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By S. Budavari, F. M. C. C. Vieira

For several decades, the deep, gold-bearing reefs of the Witwatersrand basin have been mined predominantly according to layouts characterised by longwalls with strike pillars to provide regional support. During the past few years, a new generation of mining layouts has been introduced. These layouts are based on the concept of grid-type, scattered mining with dip pillars for regional support. As the extraction of the gold reefs extends to depths below 3500 m, the hazards associated with mining at these ultra-depths are expected to increase to unacceptable levels. Therefore, the question arises as to, which of the currently applied mining layouts is most acceptable for the economic and safe extraction of an ultra-deep orebody. The purpose of this paper is to describe a research investigation carried out to evaluate the degree of acceptability, from the rock engineering view point, of four currently used mining layouts for application at a depth-range of between 3500 and 5000 m. Pendant plusieurs décades, les filons d’or profonds du bassin du Witwatersrand ont été exploités d’une manière prédominante selon des dispositions caractérisées par des longues tailles avec des piliers directionnels pour donner un support local. Les dernières années ont vu la naissance d’une nouvelle génération de dispositions de mine. Ces dispositions sont basées sur le concept d’extraction type grille, éparpillée, avec des piliers au pendage pour le support local. Comme l’extraction des filons d’or va à des profondeurs en dessous de 3500m, il faut s’attendre à ce que les risques associés à l’exploitation à ces profondeurs extrêmes atteignent des niveaux inacceptables. Par conséquent, la question se pose de savoir laquelle des dispositions de mine couramment utilisées est la plus acceptable pour l’extraction rentable et sans danger d’un gisement ultra profond. Le but de l’étude présente est de décrire les recherches exécutées pour évaluer le degré d’acceptabilité, du point de vue d’ingénierie de roche, de quatre dispositions de mine couramment utilisées pour application à un rayon de profondeur entre 3500 et 5000m. Die tiefen, goldträchtigen Riffs des Witwatersrandbeckens wurden über einige Jahrzehnte hindurch vorwiegend nach jenen Layouts abgebaut, die von Langfrontstreben mit Schlagpfeilern charakterisiert waren, um regionale Unterstützung zu gewährleisten. Während der letzten Jahre wurde eine neue Generation von Grubenlayouts eingeführt. Diese Layouts basieren auf dem Konzept der Rasterfeld-Art, gestreuter Bergbau mit Neigungspfeilern für regionale Unterstützung. Es kann erwartet werden, daß auf Grund der Tatsache, daß sich die Extraktion der Goldriffs auf Tiefen unterhalb 3500 m verlagert, die Gefahren, die mit dem Bergbau an diesen Ultratiefen verbunden sind, sich auf ein nicht annehmbares Niveau erhöhen. Folglich erhebt sich die Frage, welche der gegenwärtig angewendete Abbaulayouts für die ökonomische und sichere Extraktion eines ultra-tiefen Erzkörpers am annehmbarsten ist. Zweck dieser Arbeit ist es, eine durchgeführte Untersuchung zu beschreiben, die vom Gesichtspunkt der Gesteinstechnik gesehen, vier gegenwärtig angewandte Abbaulayouts für den Einsatz in einem Tiefenbereich zwischen 3.500 und 5.000 m evaluiert.

Jan 1, 2003

By P. B. Moran, J. W. Rozelle, J. Forbort

"The Mount Todd gold mine is located approximately 50 km (31 miles) north of the town of Katherine in the Northern Territory (NT) of Australia just off the Stewart Highway, and was operated in the 1990s until 2000 (Fig. 1). After cessation of mining operations, meteoric precipitation and acid rock drainage (ARD) associated with surface water runoff continued to report to the retention ponds across the mine site. In 2000, the mine site entered receivership and the NT government became responsible for the care and maintenance program. In 2006, Vista Gold Corp. (as Vista Gold Australia) acquired the rights to resume exploration on the mine site without the financial responsibility of the environmental liability. In January 2007, Vista Gold assumed the care and maintenance activities on behalf of the NT government, with the environmental liability remaining with the NT government, with the goal of reinitiating mining operations. A critical path component of the planned upgrades to the existing mine/milling facilities and infrastructure was the development of a cost effective treatment and dewatering program for the approximately 11 GL (2,900 million gal) of ARD stored in the Batman Pit (RP3), which has a maximum capacity of approximately 12 GL (3,200 million gal). The Batman Pit must be dewatered in order for Vista Gold to restart mining activities. The Batman Pit Lake had reached a maximum water depth of 104 m (341 ft) when Vista Gold initiated water level monitoring. It is important to note that the ARD water in the Batman Pit has been pumped into it from 2000 to 2012 under an NT government-approved environmental control program.The Batman Pit has virtually no connection to the local ground water regime due to the intensive silicification that accompanied the mineralizing event. In October 2012, just prior to treatment, the pit lake profile was characterized by uniform acidic pH (approximately pH 3.2) and oxic conditions (dissolved oxygen greater than 7 mg/L with depth (Fig. 2). The temperature decreased by approximately 4 °C (39 °F) between the lake surface and approximately 20 m (66 ft) deep, and remained near constant with increasing depth. This temperature difference is believed to be associated with an existing shallow mine bench at approximately 20 m (66 ft) deep around the perimeter of the pit. Similarly, electrical conductivity, an estimate of total dissolved solids, decreased by about 85 microSiemens per centimeter (µS/cm) from 2,800 to 2,715 µS/cm between 10 and 20 m (33 and 66 ft)."

