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By R. D. Nelson, D. H. Polonis

ThE eutectoid decomposition of 6 phase in the Cu-Sn system (32.53 wt pct Sn) has been studied by several investigators.1-6 On the basis of X-ray work, Iball and 0wen1 reported complete decomposition of 6 in powder specimens after 7 days at 300°C. Wang and Hansen followed the micro-structural changes during heat treatment of lump specimens and estimated 5 pct decomposition after 63 days at 300°C. Wang and Hansen suggested that the discrepancy between the decomposition rates of powder and lump specimens might be due to the excess surface free energy of powders. The present work involved 6 decomposition in copper-tin alloy powders containing 32.6 wt pct Sn. Alloy lump specimens exhibited entirely 6 phase after fairly slow cooling (air cooling) from above 350°C and were so brittle that they could be easily crushed to-200 mesh in a percussion mortar. X-ray diffraction study of the as-crushed powders (-200 mesh) revealed 100 pct 6 phase with very slight line broadening effects due to strain introduced by crushing. Only 6 phase was detected by X-ray diffraction in powders which were heat treated 4 hr at 450°C and air cooled to room tem-

Jan 1, 1960

By P. A. Beck, A. H. Lutts

IN the classical picture of recrystallization the growth of strain-free grains at the expense of a strained and work hardened matrix is responsible for the softening of cold worked metals on annealing. It has been commonly assumed that the release of the stored energy of cold work, softening, and re-orientation take place simultaneously, as the high angle boundaries of the new recrystallized grains sweep over the cold worked matrix. However, recent evidence indicates',' that, at least in some cases, considerable or even complete soften- ing may take place essentially without reorientation (i.e., without the migration of ordinary high angle grain boundaries, or "recrystallization'') and the release of most of the stored energy of cold work may precede softening. These cases suggest that the property changes, taking place during the annealing of cold worked metals, may be associated with various distinct processes. The study of such situations is of considerable interest, since quantitative data are still scarce. In the present work, isothermal softening curves were determined for a heavily rolled aluminum single crystal, under such conditions that no "recrystallization" occurred until after complete softening. The hardness data were correlated with X-ray diffraction pinhole patterns in order to detect any reorientation that may occur and with measure-

Jan 1, 1955

By SAIMM SAIMM

To the members of The South African Institute of Mining and Metallurgy We have audited the annual financial statements of The South African Institute of Mining and Metallurgy set out on pages 575 to 586 for the year ended 30 June 2005. These financial statements are the responsibility of the Institute?s Council. Our responsibility is to express an opinion on these financial statements based on our audit. Scope We have conducted our audit in accordance with International Standards on Auditing. Those standards require that we plan and perform the audit to obtain reasonable assurance about whether the financial statements are free of material misstatement. An audit includes: examining, on a test basis, evidence supporting the amounts and disclosures in the financial statements, assessing the accounting principles used and significant estimates made by management, and evaluating the overall financial statement presentation. We believe that our audit provides a reasonable basis for our opinion. Audit opinion In our opinion, the financial statements fairly present, in all material respects, the financial position of the Institute at 30 June 2005 and the results of its operations and cash flows for the year then ended in accordance with South African Statements of Generally Accepted Accounting Practice.

Jan 1, 2005

By Monte J. Pool, Rudolph Speiser, George R. St. Pierre, William T. Ebihara

Standard free energies of formation of WO,, W O W20058 and WO3from oxygen and the lower oxide or tungsten have been determined in the tempel-ature range of 700° to 1220°C. A tentative W-O phase diagmm is presented based on the evidence that the oxide phases, "W18O49" and "W20058", are unstable at l~atm total pressure, below 585° and 485°C:, respectively. A HE need for a complete quantitative treatment of the phase equilibria of the W-O system has been establishedby metallurgists during the past few years. In an attempt to improve the endurance of tungsten exposed to oxidizing atmospheres at high temperatures, many research groups have studied the kinetics of tungsten oxidation. The reverse process, the reduction of the tungsten oxides to metal, has also been investigated to some extent. Regardless of the particular interest, metal to oxide or oxide to metal, a knowledge of the phase equilibria of the W-O system is required to properly interpret all of the kinetic observations. It should be possible to state the stable phases of the W-O system at any given values of the chemical potential of oxygen and the temperature.

