. The apparatus was calibrated against pure zinc and pure copper and found to reproduce the accepted melting points within 1°C. Alloys were subsequently subjected to metallographic examination and those of appropriate compositions were used for X-ray diffraction analysis and for zinc vapor pressure determinations. The vapor pressures were determined by the dewpoint method. Both the differential analysis and dewpoint measuring apparatuses have been described in earlier papers.4, 5 All alloy samples were etched with Nital (0.5 to 3 pct nitric acid in alcohol) except the samarium-rich alloys. These more reactive alloys were electro-polished in a 1 to 6 pct HClO4 in methanol solution at -700c at a potential of 50 v. EXPERIMENTAL RESULTS Phase Diagram. The results of thermal analysis are indicated by the points on the phase diagram, Fig. 1. Eight compounds and four eutectics were observed. The composition of the compounds and their melting or peritectic temperatures are given on the phase diagram. The four eutectic compositions in wt pct zinc and eutectic temperatures in % are: 12 pct,- 680°C; 36 pct, 890°C; 58 pct, 850°C; and 72 pct, 900°C. The stoichiometry of the most zinc-rich compound is still uncertain, but is very likely either SmZnll or SmZnlz. However, to simplify the presentation which follows it will be referred to as SmZnll. As shown on the phase diagram the phase regions for some of the samarium-rich alloys have not been unambiguously established. A sample of pure samarium was observed to transform at 924°C and to melt at 1074"C, in good agreement with corresponding val-

Jan 1, 1968

By S. W. Nabbs, W. D. Bachman, A. W. Last

A study has been made of wet, autogenous grinding of disseminated copper ores, including testing of a large number of samples from. Kennecott Copper Corp.&apos;s Chino mine. The grindability of the various samples was found to vary over an extremely wide range, even among samples taken from diflerent locations within the same ore body. These variations in grindability can be qualitatively correlated with the composition of the rock and the fracturing and alteration that had occurred. Because of the large-scale nonhomogeneity of disseminated copper ore bodies, extensive and careful testing of autogenous grinding is needed. An accurate correlation of the grinding data from various samples with the geology of the ore body and with mining plans will indicate if variations in grinding mill throughput can be avoided in practice by mixing or blending of mill feed and if, therefore, autogenous grinding can be applied successfully. In the past 15 years, comminution by primary autogenous grinding has been investigated and applied economically in the minerals industry, particularly in iron-ore processing. Successful application of primary grinding is described by A. A. Dor,l.&apos; P. B. Dettmer,334 R. R. Turner, and W. F. McDermott." other authors describing the advantages of primary grinding include H. Hardinge,&apos; R. T. Hukki,&apos; D. S. Coyle" and F. C. Bond,"," to mention only a few. For the past several years, Kennecott Copper Corp.&apos;s Metal Mining Div. Research Center has been investigating primary grinding of disseminated copper ores. Both fully autogenous and semi-autogenous grinding has been explored. Ores from three of Kennecott&apos;s western operating properties and from several development properties have been tested. Host rocks include diabase, schist, silicated and argillaceous limestones, shale, granite porphyry, quartzite, and granodiorite. The object of this paper is to supplement the general knowledge in the field of primary grinding by presenting information gained from the studies with the various disseminated copper ores. Data are presented which show that carefully considered test programs are required to determine the practicality of applying autogenous grinding to a specific mining-beneficiation operation. Particularly this is so if more than one rock type occurs within an ore body or if there are wide variations in the degree of rock alteration within a deposit. Extensive sampling and subsequent testing of the ores within a mine are required if serious pitfalls are to be avoided and good engineering data acquired. We hope that this presentation will stimulate studies to increase the understanding of the fracturing process applicable to large pieces of rock, a factor which is important in autogenous grinding. Test Facilities and Procedures Both pilot-scale and full-scale test facilities were used to investigate primary grinding characteristics of the disseminated copper ores. A 6-ft Hardinge Cascade mill, a 24-ft Hardinge Cascade mill, and a 5-ft Aerofall mill were used in the studies. The Cascade mills were operated wet, while the Aerofall mill was operated both wet and dry. The Cascade mills were operated both for fully autogenous and for semi-autogenous grinding; the Aerofall mill was operated only semi-autogenously. Fully autogenous grinding, as the term is used in this presentation, is defined as the direct grinding of coarsely crushed rock by the action of the rock mass. Semi-autogenous grinding is defined as the grinding of coarsely crushed rock with the rock mass supplemented by a steel ball charge. Most of the test work was done using the 6-ft Cascade mill, operated under fully autogenous grinding conditions. Only limited testing was conducted with the 5-ft Aerofall mill and only one ore was tested using the 24-ft Cascade mill. This presentation is concerned, therefore, primarily with fully autogenous, pilot-scale grinding in the Cascade mill. Fig. 1 presents a diagram of the test facilities used with the 6-ft Cascade mill. Principal test equipment consists of a 6-ft Cascade mill, an 18-in. spiral classifier, an elevator, a 2 x 9-ft belt feeder, a conventional integrating kilowatt-hour meter, and a recording wattmeter. In testing, feed charges, adjusted as to weight, were deposited as evenly as possible on the belt feeder every 15 min. The weight of each charge was adjusted to maintain a mill charge of 27 to 30% of mill volume. The classifier was adjusted to obtain the desired size-consist of the final product, generally at 20% + 100 mesh. After equilibrium was established in the circuit, the feed, mill discharge, classifier sand, and classifier overflow were sampled and their size distributions determined. During the sampling periods, the power drawn by the mill was measured to determine the energy consumed in the reduction of the size consist of the ore solids. Fully Autogenous Pilot Grinding Tests on Various Kennecott Ores Typical test results obtained in the 6-ft Cascade mill with samples from Kennecott&apos;s Ray Mines Div., Chino Mines Div., Utah Copper Div., and a development property designated as Mine A are presented in Table 1. As shown, the ores that were ground most easily in the pilot-scale mill were composite mine samples obtained