Apr 1, 2016

By C. L. T. Edwards

In the preparation of this paper, the author has drawn upon experience with the operation of a blast furnace on Ioo per cent scrap burden, which he believes was the first operation of its kind in the country. It is the writer's further belief that much of the controversy attending discussions of this subject may be obviated by the presentation of a group of facts that should be known by blast-furnace men and open-hearth men alike. The following question has been raised: "Is iron made with scrap in the blastfurnace charge as good iron as that made without scrap—made with straight ore?" This question may he answered in part by the fact that the open-hearth men of the Bethlehem Steel Co. who used pig iron from a blast furnace that operated on Ioo per cent steel scrap found no difficulty in the processing of such iron other than that to be expected from the known analysis. Operation with ioo PeR Cent Steel Scrap During World War I, there was accumulated at the Bethlehem plant of the Bethlehem Steel Co. some 22,000 tons of alloy-steel turnings and borings. This scrap material was highly corroded and, like alloy-steel turnings of that period, were as a drug on the market. There was also a small tonnage of scale, rich in valuable metal. If this scrap could be worked up into pig iron, it would become readily available for use in the open hearth. In order that the pig iron should contain the maximum percentage of nickel, permission was obtained from the Management to try a Ioo per cent scrap operation. Accordingly, a blast furnace was swung over on July 25, 1919; the percentage of steel scrap increased more or less irregularly until July 29, when Ioo per cent steel turnings were charged. Except for occasional charges of nickel scale as an "cxtra," the operation continued uninterruptedly until Aug. 27, when the furnace was swung back to the regular ore—ore and scrap—operation. During this period, the furnace production ranged from a low of 568.60 tons to a maximum of 766.07 tons. At the swing-over from Ioo per cent scrap to the regular burden, the charge I hung up rather severely. While the latter condition has nothing to do with the object of the paper, a brief discussion of the experience will appear later because of the interest it may have for blast-furnace operators. In Table I, results of the first and last days of the ore charge are shown, July 25 and July 29. The last column shows a summary for the period of maximum scrap ' charge. The furnace employed had a hearth 16 ft. o in., bosh 22 ft. o in., stock line 15 ft. 6 in., bell 11 ft. 6 in., batter in., bosh angle 73' 27' and a Balrer-Neuman top, with gas-engine blowers. Analyses Range of material charged Gravel: SiO2, 64.40 to 70.54 per cent; A1203, 3.52 to 4.58; CaO, 7.27 to 9.49; MgO, 5.09 to 6.14; FeO, 1.67 to 3.34, ignition loss, 9.50 to 13.83. I

Jan 1, 1942

By C. L. T. Edwards

In the preparation of this paper, the author has drawn upon experience with the operation of a blast furnace on Ioo per cent scrap burden, which he believes was the first operation of its kind in the country. It is the writer's further belief that much of the controversy attending discussions of this subject may be obviated by the presentation of a group of facts that should be known by blast-furnace men and open-hearth men alike. The following question has been raised: "Is iron made with scrap in the blastfurnace charge as good iron as that made without scrap—made with straight ore?" This question may he answered in part by the fact that the open-hearth men of the Bethlehem Steel Co. who used pig iron from a blast furnace that operated on Ioo per cent steel scrap found no difficulty in the processing of such iron other than that to be expected from the known analysis. Operation with ioo PeR Cent Steel Scrap During World War I, there was accumulated at the Bethlehem plant of the Bethlehem Steel Co. some 22,000 tons of alloy-steel turnings and borings. This scrap material was highly corroded and, like alloy-steel turnings of that period, were as a drug on the market. There was also a small tonnage of scale, rich in valuable metal. If this scrap could be worked up into pig iron, it would become readily available for use in the open hearth. In order that the pig iron should contain the maximum percentage of nickel, permission was obtained from the Management to try a Ioo per cent scrap operation. Accordingly, a blast furnace was swung over on July 25, 1919; the percentage of steel scrap increased more or less irregularly until July 29, when Ioo per cent steel turnings were charged. Except for occasional charges of nickel scale as an "cxtra," the operation continued uninterruptedly until Aug. 27, when the furnace was swung back to the regular ore—ore and scrap—operation. During this period, the furnace production ranged from a low of 568.60 tons to a maximum of 766.07 tons. At the swing-over from Ioo per cent scrap to the regular burden, the charge I hung up rather severely. While the latter condition has nothing to do with the object of the paper, a brief discussion of the experience will appear later because of the interest it may have for blast-furnace operators. In Table I, results of the first and last days of the ore charge are shown, July 25 and July 29. The last column shows a summary for the period of maximum scrap ' charge. The furnace employed had a hearth 16 ft. o in., bosh 22 ft. o in., stock line 15 ft. 6 in., bell 11 ft. 6 in., batter in., bosh angle 73' 27' and a Balrer-Neuman top, with gas-engine blowers. Analyses Range of material charged Gravel: SiO2, 64.40 to 70.54 per cent; A1203, 3.52 to 4.58; CaO, 7.27 to 9.49; MgO, 5.09 to 6.14; FeO, 1.67 to 3.34, ignition loss, 9.50 to 13.83. I

