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By M. J. Richards, A. Szirmae, G. R. Speich

The formation of austenite from ferrite, ferrite plus retastable carbide, spheroidite, and pearlite has been studied in a series of irons, Fe-C alloys, and plain-carbon steels using fast heating techniques. In the absence of carbide, austenite nucleates at ferrite/ferrite grain boundaries; nucleation is followed by the rapid growth characteristic of a massive transfornation. The trarnsformation occurs at 950°C at heating rates of 106º C per sec and cannot be suppressed. Metastable carbide dissolves before austenite forms and does not influence the transformation kinetics. For spheroidite structures, austenite nucleates preferentially at the jinction between carbides and ferrite grain boundaries. Growth from these centers proceeds until the carbide is completely enveloped; subsequent growth occurs by carbon diffusion through the austenite envelope. For pearlite structures, austenite nucleates preferentially at pearlite colony intersections. Carbide la)?zellae dissolve at the advancing austenite interface but complete solution of carbide does not occur; the residtial carbide is eventually dissolvled or spheroid-ized depending on the carbon cuntent. The magnitude and temperature dependence of the austenite growth rate into Fe-C pearlite when incomplete carbide dissolution is assumed are satisfactorily explained by an approximate colume diffusion model. The impurities present in plain-carbon steel reduce the growth rate of austenite in comparison to that jound in an Fe-C alloy. The formation of austenite has been studied in much less detail than the decomposition of austenite. This is primarily a result of the importance of harden-ability in determining the mechanical properties of steel. Recently, more interest in the kinetics of austenite formation has resulted from the discovery by Grange1 that rapid heating techniques strengthen steel by refining the austenite grain size. Although the strengthening effect is not large, it is accompanied by no loss in ductility. In addition, interest continues in rapid heat treatment of low-carbon steel sheet for tin plate applications.2,3 Among the few systematic studies of austenite formation are the early work of Roberts and Mehl4 on formation of austenite from pearlite and recent work of Molinder5 and of Judd and paxton6 on formation of austenite from spheroidite. Also, Boedtker and Duwez7 and Haworth and paar8 have recently studied the formation of austenite from ferrite in relatively pure iron, Kidin et al.9,10 have studied the formation of austenite in 8 pct Cr steels, and Paxton has recently discussed various aspects of austenite formation in steels." The present work was undertaken to determine the kinetics of austenite formation for a variety of starting structures including ferrite, ferrite plus metastable carbide, ferrite plus spheroidal cementite, and ferrite plus pearlitic cementite. Emphasis was placed on determining the active sites for austenite nucleation, determining the temperature and time range of austenite formation, and in the case of pearlite a careful study of the growth rate of austenite was made in the absence and presence of impurities. By using a variety of heating techniques including laser-pulse heating, it has been possible to study austenite formation in an isothermal fashion over a wide range of temperatures. EXPERIMENTAL PROCEDURE The alloys studied in the present work are a zone-refined iron with 4 pprn C, an Fe-C alloy with 130 pprn C, 2 Fe-C alloys with 0.77 and 0.96 wt pct C, and a plain carbon steel with 0.96 wt pct C. The zone-refined iron and Fe-C alloys contained 60 pprn and 200 pprn total substitutional impurities, respectively. The plain carbon steel contained 2400 pprn Si, 2000 pprn Mn, and 900 pprn Cr. Various heat treatments were given to these alloys to produce different starting structures of equiaxed ferrite, ferrite plus metastable carbide, fine pearlite, and spheroidite. These heat treatments are given in Table I. A wide range of heating rates were employed in this work because many of the reactions occur so quickly at temperatures in the austenite range that they are completed during the initial heating cycle unless very fast heating rates are used. Essentially the same heating techniques employed by Speich et a1.12 and Speich and Fisher13 were used in this work. For time intervals of 2 sec to 20 hr, simple hand immersion of 0.010-in. thick specimens in a Pb-Bi bath was employed. These specimens were quenched in a 10 pct NaC1, 2 pct NaOH aqueous bath. For time intervals of 100 m-sec to 2 sec, an automatic dunking and quenching device was employed with 0.002-in. thick specimens. Again, liquid Pb-Bi baths were used for a heating medium but now helium gas quenching was employed. For time intervals of 2 to 100 m-sec a laser heating device was employed with 0.002-in. thick specimens; a helium plus fine water-droplet spray was now used for quenching. Additional information on heating times shorter than 2 m-sec was obtained by study of the zones around the centrally heated laser spot. Here diffusion of heat from the centrally heated zone raises the temperature of the specimen locally to all temperatures between ambient and the peak temperature, but for times of the order of microseconds. All the heat-treated specimens were examined by