Jan 1, 1971

By A. P. Young, F. C. Holden

IN many of the early investigations of the aging of titanium alloys, it was observed that the retained beta phase could be aged to a high hardness without an apparent change in microstructure; moreover, in some alloys it was found that even a very rapid quench produced beta having an anomalously high hardness. What appeared to be hard beta phase was designated as "beta prime," to distinguish it from the softer normal retained beta phase. Frost and co-workers&apos;.&apos; demonstrated that the increased hardness is caused by the formation of a metastable phase, omega, which can be identified from its X-ray diffraction pattern. This reaction also has been studied by Brotzen, et a1.,3 and the structure of the omega phase has been published by Silcock, et al.," and by Austin and Doig.3 The X-ray diffraction analysis by Austin and Doig indicated that omega is low in alloy content, approaching the composition of the stable alpha. The aging process proceeds by the path Thus, overaging results in the formation of alpha and the ultimate disappearance of omega. This is accompanied by a decrease in hardness and a recovery of ductility. The present work represents a study of the aging process in a high-purity Ti-6.4 Mn alloy, in which optical and electron microscopy, X-ray diffraction, and hardness data were correlated. The specimens, 1/8-in. sections cut from 1/4-in.-diam rod, were heated in argon to a temperature just into the beta field (810°C), held for 1 hr, and water-quenched. The aging samples were heated in air for 1 hr at 300 to 500°C. The specimens were ground and mechanically polished, finishing with alternate etching and polishing on a slow-speed wheel to remove disturbed metal from the surface. The surface finally was swab-etched, rinsed with tap water, and air-dried. The etchant used was an aqueous solution containing 1 1/2 pct HF and 3 1/2 pct HNO3. This procedure is identical with that used for preparing metallographic specimens, except that the optimum degree of final etching required for electron microscopy may vary from that used for light microscopy. Metal replicas, backed with silica, were prepared from plastic strip-pings for examination with the electron microscope. Complete details of the procedures used for preparing these replicas are available elsewhere." Fig. 1 shows a comparison between the structures revealed by the electron microscope and those obtained by conventional light microscopy. Hardness and X-ray diffraction data are listed in Table I. No evidence of a second phase is apparent in the optical micrographs for the as-quenched sample, or for the specimens aged at 300 or 400°C; fine alpha particles can be seen in the specimen aged at 500°C. In the electron micrographs, visible equiaxed pocks appear in the as-quenched specimen, and, with increasing density, in the samples aged at 300 and 400°C. After aging at 500°C, the structure becomes markedly coarser and develops an oriented appearance of a second phase, coincident with the appearance of visible alpha particles in the light micrographs. These observations are confirmed by X-ray diffraction and hardness data. In these specimens, aged for 1 hr, the intensity of the omega pattern increases to a maximum at the 400°C aging temperature, corresponding to maximum hardness and the maximum density of pocks in the electron micrographs. At this point the beta phase shows a definite enrichment in Mn. At 500°C, little or no omega re-