Jan 1, 1962

By Morris A. Dodd

The nodulizing process described converts finely divided zinc calcine into a dense purified nodule. These zinc nodules are suitable for direct use in zinc reduction furnaces, especially the horizontal retort type, for the production of metal or American process zinc oxide furnaces for producing lead free zinc oxide. The basic reactions are carried out in a rotary kiln fired with fuel oil or natural gas. The combustion gases travel countercurrent to the flow of the zinc bearing material. Volatile impurities are carried out of the kiln with the combustion gases and a small portion of the finely divided feed material. The gases are cooled and the larger solid particles are separated from the stream before the resultant fume is collected in a conventional cloth filter baghouse. The fume collected can be further processed by conventional wet processes for the recovery of metallic values such as lead and cadmium. In order to nodulize the calcine properly it is elevated to a point of incipient fusion. The agglomerated material falling from the kiln is cooled and then sized by passing through a screening and crushing system consisting of a jaw crusher and a roll crusher in closed circuit with a vibrating screen.

Jan 1, 1970

By John Hollow

The Fort Knox Mine is an open-pit operation, located in Alaska, which started commercial production in March 1997. A conventional Mill processes 38,000 tpd of a low grade (1.0 g/t Au) free milling gold ore with a low sulphide component (below 1%). Gold production (11% from gravity and 89% from cyanidation) exceeds 400,000 ounces per year. Short retention time (20h), low temperature (7oC) and low grade make high leaching kinetics critical for the performance of the process. In April 2001, the treatment of True North ore, which represented 14.4% of the mill throughput and contained more sulphide in the form of pyrite, arsenopyrite and stibnite, resulted in a substantial drop in the average gold extraction from 87% to 72.6%. Laboratory investigations indicated that the addition of lead nitrate increased gold extraction to 91.7%. Further investigations were initiated to determine the leaching parameters and the nature of the problem. Subsequently, a lead nitrate addition scheme was implemented at the Fort Knox Mill that resulted in 2002 gold production being in excess of 31,000 ounces higher than that estimated for the non-lead nitrate reagent scheme. This paper presents the results of laboratory investigations and modifications in plant operating practices.

Jan 1, 2003

By Keith Sutherland

AEROFLOAT 25 is a complex mixture of free cresylic acid with aryl substituted dithiophosphoric acids, sulphides, disulphides, etc. Its complete composition has not been published by the makers or patentees. It is prepared1 by the reaction between phosphorus pentasulphide and carefully selected cresylic acid. We have been informed that this acid contains no phenol and very little cresol, and that it is rich in the higher alkyl phenols, particularly the xylenols. The reaction between a phenol ROH and phosphorus pentasulphide may produce compounds of the following type:2 (i) RO S P Diaryldithiophosphoric acids RO SH (ii) R-SH Thiophenolic compounds (iii) R-S-R Diaryl sulphides (iv) R-S-S-R Diaryl disulphides (v) Substituted thianthren compounds Except for the simplest type, the thiophenols (ii), the boiling points of these compounds lie above 300° C. The amounts of the thianthren compounds formed from the five xylenols (if all these are present) will vary according to the position of the substituents. In an attempt to ascertain something of the nature of the reagent, fractional distillation under atmospheric pressure was undertaken. At the outset, hydrogen sulphide was freely evolved, and-particularly below 200° C.-the fractions had the strong odor characteristic of the

Jan 1, 1939

By Robert A. Jankowski, J. Drew Potts

The Bureau has identified several ventilation techniques and a water spray application for reducing auger mining dust levels. The dust control methods all work by improving air quantity or flow patterns at the face. Section airflow is maximized by reducing air leakage between entries. Sealing stoppings with brushed- on mortar, containing fiberglass additives, reduces leakage. Heading airflow is maximized by properly installing the brattice curtain and keeping the area between the curtain and rib free from obstructions. The brattice curtains should be hung at least 1.7 m (5.5 ft) from the ribs in 7.9 m (26 ft) wide entries and kept less than 4.6 m (15 ft) from the face. A tight curtain seal is created by using the mine air pressure to force the curtain against the timbers, roof and floor of the mine. Using both a blowing curtain and an exhausting curtain will isolate some face workers from dusty air. The Bureau designed, auger clearer spray system, can be used to force heading air across the active face. Laboratory tests indicate that the auger clearer spray system can increase face velocities by more than 0.5 m/sec (100 ft/min), while using less than 0.13 l/sec (2.0 gal/min) of water.