Jan 1, 1942

By W. H. Renison

This paper will discuss the operation and maintenance of Molycorp's Quests Mine ore conveyor system. The 1.2 m wide, steel cable-reinforced conveyor belts are capable of transporting 1089 kt/h of ore over a belt line distance of 3124 m with a vertical rise of 518 m. A portion of the conveyor runs in a 2057 m long underground decline on a 10° slope. Six 448 kW motors power the main 1.2 m wide belts.

Jan 1, 1984

By J. L. CAHUTHERS

MANY different methods are used for transporting coal from the working face to the tipple. The common methods are animal haulage, locomotive haulage, conveyor systems, and combinations of these three, supplemented, at times, by the miner pushing his own car. Haulage may also be classified into gathering and main line. Gathering haulage is the collecting of loaded cars to a point whence they are hauled to the tipple. In this there is perhaps more lost-time than on the main line because there are .more units in the operation to be synchronized. Also, conditions on the gathering haul- age, such as bad track, poor timbering, insufficient motive power, etc., cause delays; usually these are corrected on the main-line haulage.

Jan 1, 1931

By A. H. Redfield

BITUMENS have been defined by Abraham1 as substances of variable color, hardness, and volatility; composed principally of 'saturated hydrocarbons, substantially free from oxygenated bodies, sometimes associated with mineral matter; the nonmineral constituents being fusible and largely soluble in carbon disulphide. The principal bituminous substances other than petroleums known to commerce are: (1) native asphaltites, such as gilsonite, grahamite, and glance pitch, distinguished by their hardness and relatively high fusing point; (2) native asphaltic pyrobitumens, such as wurtzilite, distinguished by their infusibility and comparative insolubility in carbon disulphide; (3) mineral waxes, such as ozokerite, distinguished by their high content of crystallizable paraffins; (4) native asphalts, containing varying amouqts of mineral matter, distinguished by the comparative absence of crystallizable paraffins; and (5) bituminous residues obtained from the distillation of petroleum. GILSONITE Gilsonite, a black native asphaltite 98 to 99.9 pct pure, characterized by a hardness of 2 (Mohs), a specific gravity of 1.01 to 1.10, and melting point of 230° to 400°F (K and S), is mined commercially only in northeastern Utah. It occurs in parallel, almost vertical veins, from a few inches to 18 ft wide and 8 to 30 miles long, that cut across the Tertiary limestone and shale of the broad Uinta Basin. The bitumen is believed to have been distilled by heat from oily material in the underlying Green River shale. At the Rainbow mine of the American Gilsonite CO.,† the gilsonite, formerly extracted in an opencut by a system of slopes and benches, is now mined by shafts and horizontal drifts. Little timbering is required. The mining methods of the Raven Mining Co., Chicago, and of the American Asphalt Association and the Utah Gilsonite Co., both of

Jan 1, 1949

By John W. Sweeney

Florida is usually thought of as a vacation State, not a mining State, yet Florida has been the leading producer of phosphate rock in the United States for 85 consecutive years and in 1978 supplied over 80 percent of the national output and over 30 percent of the world's output. It is estimated that the Central and South Florida phosphates contain 412,800 short tons (374,000 metric tons) of U308, that 115,000 short tons (104,000 metric tons) of U308 will be contained in the wet-process phosphoric" acid procsssed in the United States through the year 2000, and that 54,000 short tons (49,000 metric tons) of U308 will be available from Florida. In 1978, two companies in Florida and one in Louisiana went on-stream recovering uranium oxide from wet-process phosphoric acid and three others started construction of U308 extraction facilities. These six companies will have the capacity to recover 2,137 short tons (1,900 metric tons) of U308 annually in 1980, which will account for about 15 percent of our domestic requirements.