Jan 1, 1959

By S. T. Wlodek

The surface and subscale oxidation reactions were followed by means of continuous weight-gain and metallographic techniques over the range 1600" to 2200°F (871° to 1204 °C) for up to 400 hr. Full identification of all scale and subscale reaction products was obtained by electron and X-ray diffraction. At or below 1800°F (982°C) a linear rate of reaction (QL = 46.0 kcal per mole) governed the oxidation process, extending for up to 100 hr at 1600°F (871 "C). During linear oxidation the surface scale consisted of an amorphous SiO2 film overgrown with Cr 2O 3 and NiCr204. This initial linear process was followed, and above 1800°F completely replaced, by two successive parabolic rate laws (Qp = 60 and 57 kcal per mole). This parabolic reaction involved the formation of a complex scale consisting of Cr2 O3 and smaller amounts of NiCr2O4. Parabolic oxidation appeared to coincide with the disruplion of the silica film present during linear oxidation and was followed by subscale (internal) oxidation of crystobalite and NiCr2O4. The balance between the subscale and surface oxidation reactions controls the oxidation of this commercial alloy. The amorphous silica film appears to result in the linear rate and diffusion through Cr2O3 is the more likely rate-limiting step during parabolic oxidation. THE oxidation of a multicomponent composition is a complex phenomenon not presently amenable to a rigorous classical interpretation. Nevertheless, even a qualitative understanding of the scaling and subscale reactions that occur in a commercial composition can illuminate the reactions that limit its high-temperature stability in an oxidizing environment. This study of the oxidation of Hastelloy Alloy X presents the first of a series of studies with the above approach in mind. Hastelloy X exhibits one of the best combinations of strength and oxidation resistance available in a wrought, solution-strengthened, nickel-base alloy. Although during long time exposure some precipitation of M6C and M23C8 carbides as well as a complex Laves phase occurs, the amounts are probably small enough to have no appreciable effect on the chemistry of the matrix. Radavich has identified the oxidation products on Hastelloy X oxidized for 5 min to 10 hr at 1115°F as NiO and the NiCr2O4 spinel. Oxidation for 5 to 15 min at 1500°F produced a scale of spinel, NiO, and a rhombohedra1 phase, probably Cr2Os. Sannier et 2. have reported continuous weight-gain data for Hastelloy X at 1650" and 2010°F and internal-oxidation measurements after 150 hr at 2010°F. In addition, much of the data on binary Ni-Cr alloys recently reviewed by Kubaschewski and okins&apos; and Ignatov and Shamgunova4 as well as studies of binary Ni-Mo alloys5 are also pertinent to the oxidation of this composition. EXPERIMENTAL Continuous weight-gain measurements and metallographic measurements of subscale reactions were the main experimental techniques used in this study. X-ray and electron diffraction backed up by a limited amount of electron-microprobe analysis served to characterize the nature of the scale- and subscale-reaction products. Two heats of commercial sheet of the composition given in Table I and identified as A and B were used in the bulk of this study. Internal-oxidation measurements were made on a third heat of material in the form of a 0.5-in.-diam bar. In order to assure homogeneity, all heats were reannealed 4 hr at 2175°F prior to sample preparation. weight-Gain Measurement. All specimens (1.5 by 0.4 by 0.03 in.) were abraded through 600 paper, electropolished, and lightly etched in an alcohol-10 pct HCl solution. An electrolyte of 150 cu cm H,O, 500 cu cm HsPO4 (85 pct conc), and 3 g CrO3 at a current density of 0.9 amp per sq cm or a solution of 10 pct HaW4 in alcohol used at 4 v and 0.3 amp per sq cm was used for electropolishing. The resultant surface exhibited a finish of 3 ± 1 p rms. Continuous weight-gain tests were made at 1600°, 1700°, 1800°, 1900°, 2000", and 2200°F on auer&apos; type balances capable of recording a total weight change of 110 mg with an accuracy of k0.1 mg. All tests were made in air dried to a dew point of -70°F and metered into the 2-in.-diam reaction