Jan 1, 1970

By Thomas Lien

Session 1 - 1 :45 p.m. June 27,1983 Session Chairman: Robert Diederich, RMCMl New Mexico Vice President MR. DIEDERICH: My name is Bob Diederich. I'm the Vice President of the Rocky Mountain Coal Mining Insitute from the state of New Mexico. I'm manager for the Navajo Mine down in Farmington, New Mexico and also project manager for a new mine that we're starling out called the LaPlata Mine in the same area. I'd like to introduce our first speaker, Tom Lien, who is the Director of Engineering, Western Division of AMAX Coal Company. He is responsible for mine engineering, environmental engineering, permitting, reclamation, planning, design and construction plant engineering, geology exploration, quality control and real estate for the western division. He joined AMAX Coal in 1980 as manager of preparation and quality control and was promoted to the present position in August of 1980. He has a BS in geological engineering from the University of Mines. He previously worked for Kennecott Copper Corporation in Hayden, Arizona from 1965 to 1970. He worked as a mine geologist, mill metallurgist, mine production and quality control engineer. Then he also worked for Kaiser Steel Corporation, Eagle Mountain Iron Ore Mine in California, 1970 to 1977. Worked as a process engineer and plant superintendent at that place. Then he was employed by Mountain States Research and Development, Tucson, Arizona, 1977 to 1980. Worked as a senior process engineer and operation superintendent. So without further ado, I'd like to introduce Tom Lien who will talk to us about in-pit coal crushing. MR. LIEN: Thank you. The official title of my paper is actually: "The Potential for In-pit Crushing and Conveying in the Powder River Basin." The Powder River Basin is one of the world's greatest coal fields and will be an important source of energy for centuries to come. About one-eighth of the total coal producing in the United States comes from the Powder River Basin, and 80% of that production came out of Campbell County, Wyoming last year. The growth potential for this region is projected to be enormous. New technology as well as existing technology will make this region competitive in the decades ahead. Understanding the current status of the Powder River Basin, both the mines and the possible methods for changing the means of material movements within the mines is the subject of this paper. Each mine is unique and must be addressed specifically with respect to the application of new and innovative technology. There have been many articles written in the recent mining journals on in-pit crushing and conveying applications, including costs which are very interesting, enlightening and I recommend it to any company considering such installations. To give a general setting, coal production in the United States for 1982 was about 800,000.000 tons. This production came principally from six regions, the Applachian fields of the eastern states where bituminous and anthracite coals have been mined for almost three centuries, the Midwest located in the upper Mississippi Valley are often typified by huge dragline operations of which AMAX Coal has three right now, the Gulf Coast lignites. feeding mine-mouth power plants, the vast bituminous and sub-bituminous departments of the four-comer states of Arizona, New Mexico, Colorado, and Utah and finally the Northern Great Plains lignite fields currently being developed and the sub-bituminous fields of Wyoming and Montana including the Powder River Basin. The Powder River Basin is located in northeastern Wyoming and southeastern Montana. It is the darker brown right in the center of the photograph. It covers about 18,000 square miles. It's about 200 miles north to south and 90 miles east to west. The shallowest coal beds are on the east, north and northwest edges of the basin with deeper coal beds toward the center in the same kind of relationship. There are about one trillion tons of coal in the Powder River Basin, that's the one with 12 zeros after it. A little bit there. It's distributed in many, many seams at varying depths to the current topography. In 1982 about 100,000,000 tons of coal were mined from this basin and 81,000,000 tons came out of Campbell County, Wyoming. The mines that are shown in this picture are basically the mines around the basin edge. The principal area that I'll be talking about is the eastern edge down where it shows Gillette and Campbell County. Geologically the Powder River Basin is the Fort Union formation. The most productive seam is called the Wyodak or Smith-Roland or the Anderson-Canyon, depending upon where you're located. There is about 75 billion tons of coal in the Wyodak of which about 23 billion are considered surface minable today. the coal is Paleocene in age, about 40,000,000 years old. It was once an old Delta region very much like the Mississippi River delta is today. During the 100,000 years that this deposit was formed peat accumulated to a thickness of about 365 feet. It was later buried and turned into sub-bituminous coal. All of the coal mined in the basin is for the electric utility industry. This coal has a typical BTU content of about 8,300 plus or minus about 4,500 BTUs. Sulfur coal is generally less than'/ %, thereby making this coal compliance coal in terms of sulfur dioxide emissions, 1.2 pounds of sulfur dioxide per million