Jan 1, 1979

By Henry B. Estabrooks

As was the case with many other mining companies, the management of Tennessee Copper Company' viewed with some interest reports in various trade and technical journals of the successful use of ammonium nitrate-fuel oil mixtures as an explosive in coal stripping., construction, open pit mining and finally in underground salt mines. The apparently simple transition from the "interested" stage to a point where in excess of 80% of the ex¬plosives used is ammonium nitrate may be of interest. The Tennessee Copper Company operates five mines in the "Copper Basin" of Polk County which lies on the extreme southeast corner of Tennessee. The ore is a massive sulfide principally pyrrhotite and pyrite with some tough lime silicates as gangue while the walls are greywacke interbedded with schist. Average total production is 28,000 tons per 5-day week and an average of 422 feet of development per week. Development work consists of 8'x 8?, 6?x 7?, 5'x 7? and 4'x 6? drifts driven with drifter drills using 1-5/8" carbide detachable bits and 5'x 5' raises driven with stopers using 1-5/8" bits.

Jan 1, 1962

By Russell B. Dix

Laboratory column leach tests are used routinely to evaluate gold and silver ores for heap leaching. The tests are fairly straightforward. The ore to be tested is placed into an appropriately sized leach column. Then an alkaline cyanide solution is applied to the ore and the effluent solutions are collected. The effluent solutions are then put through a column of activated carbon to recover the gold and silver in solution. Zinc precipitation may also be used for silver ores or if the ore contains a significant amount of soluble silver in addition to gold. The length of the leach time in the tests is mainly a function of how quickly the precious metals are recovered from the ore. At the completion of the tests, the tailings are assayed and the recoveries determined from the calculated head grades. The amount of ore to be used for a test is partly a function of its crushed size. The coarser the size, the larger the sample size for the column. A typical small-scale laboratory column test on ore crushed to -12 mm (-0.5 in.) would usually be conducted in a 100- to 150-mm-diam (4- to 6-in.-diam) column 2 m (6.6 ft) in height, containing about 20 to 45 kg (44 to 99 lb) of ore. A comparable test on the same ore crushed to 50 mm (2 in.) would be conducted in a 200- to 300-mm-diam (8- to 12-in.-diam) col¬umn 2 m (6.6 ft) in height, containing about 80 to 200 kg (176 to 441 lb) of ore. Larger-scale tests can be run by increasing the column heights and diameters. Conducting larger-scale tests results in increased costs for sample procurement, preparation, and running of the tests. The additional data obtained from

Jan 1, 1989

By S. Saha

Minerals in coals cause operational problems in various coal utilization technology as well as act as diluents towards its heat value. Indian coals contain higher percentage of mineral matter compared to foreign coals due to its formation through drift origin. Efficient utilization of high ash coal in environment friendly manner is the driving force for the development of clean coal technology. Coal gasification offers practical means meeting stringent environment controlled parameters for producing power, chemicals, liquid fuels and clean energy on a single platform. It is well accepted that gasification technology has sufficient potential for sustainable development of clean coal technology. Gasification reactivity depends on the coal properties and operational parameters as well as minerals present in coal. Minerals not only create operational problems but on the other side also catalyze the gasification and combustion reaction. Therefore to understand the effect of minerals on the gasification reactivity, present paper proposes to study the variation of gasification reactivity with various gravity cuts such as, 1.4-1.5; 1.6-1.7; and 1.8-1.9 of a selected coal. Gasification reactivity has been studied by thermo gravimetric analyzer (TGA) using CO2 as gasifying agent. Surface properties such as, true density, apparent density, porosity, surface area of different gravity cut samples have also been measured to find out the effect of gravity separation on gasification reactivity. Keywords: gasification, reactivity, mineral matter, gravity separation