Jan 1, 1964

By John J. Collins

[CONTENTS PAGE Purpose and scopeI Core samplingI Sludge sampling.....7 Combining core and sludge assays13 Deviation and surveying of drill holes17 Conclusions24 Acknowledgments25 Bibliography25] PURPOSE AND SCOPE The purpose of this paper is to review a few of the problems involved in the interpretation of diamond-drill core and sludge samples, with the hope that it will stimulate discussion and bring forth some of the information on the subject that has been accumulated by various organizations. The release of these data would benefit the mining industry as a whole. Particularly valuable would be the details of drilling programs that have been tested by actual mining operations, as most of the available published material is unsupported by production records. The scope of this paper is limited to a few of the commonest problems of interpreting diamond-drill core and sludge samples and the deviation of holes. It does not include the description of routine operations performed by the drillers, nor the estimation of ore reserves. The methods of calculating reserves have been described 25,26,31,32.50,65 but more information is needed on the results shown by actual mining operations. CORE SAMPLING The accuracy of diamond-drill-hole sampling depends in large measure on the percentage of core recovery. As a general rule, the longer the run, the poorer is the recovery. Although the longer runs allow cheaper drilling costs, the determining factor should be the percentage of recovery desired. The problem is one of balancing accuracy against costs. Holes of larger diameter may show improved recovery, except as Joralemon* points out that there are some conditions in which small holes give better core recovery than large holes. This is particularly true, he says, in rocks like shattered quartzite, where larger core tends to break in angular fragments and grind up, while the diameter of smaller core is less than the distance between joint planes and grinding is not so serious. In a large measure the length of the sample run should depend on the type of deposit. Favorable drilling may allow uniform sample runs of 5 ft., 10 ft., or perhaps 20 ft. In veins or beds of relatively narrow widths and changing characteristics, however, the ore zone may be overrun and ground in a core barrel full of waste rock. Generally it is desirable to break the sample at a change in the rock or ore. Such a change, often indicated by a color change in the sludge water, should be a signal for the driller to make a "dry" block and to pull out that run regardless of the length. Such a procedure permits the segregation of waste from ore samples so