Jan 1, 1983

By James W. Dally, William L. Fourney, Anders Ladegaard Peterson

For the past 3 years, the authors have been conducting research sponsored by the National Science Foundation (RANN) to improve the process of excavation by drilling and blasting. The approach followed has been experimental where the development of stress waves and fractures initiated at the bore hole have been investigated in order to obtain a complete understanding of the dynamic fracture process. The second step in the approach has been to introduce modifications in the drill and blast procedure which will permit closer control of the fracture process. The laboratory investigations involve high speed photography where the dynamic fracture process is recorded with a Cranz-Schardin 1, 2 multiple-spark camera. The camera is equipped with 16 spark gaps which are pulsed at 25 K volts to produce an intense but very short (0.5 sec) flash of light. The camera is capable of recording 16 photographs of a dynamic event at framing rates which can be varied from 30,000 to 1,500,000 frames per second. The exposure time is sufficiently short to stop motion associated with detonating explosive charges and to make visible the details of the fracture process at a bore hose. The bore hole in a massive intact rock formation is modelled with a two dimensional plate containing a circular hole to represent the bore hole. The model material employed is a transparent polyester known commercially as Homalite 100.* This polymeric material is extremely brittle as evidenced by its extremely low fracture toughness of [ ]. The fracture toughness is a measure of the ability of a material to resist the propagation of flaws or small cracks. In comparison, Schmidt3 has recently measured the fracture toughness of Salem limestone and determined [ ]. Thus, the Homalite 100 should closely model the brittle nature of rock where fractures occur at small flaws and propagate without any apparent plastic deformation. Homalite 100 is also birefringent, which indicates that it becomes optically anisotropic when subjected to either static or dynamic loads. Circularly polarized light is transmitted through the loaded Homalite 100 model in a polariscope4 and the birefringence produces an optical interference pattern which is called a fringe pattern. For dynamic photoelasticity, the multiple-spark camera is equipped with polaroid filters to produce the circularly polarized light required to generate the photoelastic fringe patterns. An example of a singlespark frame showing a fringe pattern from a typical experiment is presented in [Fig. 1]. The photograph was taken 0.000072 sec (72 sec) after the detonation of the explosive charge. The circular fringes are due to the outgoing dilatational or P type stress wave and travel with a velocity of 85,000 in. per sec (2260 m/sec) in the Homalite 100. The P wave is followed by a second lower velocity stress wave known as the shear or S type wave which propagates at a velocity of 49,000 in. per sec (1245 m/sec). In the local neighborhood of the bore hole, several radial cracks are visible. These cracks propagate at essentially a constant velocity of 15,000 in. per sec (380 m/sec) prior to arrest. The fringes about the crack tips and in the local region of the bore hole are primarily due to the residual gases contained in the bore hole after the explosive charge was detonated. Sixteen frames similar to this one are recorded during the experiment to give full field visualization of the dynamic event at 16 discrete times over its duration. The fringe order number N is related to the difference in the principal stresses of and 02 according to a stress optic law4: [ ] where f0 = material fringe value, and h = model thickness. The wholefield dynamic-fringe patterns provide a basis for simultaneously observing the interaction between propagating cracks and the stresses which drive these cracks. Fracture Control Experiments Improvements in the efficiency of the drill and blast procedures must involve close control of the fracture process following the detonation of an explosive charge in a bore hole. By control it is implied that the number of cracks initiated and the location of each crack on the wall of the bore hole can be specified. Control also, involves orienting each crack and maintaining the crack path and velocity until the specified crack length is achieved. If the entire fracture process can be controlled, then rounds can be designed to optimize volume removed. fragment size and minimize costs. One area of blasting where fracture control is vitally important is in underground excavation where the strength and stability of the rock walls must be maintained and smoothness and precision of the walls must be achieved. The smooth blasting method is one of the most commonly employed procedures for achieving some degree of fracture control. In smooth blasting, the central region of material is first removed, and then the final row of closely spaced undercharged or cushioned holes are fired to remove the final volume and produce a smooth wall. In some instances, unloaded or dummy holes between the loaded holes are recommended to guide the fracture plane. This investigation pertained to an evaluation of 3 features of the smooth blasting process. These included (a) the effect of stress reinforcement on fracture by simultaneously firing 2 charges; (b) the influence of a dummy hole on control of the fracture planes between 2 simultaneously fired charge holes; and (c) the influence of dummy hole spacing on fracture plane control.