Sep 1, 2012

The Institute was founded in 1894 as the Chemical and Metallurgical Society of South Africa. In 1904 it was reconstituted as the Chemical Metallurgical and Mining Society of South Africa and in 1956 it became the South African Institute of Mining and Metallurgy. The objects of the Institute are to advance the science and practice of mining and metallurgy, to afford opportunities for the interchange and recording of knowledge of mining and metallurgy and to ensure high standards of professional conduct and competence. Membership benefits include monthly issues of the Journal of the Institute, General Meetings at which papers are read, symposia, excursions to mining and industrial concerns and the use of club facilities at Kelvin House. Technical journals received on an exchange basis are available to members at the Johannesburg Public Library. The current membership of the Institute is over 1 600. Membership applications are accepted from suitably qualified persons and the requirements for entrance to the various grades of membership are summarised below. Fellows shall not be less than 30 years of age, shall be university graduates in pure or applied science or shall produce evidence to the satisfaction of the Council that they have successfully completed a co-ordinated course of study in pure or applied science of at least three years' duration at an approved university or institution deemed by the Council to be of equivalent status. Members shall have been employed in senior technical positions in important mining or metallurgical undertakings for at least five years or they shall have practised as mining or metallurgical consultants for at least five years. They shall be practising their profession at the time of application. Entrance fee R10,00, Annual subscription R20,00. Letters of designation: F.S.A.I.M.M. Members shall be not less than 25 years of age and shall be university graduates in pure or applied science, or shall have successfully completed co-ordinated courses of study in pure or applied science of at least three years' duration. They shall have been engaged in work of an approved technical character in the mining or metallurgical industries, of which not less than two years shall have been in positions of responsibility. A candidate shall be practising his profession at the time of his application. Entrance fee R8,00, Annual subscription R18,00. Letters of designation: M.S.A.I.M.M. Associates shall be not less than twenty-five years of age, and shall have been engaged in positions of responsibility in, or associated with, the mining or metallurgical industries for periods of not less than three years. If, however, the candidate for admission to the higher grade of Associate, is at the time of his application, already a Student, he need satisfy the Council only that he is, at the time of his application, engaged in a position of responsibility in or associated with the mining or metallurgical industries. In all cases the applicants shall satisfy the Council that they are fit and proper persons to become Associates. Entrance fee R8,00, Annual subscription R18,00. Graduates shall be not less than 21 years of age and shall be university graduates in pure or applied science, or have completed co-ordinated courses of study in pure or applied science of at least three years' duration at an approved university or institution. They shall not remain Graduate members after attaining the age of 30 years without the permission of Council. Entrance fee R2,00, Annual subscription R12,00. Students shall be persons not less than 18 years of age who are being educated or trained in a manner approved by the Councils to occupy a technical position in or associated with the mining or metallurgical industries and who, furthermore, shall not have attained the qualification required for a higher grade of membership. They may remain Students until they have obtained the necessary qualifications for transfer to a higher grade of membership, but not after the end of the Institute's financial year in which they attain the age of 28 (twenty-eight) years. They shall then transfer to a higher grade to retain membership of the Institute. The Council may relax the provisions of this clause in such cases as it considers appropriate. Entrance fee nil; Annual subscription R5,00. Other. The Council has the power to elect to the grade of Fellow or Member, candidates who may not fulfil all the requirements for entrance to these grades but whose status, professional achievements and practical experience in mining or metallurgy justify such election. Applications. Requests for membership application form should be addressed on the attached form to the Secretaries South African Institute of Mining and Metallurgy, P.O. Box 61019, Marshalltown, Transvaal.

By K. E. Manchester, A. H. Clark

Hall measurements have been made on three groups of silicon samples, which were implanted with boron, aluininunz, and phosphorus ions. Boron and phosphorus implants show essentially bulk properties when annealed under previously determined conditions, i.e., 15 min at 80O°C for boron and 10 mirz at 600°C for phosphorus. Bulk properties were also observed in alu?rrilzum-implanted samples when ion concentrations were below 1015 cm-'; however, strange low- terrzperalure behavior and a falloff in cavrier concenlralion was observed for sarnples with ion concenlration abozle loL5 ern-'. This can be attributed to solid-solubility effects since the average bulk concentration f0.r the sample exhibiting bulk properties was below the reported solubility limit for aluminum, while the other samples were above the limit. A detailed study of the annealirzg process indicates that mobilities reach bulk c~alues in phosphorus implants at 250DC, in alutninurn ivnplanls a1 500°C, and in boron implants at 700°C. A simple rnodel has been proposed to fit the annealing data. THE technique of semiconductor doping by direct injection of energetic ions is being actively investigated at this Center. Previously published results from this laboratory have been concerned with the investigation of implanted profiles in silicon and the electrical properties of the resultant p-n structures.1'2 Diode properties have been reported for junctions produced by implantation of boron ions and aluminum ions in n-type silicon and phosphorus ions in p-type silicon. Reverse current characteristics, as well as electron diffraction data, have indicated that the damage to the structure produced by the energetic ions during the stopping process can be minimized by a gentle anneal. The time-temperature conditions for this process have been determined by sheet resistivity studies.3 Electrical profiles of implanted phosphorus ions which have been determined by a differential sheet conductivity technique are based on the assumption that the carrier mobilities and activation energies for ionization are the same as in bulk silicon. These assumptions have been substantiated experimentally by measurement of Hall constants and resistivities of these implanted structures and the results are presented herein. After establishing the bulk property equivalence of the annealed layers, a more detailed examination of the annealing process was undertaken. The sheet resistivity technique initially used to determine anneal conditions has been broadened to include Hall constants and mobility behavior with anneal conditions. PROCEDURE The three groups of samples prepared for Hall measurements were implanted with mass-separated beams of boron, aluminum, and phosphorus ions. Ion concentrations are equivalent surface concentration and they have been determined from beam current measurements during implantation. Implant conditions for the samples are shown in Table I. In all cases, the samples had previously been angle-lapped on one edge and the junction delineated with an HF-copper sulfate solution. The surfaces of the implanted regions of the samples were exposed to the delineating solution and in the case of the "n"-type phosphorus-implanted areas it was later discovered that a finite amount of surface could be etched by this solution. This is discussed in the section on results. The implanted samples were ultrasonically cut into the conventional six-probe configuration shown in Fig. l.* In the annealing studies, for which the meas- urements were taken at room temperature, the sample holder was a Bakelite jig incorporating tungsten mechanical probes. The other measurements were taken using a conventional liquid-nitrogen cryostat, in which the temperature could be varied from 77" to about 400°K. For these measurements, 10-mil aluminum dots were evaporated onto the samples and alloyed with the silicon at 530°C for 15 min. The samples were then mounted in flat-packs, and gold wires were bonded to the aluminum dots. The electrical measurements were made with either a Leeds & Northrup K-5 Potentiometer, a Keithley 147 Nanovolt Null Detector, or a Keithley 601 Electrometer. No external bias was applied to the substrate. There is an inherent self-biasing arrangement at one or the other of the current contacts (depending upon the polarity), and this suffices to limit the current flow to the implanted layer. The magnetic field was generated by an A. D. Little electromagnet. Except for occasional checks of linearity of the Hall voltage with