Jan 1, 1946

By John J. Collins

Page Purpose and scope..................... 521 Core sampling......................... 521 Sludge sampling....................... 527 Combining core and sludge assays ...... 533 Deviation and surveying of drill holes. . . . 537 Conclusions........................... 544 Acknowledgments...................... 545 Bibliography.......................... 545 Purpose and Scope The purpose of this paper is to review a few of the problems involved in the interpretation of diamond-drill core and sludge samples, with the hope that it will stimulate discussion and bring forth some of the information on the subject that has been accumulated by various organizations. The release of these data would benefit the mining industry as a whole. Particularly valuable would be the details of drilling programs that have been tested by actual mining operations, as most of the available published material is unsupported by production records. The scope of this paper is limited to a few of the commonest problems of interpreting diamond-drill core and sludge samples and the deviation of holes. It does not include the description of routine operations performed by the drillers, nor the estimation of ore reserves. The methods of calculating reserves have been described,25,26,31,32,50,65 but more informa- tion is needed on the results shown by actual mining operations. Core Sampling The accuracy of diamond-drill-hole Sampling depends in large measure on the percentage of core recovery. As a general rule, the longer the run, the poorer is the recovery. Although the longer runs allow cheaper drilling costs, the determining factor should be the percentage of recovery desired. The problem is one of balancing accuracy against costs. Holes of larger diameter may show improved recovery, except as Joralemon* points out that there are some conditions in which small holes give better core recovery than large holes. This is particularly true, he says; in rocks like shattered quartzite, where larger core tends to break in angular fragments and grind up, while the diameter of smaller core is less than the distance between joint planes and grinding is not so serious. In a large measure the length of the sample run should depend on the type of deposit. Favorable drilling may allow uniform sample runs of 5 ft., 10 ft., or perhaps 20 ft. In veins or beds of relatively narrow widths and changing characteristics, however, the ore zone may be overrun and ground in a core barrel full of waste rock. Generally it is desirable to break the sample at a change in the rock or ore. Such a change; often indicated by a color change in the sludge water, should be a a for the driller to make a "dry" block and to pull out that run regardless of the length. Such a procedure permits the segregation of waste from ore samples so

Jan 1, 1946

By J. A. Fellows, Earnshaw Cook, H. S. Avery

THE increased use of heat-resistant alloys (26 per cent Cr, 12 per cent Ni, 16 per cent Cr, 35 per cent Ni, 12 per cent Cr, 60 per cent Ni, etc ) in recent years has been accompanied by continued demands by the consumer for improvement in properties The manufacturers of these materials, of necessity, have resorted to more critical studies of the basic causes of variations in fundamental properties To keep pace with the need for augmented information, previous techniques of investigation have been refined and new test methods and instruments created It is the purpose of this paper to describe the apparatus, experimental technique and precision of measurement that were employed in the exploration of properties of heat-resistant alloy presented in a companion paper† The detailed precautions in experimental procedure essential for attainment of accurate and comparable data at elevated temperatures cannot be too highly emphasized THE CREEP TEST Since the fundamental concepts of creep testing have been recognized for the past 20 years, present interest is restricted to refinements in equipment and experimental procedure The creep-test laboratory constructed for the American Brake Shoe and Foundry Co is designed to permit 1000-hr tests in the temperature range of 1200° to 2200°F with a choice of applied stress extending from 50 to 50,000 lb per sq in in increments of 50 lb per sq in To simplify the calculation of weight combinations, the loading beams (Fig I) are provided with counterweights to eliminate the tare The precision of the knife-edges (10 1 ratio) is such that a 10-gram weight placed on the upper specimen grip will be exactly balanced by a 1-gram weight attached to the loading shackle Check of the loading weights proved the maximum error to be 0 06 per cent Axial alignment is provided by crossed knife-edges in the grip seats The furnaces are similar in design to those of Crane and Company, Battelle Memorial Institute, and the U S Steel Corporation An 80 per cent Pt, 20 per cent Rh No 19 B and S gauge wire is used as the heating element The details of construction are shown in the radiograph reproduced in Fig 2 The windings are held in place on an alundum tube with alundum cement, the distribution of turns is adjusted to give a maximum temperature gradient of ± 1°F over a 4-in gauge length (Fig 3) To avoid sharp temperature differentials at the ends of the test section, the length of the furnace is four times the gauge length Sixteen-inch coupons are used to reduce the temperature of the threaded