Jan 1, 1979

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

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

Jan 1, 1970

Council for 2001/2002 President A.A.B. Douglas President-elect S.J. Ramokgopa Vice-presidents Senior F.M.G. Egerton Junior Prof. T.R. Stacey Immediate Past-president Dr L.A. Cramer Honorary Treasurer J.A. Cruise

Jan 1, 2002

Council for 2004/2005 President F.M.G. Egerton President-elect W.H. van Niekerk Vice-presidents Senior R.P.H. Willis Junior R.G.B. Pickering Immediate Past-president T.R. Stacey Honorary Treasurer J.A. Cruise

Jan 1, 2005

NORTH AMERICA ALASKA Anchorage.-Strandberg, H. College.-Barber, R. J. Burns, W. T. Henricksen, R. Henton, H. M. Joesting. H. R. Comet.-Albertson, F. E. Douglas.-Cahill, W. E. Carlson, L. Fairbanks.-Andregg. J. N. Coats. R. R.

Jan 1, 1938

By Thomas O. Llewellyn

This U.S. Bureau of Mines (US8M) report presents a concise review on sources, processes, supply, and historical use patterns of domestic cadmium. It also covers a preliminary estimated cadmium material balance for the year 1989.

Jan 1, 1992

By W. P. Tassie

"The location of Quintette Coal Limited's pit and geologic nature of the dedicated reserves combine to introduce a number of factors which can adversely affect waste dump stability. These factors include steep terrain, variable material type and particle size, and precipitation. This paper describes the solutions to related waste dump problems. A brief geotechnical review of the waste dumps is given and Quintette's operational strategies during periods of dump stability and instability are discussed. Despite best efforts, failures occasionally occur and discussion is included on Quintette's success at bringing failed dumps safely back into use. IntroductionQuintette Coal Limited (QCL) property is situated on the east slope of the Rocky Mountains near Tumbler Ridge, approximately 120 km southwest of Dawson Creek, British Columbia. This surface mine is located on mountainous terrain and is subject to climatic extremes of wind, precipitation and temperature.QCL has a current annual production of 4.75 million tonnes of metallurgical coal with a stripping ratio of 7: 1. This correctly implies the extraction of over 33 million bank cubic metres (BCM) of waste which is placed on numerous waste dumps at various locations adjacent to the pit.There are three active pits - Mesa, Wolverine, and Shikano (Fig. 1). Shikano is a newly developed pit located near the plant site. It is not subject to the unique dump management problems prevalent at the Mesa and Wolverine pits, therefore the Shikano waste dumps will not be discussed . The dumps for the Mesa and Wolverine pits, however, have been technically and operationally challenging since shortly after mining began in 1983. Their characteristics, associated site specifics and management, therefore, are the subject of this paper.Waste dump locations for the Mesa and Wolverine pits are shown in Figures 2 and 3 as is the relative waste distribution . Each sub-pit is connected to a principal waste dump in reasonable proximity and to an alternate dump should disposal of waste be necessary at that location. The waste dumps are constructed from the top elevations initially to coincide with subpit development. In the initial years (1-7) waste dump heights will be up to 250 m, then, as the mine matures (8-20 years), waste dump maximum heights will continuously decrease to coincide with lower bench elevations"