Jan 1, 1969

By J. Alfred Berger, O. M. Katz

Selected samples of hydrided Zr-Hf alloys were rapidly quenched to voom temperature and exrtrnined metallographically, by X-ray diffraction, and through micro hardness studies to confirm high-temperutuve data Confirming experiments sllowed that there were five phases in this Lernary system: 1) hextrgonal with lattice parameters similar to that of the initia1 Zr-Hf alloy but slightly enlarged due to dissolved hydrogen; 2) fee with properties of a brittle, intermediate, hydride compound; 3) fct with c/a crvoltnd 1.07 and which appeared as a neetilelike precipitale; 4) hexagonal, designated ?, with c/a ratio of 2.37; and 5) orthorhombic, designated X, with a = 4.67, b = 4.49, and c = 5.093 and whose tnicro-st?ruct~ival nppetrl-nnce depcncled o/i, heat lvecrt~r~ent. The tetragonrrl phase never crppeal-erl witkorct the cubic hydricle. Abpecrrtrnce of 0 and A also tlependet on the hafnium content of the zirconium. A previous paper' on the Zr-Hf-H system described the thermochemical data obtained with a high-vacuum, high-sensitivity mirrogravimetric apparatus. This data presented a fairly complete picture of the phase relationships at elevated temperatures. However, it could not establish the actual crystal structures, lattice parameters, or metallographic disposition of the hydride phases. The present complementary study utilizes X-ray powder patterns along with light and electron microscopy to characterize completely the five hydrided phases found in Zr-Hf-H alloys quenched to room temperature. Crystallographic features of the zr-Hf,2,4 zr-H,5-7 and Hf-H8 systems have been summarized in Table I. Designations of a, ß, and ? were retained in the Zr-Hf-H system for the phase regions through which the pressure-composition isotherms always sloped. However, it was not firmly agreed that these were single-phase regions.' In fact, the region designated y always contained a cubic as well as a tetragonal phase after quenching to -196°C. MATERIALS Preparation of the high-purity Zr-Hf alloys has been described.' The four zirconium alloys which were hydrided contained 37 wt pct Hf (23 at. pct), 51 wt pct Hf (37 at. pct), 73 wt pct Hf (58 at. pct), and 91 wt pct Hf (82 at. pct), respectively. These were designated B-2, B-4, B-6, and B-8. Photomicrographs of the initial alloys showed the material to be quite clean as would be expected from the precautions exercised in producing them. However, there were a number of annealing twins but no other subgrain structure. In addition to the four original alloys, fifteen hydrided samples were observed at room temperature. Hydrogen compositions are given at the top of Tables I1 to V. APPARATUS The phases present at elevated temperatures were studied by quenching hydrided samples to room temperature by two different methods, both under vacuum: 1) fast cooling of the sample tubes of the microgravimetric apparatus1'9 with flowing air and 2) rapid quenching into liquid nitrogen. The cooling rate for 1) was 750° to 250°C in 30 sec. Since the microbalance chamber was not designed to permit very rapid cooling of a hydride sample, all liquid-nitrogen quenching was done in an auxiliary experiment. The auxiliary quenching apparatus consisted of a small-bore, high-temperature furnace, a sealed SiO2 tube containing the sample, and a dewar quenching flask filled with liquid nitrogen. The hydrided sample, previously quenched in the microgravimetric reaction chamber, was placed in a platinum boat in a vacuum-degassed SiO2 tube. A zirconium wire getter and degassed SiO2 rod, to reduce the internal volume, were also in the tube. After sealing the tube under vacuum the zirconium getter was heated to absorb the last traces of gas. Only the sample was heated at the reaction temperature for the desired length of time, and then the tube dropped through the opposite end of the furnace into the dewar. A quenching rate of 200" to 400° C per sec was estimated. Analyses of samples after the auxiliary experiment also showed practically no increase in oxygen or nitrogen content from heating in the SiO2 tube. All of the samples were examined at room temperature by the X-ray powder method. The majority of the powder patterns were obtained with double nickel-filtered CuKa radiation after 8- and 16-hr exposures in an 11.48-cm-diam camera. Cobalt and chromium radiation were also used to spread out the high d value end of the Pattern. Such patterns readily identified the minor phases. NO oxide or nitride lines were found. Where sharp back-reflection lines existed it was possible to reduce the