Jan 1, 1942

By W. David Menzie

An inventory of the mineral resources of a region is a most desirable basis for planning their systematic development. In the absence of appropriate samples upon which to base a reliable estimate of this inventory, investigators have resorted to a number of indirect approaches; the unit regional value concept forms the basis for one such approach. The produced resources of a region are a function of the inventory multiplied by the degree of intensity (some fraction between zero and one) of development of the resources. A convenient means of expressing this amount is to accumulate the weight and value of resource and to prorate the totals over the area represented by the unit sampled; i.e., weight and value per unit area or tonnes/km.2 and U.S. 1967 $/km.2. When this approach is used for the U.S.A., South Africa and Rhodesia the average value of (log.) U.S. 67$/km.2 for the states of the U.S.A. and South Africa are similar ($55,176 U.S.$67/km.2 and $52,538/km.2 respectively); whereas, Rhodesia is relatively underdeveloped (10,301 U.S.67$/km.2). On the other hand if the individual commodities are used as a basis for comparison, the U.S.A. shows the greatest diversity (67 commodities out of a possible total of 75) with South Africa next (50 commodities) and Rhodesia last with only 26 commodities. Since it is known that area and value are not simply related perhaps diversity is directly related to area. Again the kinds of commodities which have contributed the greatest value to the totals differs, with fuels (petroleum, coal and natural gas) predominating in the U.S.A. totals; whereas, South Africa and Rhodesia obtain their greatest contributions from gold, diamonds (S. Africa) coal and asbestos. In each case the greatest value of resources is produced from sedimentary rocks. As a somewhat different kind of example, consider a single resource (petroleum, the most valuable in the U.S.A.) over a large area (number of sampling units) the U.S.A. by state. Furthermore, assume that the model frequency distribution of the amount of petroleum (tonnes/km.2) is log normal. Then the present production fails to match the model first because 17 states (34% out of 50) have not produced petroleum. Of the remaining 33 states, Louisiana has produced the most (15,177 tonnes/km.2) and the frequency histogram is still skew. A spatial analysis of the distribution of the amount of petroleum by state shows that the samples (i.e., states) are not random; the non-producers and producers tend to cluster. It remains to be seen if this is indeed a true representation of the incidence of petroleum or is a function of the exploration philosophy currently in vogue.

Jan 1, 1977

By L. B. Herring, Harold Decker

The area under discussion includes districts 2 and 4, so designated by the Texas Railroad Commission,‡ and comprises the following 26 counties: Bee, Brooks, Calhoun, Cameron, De Witt, Duval, Goliad, Gonzales, Hidalgo, Jackson, Jim Hogg, Jim Wells, Karnes, Kenedy, Kleburg, Lavaca, Live Oak, Nueces, Refu-gio, San Patricia, Starr, Victoria, Webb, Willacy, Wilson and Zapata. Drilling and Production The total number of wells drilled in this district during 1944 is shown in Table 5. This indicates a 41 per cent increase in all drilling over 1943. In spite of every handicap created by the war, the oil industry more than met the crisis in its drilling program in this district. Last year we thought it would be impossible to add to the total number of wells drilled in the previous yea;, but in spite of the lack of manpower and materials this large increase was shown. During 1944 this district produced 124,201,978 bbl. of oil from 11,607 producing wells, whereas in 1943 it produced 88,421,693 bbl. from 11,211 producing wells. This is an increase of 35,780,285 bbl. of oil. Some decline has been shown in the older fields arid undoubtedly the majority of wells in the district are producing at too high a rate. However, this district has been able to meet any demands made on it by the Petroleum Administration for War for increased production. Discoveries and Extensions There was a slight decline over the preceding year in the number of wildcats drilled, but the number of discoveries increased. In most instances subsequent development has not been sufficient to arrive at preliminary valuations but it appears that the over-all oil discoveries were below normal and that the gas discovcrics were relatively more important. The Canolis and Tijerina pools, Jim Wells County, and the Dan Sullivan pool, Brooks County, appear to be the outstanding oil discoveries of the year. No important oil reserve was discovered or developed in the Wilcox horizons, although considerable wildcatting was done. Important extensions were made at Agua Dulce and Stratton in Nueces and Teleberg Counties. Production was increased appreciably at Willimar, Willacy County, and Seeligson, Jim Wells County, in sands and areas which were apparently productive at the beginning of the year. Refining Most of the major work in the changing and completing of refineries was finished this past year. Some of the most modern refineries in the industry are now located in the Corpus Christi area. Gas and Condensate No new recycling projects were commenced during the year, nor were there any important additions to the old operations. No major gas reserve was developed, although several new discoveries suggest potentid possibilities.