Jan 1, 1989

By Pollard D. M

The BHAS copper-lead matte leach process, developed in Port Pirie, South Australia, has been in operation on a commercial scale since 1984. This process is one of many developed in the past decade to recover copper via the hydrometallurgical route. In essence, the process involves a single stage agitative leach of 100% -75um matte in an acid chloride-sulphate lixiviant with gaseous oxygen as oxidant. The leach temperature is controlled at 85¦C to assist with reaction kinetics. After leaching, divalent copper is recovered from solution via a conventional solvent extraction/electrowinning circuit to yield cathode copper for sale The BHAS matte leach process has several unique characteristics cf. other commercial oxidative acid leach processes. These include i) the use of gaseous 02 as oxidant ii) H2SO4 addition as make up acid iii) Relatively low chloride ion addition (30 g/1) For this reason, extensive work was carried out into the optimization of leaching by Meadows (1985). Variables investigated included ò Oxygen flow-rate ò Agitation intensity ò Leach temperature ò Free acidity ò Chloride ion concentration ò Matte particle size The current work was conducted to determine the effect of cooling rate, and hence mineralogy, on the leaching characteristics.

Jan 1, 1989

For the last 4 years political leaders have found a new subject on which they can safely generalize wihout creating opposition. That sub- ject is commodity prices. Since the four-fold increase in OPEC crude oil prices, public atten- tion has focussed as never before on the subject of commodity prices. To advocate stabilizing com- modity prices is akin to supporting motherhood or peace or better schooling. Everyone gives lip service to the concept that stable commodity prices are desirable. It is easier to state the principle than to achieve it. Prices that are stable in terms of one currency may be clearly unstable in terms of a different currency. This has become a troublesome matter since 1971, when the world's leading trading currencies were set free to float in relation to each other. To illustrate: consider copper prices in the interval between the end of 1974 and the end of 1977. During those 3 years, the copper price in pounds sterling on the London Metal Exchange rose by 25 per cent. During that period, however, in terms of the USA dollar equivalent, the price was unchanged. And during that period, in terms of the German mark equivalent the price shrank by 13 per cent while the Japanese yen equivalent f.11 1k

Jan 1, 1978

By Arthur P. Chase

The Tacubaya subway tunnels form a link in the Sistema de Transporte Colectivo (Mass Transit System) of the Federal District of Mexico. One tunnel 327 m long lies to the north of Tacubaya Station, the other 811 m long, to the south. The tunnels are designed to accommodate two subway trains, one inbound and one outbound, within the inside diameter of 8.5 m. The designers of the tunnel lining and tunnel driving equipment were faced with the problem of driving two subway tunnels 9.0 m outside diameter through water-bearing soil, directly beneath a high speed limited access highway, a main drainage sewer, structures including residences and office buildings, on an alignment curved for 67 per cent of tunnel length, and on grades as high as 7 per cent, within a limited budget, with minimum damage to property and disruption to the community. Planning The following decisions affecting tunnel design and construction were made at an early stage of planning: Shield. To enhance safety, a tunnel shield would be used for tunnel driving. Dewatering. The water-bearing soil could be effectively dewatered; consequently, tunnel driving could be performed in free air. Hand Mining. To minimize investment in equipment, hand mining would be used to excavate and support the face.

Jan 1, 1997

By C. H. Jr. Herty

IT Is well known that manganese will desulfurize molten iron through the formation of manganese sulfide, which, being only slightly soluble in the metal, rises to and enters the slag where it remains as such (mixer slag) or is converted partly or wholly to calcium sulfide (basic open¬hearth slag). The reaction by which manganese sulfide is formed is Mn + FeS = MnS + Fe. This reaction is generally considered to be reversible, although McCance1 states- that such is not the case. The experimental work was carried out to confirm the results of Rohl2 as to the reversibility of the reaction. Six trials are herein considered. In each 200 grams of electrolytic iron free from manganese and sulfur were melted in a pure zirconia crucible in a carbon resistance electric furnace. To the melted iron in runs 2, 3 and 4 metallic manganese and manganese sulfide were added to give 1, 2 and 4 per cent. manganese respectively, with 0.50 per cent. of sulfur in each. In runs 5, 6 and 7, metallic manganese andiron sulfide to give the same proportions of manganese and sulphur as in runs 2, 3 and 4, were added. The melts were held molten in the furnace for an hour and then cast into small ingots.