Jan 1, 1965

By L. Xiao

Using K-oleate as a collector and alizarin red S (ARS) as a modifying agent, selective depression of dolomite from francolite was achieved in alkaline medium (pH> 9.0). Interestingly. ARS was also found to depress igneous apatite in alkaline medium in oleate flotation of apatite and calcite. Adsorption of ARS alone on dolomite was found to be higher than that on francolite. Complimentary adsorption experiments of ARS and K-oleate showed that oleate adsorption was increased in the presence of ARS on francolite but not on dolomite. K-oleate adsorbs on both francolite and dolomite and causes precipitation of calcium and magnesium oleate. Such precipitates can be pre-vented when ARS is present. The results are explained in terms of competitive adsorption for the active sites on francolite and dolomite by oleate and ARS, and their interactions with dis-solved mineral species in solution.

Jan 1, 1988

The United States is being invaded... by foreign capital: marks, francs, pounds, yen, lire, riyals, you name it: On almost any given day the Wall Street Journal will contain a new report on a department store, newspaper, bank, hotel, agricultural land, supermarket chain, et a1 being bought into by overseas investment. Xenophobic reaction to this revised national condition has begun to set in. Several farming states have and are introducing legislation to monitor foreign land buying. In Congress a mandatory reporting requirement for foreign purchasing is currently under study. In a relatively short period the country has reverted from world leader- ship in overseas acquisition and involvement to become itself one of the prime investment opportunities in existence. Interestingly enough, one of the industries where this changed status has become most apparent and pronounced is within that small/regional/segmented/low value-to-weight bulk commodity enterprise ranked as the backbone of the nation's construction activity. U.S. cement making is not only undergoing internationalization in its ownership, but also in its product, equipment, and engineering expertise.

Jan 1, 1979

By John E. Shelton

Flotation tests were conducted on two samples of copper-nickel-bearing ore from the Harriman area near Rockport, Maine. The samples were taken by Roland F. Beers, Inc., for tests in cooperation with the Federal Bureau of Mines. One sample, analyzing 0.38 percent copper and 0.71 percent nickel, was essentially a disseminated pyrrhotite in hypersthene. Pentlandite, chalcopyrite, pyrite, and millerite were intimately associated with the pyrrhotite. Locked series flotation tests on this sample showed that 87.4 and 73.4 percent of the copper and nickel, respectively, could be recovered in a bulk concentrate that analyzed 3.47 percent copper and 5.34 percent nickel. The copper-and nickel-bearing minerals were collected with sulfhydric collectors in an alkaline circuit. Cleaning the concentrate three times with the addition of pH regulators was sufficient to attain the previously stated grades. The second sample had analyses of 0.31 percent copper and 0.85 percent nickel; talc and hypersthene were the main gangue minerals. Bulk sulfide flotation, made while depressing the talc, resulted in a concentrate that analyzed 2.33 percent copper and 4.70 percent nickel at recoveries of 70.0 and 54.4 percent, respectively. Flotation of the talc followed by bulk sulfide flotation resulted in a nickel content greater than 5 percent. A locked-series test showed that 60.7 percent of the copper and 73.7 percent of the nickel could be recovered at grades of 1.29 and 4.85 percent, respectively.