Jan 1, 1945

By J. V. Gluck

In the present paper," as in earlier publications, V1&apos;2"1 the authors present experimentally obtained relations for the free energy of solution of oxygen in various metals as a function of temperature. The relations are linear and given in slope-intercept form. The authors give values for the "probable error" of the slope and intercept, citing Birge&apos;s method17 for obtaining these quantities. Birge discussed these methods as early as the period 1929-32.25726 In Ref. 25 the probable error is defined as "the numerical value which the true error is as likely to exceed as not", i.e., 50 pct probability. Ref. 26 is even more detailed and includes a lengthy discussion of the applicability of the Gaussian distribution to experimental data. However, the use of the "probable error" has come under some question. For example, Ezekiel&apos;~ states that "the use of the probable error tends to make the results appear rather more accurate than the standard error". Moroney28 stated (in 1951) that "the term &apos;probable error&apos; was used in former times to denote 0.67450 (50 pct confidence using the normal distribution)....the term, a poor one, is now very much obsolescent. It has been said that &apos;it is neither an error nor probable&apos; ". Bennett and Franklin in 1954" in discussing statistical confidence levels mentioned that "the customary use of the 5 and 1 pct levels is not completely a matter of chance; these levels probably represent a balance between the desire to be correct and the desire to be specific. Also in most people&apos;s minds the borderline between something that could possibly happen and something that is not likely to happen is somewhere in this region. Certainly 5 or 1 pct confidence limits are to be preferred to limits such as the &apos;probable error&apos; which has only a 50/50 chance of including the correct value." It may readily be seen that the 95 pct confidence level, which is given by 1.960 (assuming large sample size), thus corresponds to approximately three times the "probable error". Indeed, limits of one standard deviation correspond only to about 68 pct probability. Even if we triple the tolerances given by the authors in the equations of their Table 111, the deviations do not become excessive and would not detract from their excellent results. It could then be said with a high degree of confidence (95%) that the precision of the slope and intercept values is about *3-5 percent of their absolute levels rather than the 1-2 percent implied in Table 111. This is then compatible with an earlier suggestion of Gokcen~&apos; of a probable accuracy of 5 pct in dissolved oxygen measurements by the technique employed. Alternatively, the authors could have

Jan 1, 1967

By J. F Haseman, J. E. Davenport

In the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1,2 results in inadequate beneficiation. Investigation of Flotation Methods Flotation has been used in &apos;Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate. 3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P2O5 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P2O5, 20; SiO2, 40; CaO, 29; Fe2O3, 2; Al2O3, 3; F, 2.3. The particle-

Jan 1, 1948

By J. F. Haseman, J. E. Davenport

In the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1,2 results in inadequate beneficiation. Investigation of Flotation Methods Flotation has been used in &apos;Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate. 3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P2O5 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P2O5, 20; SiO2, 40; CaO, 29; Fe2O3, 2; Al2O3, 3; F, 2.3. The particle-

Jan 1, 1948

By J. F. Haseman, J. E. Davenport

IN the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1-2 results in inadequate beneficiation. INVESTIGATION OF FLOTATION METHODS Flotation has been used in Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate.3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over-all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P205 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P205, 20; Si02, 40; CaO, 29; Fe203, 2; A1203, 3; F, 2.3. The particle-