Jan 2, 1925

By Richard W. Hutchinson, Christopher Osterman

The Selebi-Phikwe deposits are hosted by amphibolite to granulite facies gneisses of the Limpopo belt. The sulfide orebodies are generally stratiform and conformable to the foliation of the wallrock gneiss. Sulfide remobilization has occurred during metamorphism and has produced discordant sulfide veins, called piercement structures, which project into the hanging wall and footwall gneisses. These piercement structures are compositionally similar to the stratiform sulfides, consisting mainly of pyrrhotite, pentlandite and chalcopyrite. However, texturally the piercement sulfides are coarse grained and inclusion-free, whereas the stratiform sulfides are fIne grained and packed with tectonically produced wallrock inclusions. The piercement structure sulfides and adjacent stratabound sulfides were sampled and analyzed for major base metals and PGE's for comparative purposes. Most of the piercement structures were enriched in Cu, Ag and Au, whereas they were generally depleted in Ni, Co, Fe and PGE's. Changing elemental associations were observed from the stratabound sulfides and along the length of the piercement structures. This and the degree of depletion or enrichment of various major and minor elements in the sulfide piercement structures may reflect the relative mobilities of metals under metamorphic conditions. The similarities between sulfIde piercement structures and the footwall Cu-PGE veins in some Sudbury deposits are reviewed.

Jan 1, 1994

By Steven J. Page, Robert A. Jankowski, John A. Organiscak

Quartz dust is one of the most significant ongoing health concerns in coal mining today. Since initial verification of this health risk in coal mines by the Mine Safety and Health Administration (MSHA) during the early 1980’s, the Bureau of Mines has conducted numerous studies in underground mines to identify the sources of quartz dust and its fundamental characteristics. The most significant underground quartz dust source was found to be the continuous miner excavating rock. The percentage of quartz was also found to be higher in the smaller size fractions of the dust when significant amounts of rock was mined by the continuous miner. Furthermore, the quartz content in the continuous miner return remained fairly consistent at greater distances from the face. Another important factor speculated to affect quartz dust generation and its size characteristics is the morphology of the quartz bearing rock. The quartz content of dust between mines had notable differences with respect to the differences between kaolinite content in the dust. Low kaolinite content was associated with higher quartz content and increased size segregation of quartz into smaller dust particles, indicating that the free silica bonding matrix in the rock is most likely another factor responsible for quartz dust produced through mechanical comminution.

Jan 1, 1991

By P. Diehl

The increasing public interest in a useful framework for an assessment and classification of mineral resources (1) is documented by the recently published international classification scheme of the United Nations (1979) and the revision of the classification system of USBM/USGS (2) (1980). Moreover, various classification systems are currently discussed by the mining industry. In spite of major improvements in respect of the concepts, definitions, and taxonomy of ore classification, it has been recognized that the recent practice to classify mineral resources does not create truly reproductible results, i.e. an assessment to the various categories which is free from personal bias or corporate interests. Following are geostatistical methods proposed for a standardization and quantification of various terms in ore classification. After critical review of present practices, we will establish a precise link between the classical terms such as "measured", "indicated", and "inferred" reserves and the geostatistical concept of estimation errors. Moreover, we will demonstrate, how the basic criteria of economic feasibility, geologic assurance and recovery can be taken into account by geostatistics. A specific algorithm will be proposed to block out and to classify reserves according to the degree of geologic assurance. . We will also show computerized methods which perform complete classification procedures, and will also exemplify them with typical case studies.

Jan 1, 1982

By John P. Ashby

Mine production blasting is a process of destruction of rock masses in order that ore may be extracted. Many open pit operations are faced with the apparently conflicting requirements of providing large quantities of fragmented rock for the processing plant and of minimizing the damage inflicted upon the surrounding pit slopes. A reasonable compromise between these two conflicting demands may often be found by means of simple engineering of production blasts. Many operations tend to rely too heavily upon traditional wall control techniques to alleviate a problem already created by the production blasts. Various production trends including use of larger blastholes and patterns have tended to aggravate the situation. These practices only serve to emphasize the need for engineered improvements. Central to blasting engineering is the process of systematic trial and evaluation. Changes to the blasting system should be made singly, starting with the simplest. This paper discusses simple changes involving drill pattern control, effective use of blastholes and explosives, adequate delaying including use of down-the-hole sequential techniques, effective utilization of free faces, and avoidance of "choked" situations. By directing explosive effort where it is needed, such changes generally result in a significant improvement of slope conditions and of production blasting cost and performance.