Jan 1, 1961

By Rafael Malagutti Leandro

O objetivo deste trabalho é avaliar a influência do tempo e temperatura de tratamento térmico de solubilização no tamanho de grão e na fração de ferrita de um aço inoxidável dúplex UNS S31803 (SAF 2205). A chapa do aço em estudo foi laminada com 22% de redução de espessura antes de ser solubilizada a 1.000°C, 1.100°C e 1.200°C, por tempos de até 96 h. A fração de ferrita das amostras foi obtida por medidas magnéticas em ferritoscópio e por estereologia quantitativa. Já o tamanho de grão foi determinado por estereologia quantitativa de imagens obtidas por microscopia óptica. Observou-se o aumento da fração de ferrita em função da temperatura de tratamento, como esperado, e foi realizado comparativo entre as duas técnicas de determinação da fração de ferrita com simulações de equilíbrio em Thermo-Calc®. O intercepto médio das amostras solubilizadas a 1.000°C manteve-se praticamente constante em todos os tempos de solubilização estudados, em aproximadamente 9 micrômetros; após 5h de solubilização a 1.100°C, e após 30 min a 1.200°C, contudo, observou-se significativo crescimento de grão nas amostras.

Aug 7, 2018

ALBERT J. PHILLIPS, Chairman E. H. Dix, JR., Past-chairman R. H. LEACH, Vice-chairman R. F. MEHL, Vice-chairman W. M. COME, Treasurer LOUIS JORDAN, Secretary 29 West 39th St., New York, N. Y.

Jan 1, 1937

By B. Yarar, J. Kaoma

[Introduction The critical surface tension of wetting of hydrophobic materials has been investigated extensively by Zisman et al. (1973) and relates the spreading of a liquid on a solid to the surface tension of this liquid. Thus, the Zisman ap¬proach shows that a plot of Cos 0 versus ?LV where, 0 is the contact angle and ?LV is the surface tension of that liquid in contact with its own vapor at equilibrium, produces a straight line which can be extrapolated to Cos 0 = 1. The intercept corresponds to a given ?LV, termed the critical surface tension of wetting. In instances, however, Cos 0 - ?LV plots show deviations from linearity. Alternative approaches have also included the plotting of adhesion tension (YLV COs 0) versus ?LV that can be interpreted in terms of critical surface tension data (Lucassen-Reynders, 1963). It has, on the other hand, been customary to correlate contact angle (0) with flotation data, although it is recognized that surface inhomogeneities and related contact angle hysteresis limit the interpretations pertaining to solid surface structures purely based on contact angle measurements. It is nonetheless true that 0 = 0, when the liquid is water, corresponds to a hydrophilic solid. Determination of ?c by contact angle measurements in conformity with the Zisman approach can encounter experimental difficulties, especially when solids are in a powder form. Following the reasoning that 0 = 0 corresponds to the hydrophilic state of a solid, froth flotation experiments can be expected to produce critical surface tension of wetting values for powders. The technique that consists of plotting recovery (% R) versus YLV data and extrapolating the linear part of the curve to % R = 0, has been described elsewhere (Yarar and Kaoma, 1984). This paper describes some of the results obtained by the extrapolation technique mentioned above, using various sulfide minerals. Experimental Materials Molybdenite (MoS2), sphalerite (ZnS), chalcopyrite (CuFeS2), chalcocite (Cu2S), and galena (PbS) were handsorted from massive samples and ground by mortar and pestle, and the -147 Fan fraction used in experiments; sulfur (S) was a C.P. Fisher Scientific Co. product. Method One gram powder samples were conditioned for three minutes in methanol-water solutions of predetermined surface tension values and the percents recovered by flotation using a Partridge-Smith Cell (Partridge and Smith, 1972), were plotted against each ?LV value. Results and Discussion The percent recovery versus ?LV for two naturally hydrophobic solids, i.e., natural molybdenite, MoS2 and sulfur are given in Fig. 1 where it is seen that ?c for sulfur is obtained as 31.5 mNm-1 whereas for MoS2, it is 26 mNm-1. Figure 1 also shows that the new flotation technique described, for estimating ?c is in good agreement with the Zisman method. Curves obtained by the same method for the minerals Cu2S, PbS, and ZnS are given in Fig. 2, where it is seen that for Cu2S, ZnS, and PbS, the extrapolation to % R = 0 produces ?c values of 36, 39, and 49 mNm-1, respectively. CuFeS2, on the other hand, gives a curve that does not permit such an extrapolation to obtain a ?c value with certainty. Nonetheless, treatment of the powder with 8.95 X 10-5 molar Na2S solution followed by washing with distilled water prior to flotation effects the curves as shown in the insert of Fig. 2. The type of curves given in Fig. 1 appear to be typical of hydrophobic solids as can also be observed in Fig. 3. The composition of solid and minute quantities of inclusions seem to play an important role as regards the surface properties of solids as reflected by the slope of the linear part of the % R - ?LV plot and thus the value of ?c obtained by this technique. For example, MoS2 of three different stoichiometric compositions produced the data given below. [MoS2MoS2.14Mo1.002S2 ?c mNm-126.029.031.0 dR/d?8.95.23.0]]

Jan 1, 1985