Jan 1, 1947

By H. J. Booss

THERE is a considerable lack of data on thermody-namic properties of hard-metal alloys. Only two papers 1,2 give mean values of specific heat in an unknown temperature range; more recently the author3 gave more detailed information on the mean specific heat in the temperature range from 20" to -190° C.. EXPERIMENTAL 1) Method—Experimental procedure followed a method suggested by Oelsen and coworkers.4 It requires samples of some 100 g of weight. The time of measurement is prolonged (30 min) but one run allows the measurement of a lot of positions on the heat-content curve. Results4 were found to be in good agreement with "best values" given by kelley,5 Kubaschewsky and Evans,6 etal.

Jan 1, 1960

By J. K. Gerlach, E. D. Gock, S. K. Ghosh

Previous experience in leaching chalcopyrite with dilute sulfuric acid has generally resulted in rather poor extraction. However, the reaction rate of the leaching process can be significantly improved by an increase of the crystal lattice defects of the chalcopyrite. It was found that subjecting chalcopyrite to a milling action involving a strong impact, permits 99 percent of the copper to be extracted in a subsequent single leaching stage with at least stoichiometric quantities of sulfuric acid in the presence of oxygen at 100°C. Simultaneously, about 80 to 85 percent of the sulfide sulfur content is converted to elemental sulfur. The dependence of the reaction rate of chalcopyrite on the pretreatment, temperature, oxygen pressure and acid content will be discussed.

Jan 1, 1973

By Peter Allen, John R. Stuermer

COUNTRY RISK ANALYSIS A company operating in a foreign country assumes all the risks that it would at home. However, beyond these, it assumes risks that arise from the unique political economic, financial conditions of the foreign country which it operates. These risks are known as "country risks". The importance of these risks has grown appreciably in the recent past (particularly in the aftermath of Iran&apos;s revolution in the late 1970s and Mexico&apos;s debt rescheduling of 1982-83), and as a result, country risk analysis has become an increasingly useful management tool. Country risk analysis seeks to assess the probability that countries will service their aggregate foreign liabilities on time, as well as the probability that country conditions could inhibit the debt service capability of private sector debtors. This involves assessment of both the ability and the willingness of countries to meet their external liabilities in a timely fashion. "Ability" is linked economic and financial variables: principally, a country&apos;s economic structure, foreign exchange generating capacity and debt service burden. On the other hand, "willingness" is tied to political factors: mainly the economic/financial orientation of government policy and the stability of economic decision making. Ability: Economic, Financial Conditions and Prospects Judgement of a country&apos;s ability to meet foreign financial obligations begins with a detailed study of a country&apos;s external balance sheet which compares the country&apos;s foreign liabilities and assets. Table 1 below presents country balance sheets for Mexico, Korea and Indonesia. [TABLE 1 The External Balance 3-,-et 1983 (billions of dollars; Indonesia Mexico Korea Foreign Liabilities 31.1 67.6 40.5 By Source: Bank 13.1 75.5 26.3 Non-bank 18.0 12.114.2 Foreign Reserves 9.3 4.7 7.0] The large magnitude of Mexico&apos;s external debt (nearly $90 billion), compared to foreign assets of approximately $5 billion, illustrates the weakness of this country&apos;s external payments position, and the main reason that it has been unable to meet interest and principal payments and been forced to reschedule its foreign debt. On the other hand, the relationship between Korea&apos;s external liabilities and assets ($50.5 billion and $7 billion, respectively), is more favorable and helps explain why Korea has been able to meet debt service payments and maintain its access to the international capital markets. Indonesia has even a more favorable relationship between its foreign assets and liabilities, $31 billion and $9 billion, respectively, and its debt is mostly lower interest rate, non-bank debt, making for a less onerous debt service burden. After the balance sheet, the analysis turns to a comparison of the country&apos;s foreign exchange earning capacity and its prospective financing requirements, that is, the country&apos;s future earnings stream

Jan 1, 1985