Jan 1, 1983

By W. C. Hittinger, J. McGlassan, J. W. Peterson

THIS paper describes the result of an investigation of the production of thin alloyed layers on a thicker substrate of pure germanium as one step in the manufacture of transistors.' The technique of alloying materials with semiconductors is a means of introducing small amounts of impurity elements in the alloyed regions of the crystal lattice of the semiconductor, the presence of which make possible the junction type transistor.2 Impurity elements are added which will produce regions having differing electrical resistivities and means of carrying current. Impurities are of two types: those which produce an excess of electrons (negative or n-type), and those which produce a deficiency of electrons (positive or p-type) in comparison to the pure semiconductor. The boundary between two such regions is designated a p-n junction. Elements in Group V of the periodic table, having five valence electrons, give rise to an excess of electrons when present substitutionally in the semiconductor lattice (usually germanium or silicon from Group IV). Similarly, elements in Group III create a deficiency of electrons. Germanium and silicon which are substantially free of impurities of either p or n-type are designated intrinsic (i-type) semiconductors. The usual junction ----trarnsistor has

Jan 1, 1958

The reported values of the standard reduction potentials of the Au(NH3)2+/Au(0) and Au(S2O3)23-/Au(0) couples are reviewed to emphasise the experimental difficulties associated with the measurement of reliable values for these quantities. The standard reduction potential for Ag(I)/Ag(0) in ammonia and thiosulphate solutions measured at ionic strength 0.1 and 25oCis used to predict the following relevant stability constants () for gold complexes based on the linear free energy correlations: Au(NH3)2+2 = 1013, Au(S2O3)23-2 = 1024 and Au(NH3)43+4 = 1033. The equilibrium constant ( 10-11) for the reaction Au(S2O3)23- + 2NH3 = Au(NH3)2+ + 2S2O32- based on these stability constants is in reasonable agreement with the measured value (10-9) based on potentiometric and spectroscopic data when excess ammonia was added to a solution of Na3Au(S2O3)2. This also indicates that the stability constant of Au(NH3)2+ is 9 to11 orders of magnitude less than that of Au(S2O3)23-. Relevant species distribution and Eh-pH diagrams based on the revised data are presented.

Jan 1, 2003

By Sparrow GJ, Woodcock JT

A sample of pyrrhotite flotation concentrate pulp assaying 15.3 g/t An, which had been prepared from an old tailing dump, was treated to extract the gold. Examination of the concentrate with an SEM showed that exposed or free native gold, 5-40 ¦m in diameter, was present but no locked gold was observed. Direct cyanidation of the concentrate to extract the gold showed that the flotation reagents present inhibited gold extraction. A range of pre-treatments was used to overcome these effects.Direct cyanidation for 48 h extracted only 26% of the gold, indicating possible contamination of the gold in the concentrate with flotation reagents. Pre-treatment of the concentrate pulp by aeration for 24 h, or by 72 h aeration with an addition of 3.65 kg/t of sodium sulphite, gave little improvement in gold extraction by subsequent cyanidation. Pre-treatment by filtration, or by aeration with 12.5 L/t 100 volume hydrogen peroxide, followed by cyanidation for 48 h gave gold extractions in the range 67-77%. Combining filtration and an aeration period with hydrogen peroxide addition raised the subsequent cyanidation extraction of gold to 88%, giving a tailing assaying 1.5 g/t Au.

Jan 1, 1990

By Boyle L. A

A positive displacement pump is being dev eloped for pipeline transportation of coking coal to a -50 mm export specification, at a con- centration and velocity which will allow trans- portation in the plug flow region, minimising degradation of the particle structure, and pipe= line friction losses which will be lower than with turbulent or laminar flow. Pump operation is by hydraulic powered free pistons in a rota- ting cylinder block, with cylinders overlapped by stationary inlet and discharge ports in a water lubricated and sealed rotary valve. Before this development coking coal was not considered suitable for transportation in slurry form, for the valves in the double act- ing duplex piston pumps used for slurry pipe- lines, cannot be divorced from the slurry and require that the solid material must be reduced to a fines'consistency before introduction into the pump suction. The hydraulic transportation of 10 million tonnes of -50 mm coking coal, 50 km from the Bargo area west of the escarpment to an offshore shiploader at Port Kembla is feasible. Study shows that economic factors alone would dictate that the development of a pumping system for long distance transportation of coking coal is necessary. The greatest impact from the slurry transportation of coking coal would be the

Jan 1, 1976