Jan 1, 1988

By Pierre Pelletier, Christian Pichetie, Dany Pelletier

"IntroductionDuring 1994, Barrick Gold Corporation, or American Barrick at that period, acquired Lac Minerals Ltd. properties. One of them was the Bousquet Complex. The Bousquet Complex is divided into three distinct sites: one mine (Bousquet 2), one mill (Est-Malartic Division) and one analytical chemistry laboratory (Les Terrains auriferes).The Bousquet 2 orebody was discovered in 1986. An extensive drilling program was initiated during 1987. Its purpose was to define the orebody extent. At the end of 1997, gold production from Bousquet 2 mine reached 1 290 000 oz of gold from a cumulative tonnage processed of 4 175 000 M t.From 1980 to 1990, the Est-Malartic mill was used for the treatment of ore from the following mines: Bousquet 1, Doyon, Francoeur (Richmont mines), Buckshot and 1-20 stope. At present, the mill receives ore from the Bousquet 2 mine.HistoryThe Est-Malartic mill is located close to the town of Malartic in northwestern Quebec. The mill has been in production since the mid-1930s processing mainly clean, free mill gold ore. Active mining from the Est-Malartic mine ceased in 1979 and the mill was switched to custom processing of ore from La Mine Doyon located some 45 km to the west. This custom processing arrangement continued until 1987 when the expanded mill at Doyon became available to process all of the Doyon production. In 1987, Lac Minerals commenced development of a second mine at the Bousquet site located 40 km west of Malartic. This mine, Bousquet 2, is the down dip extension of the La Ronde mine developed by Agnico Eagle. During pre-feasibility evaluations, it was determined that it would be economically more attractive to modify the existing Est-Malartic mill than building a new facility on site. The Est-Malartic mill modifications for processing the high-copper/gold ore from the Bousquet 2 mine began in late 1989 and new circuits were commissioned in mid-1990. Modifications included the addition of both gravity and flotation concentration circuits to recover coarse, free gold and the copper. The existing leaching circuit continued to operate on the copper flotation tailings with some minor modifications to the solution handling systems. A S02/Air cyanide destruction circuit was also added at that time."

Jan 1, 2000

The effect of the acid anion on chalcopyrite (CuFeS2) leaching has been examined, under care-fully controlled conditions, in a range of acids in stirred reactors and columns at pH 1. Solution speciation calculations were performed to determine possible rate controlling species. The relative leach rates of Cu from chalcopyrite can be explained quantitatively entirely on the basis of solution speciation without the need to invoke any rational relating to surface structure/speciation. The rate of Cu release is first order with respect to Fe3+ activity. The stirred batch reactor experiments at fixed pH 1, Eh of 750 mV (SHE) and 75? showed that chalcopyrite leach rates were rapid in both HClO4 and H2SO4 with 100 % Cu recovery after 73h and 97 h respectively. However, the leach rate in H2SO4 was slower as compared to the HClO4 leach due to a slower increase in the activity of Fe3+ due to the formation of FeSO4-. The presence of Cl- in the leaches (HCl and H2SO4 with NaCl added) led to a low activity of Fe3+, due to the high activity of FeCl2+, resulting in very low leach rates. For the column leach experiments dissolution of Fe and Cu from the pH 1 HCl eluted chalcopy-rite column with quartz proceeded at several times the rate of any other column (chalcopyrite with and without quartz and/or pyrite leached with H2SO4, HClO4, HCl). Three observations indicate that galvanic interactions are not leach rate controlling role: a) The rate of leaching was not uniformly increased for the non-quartz containing columns where the two sulfide minerals, chalcopyrite and pyrite, are in physical contact as compared to the analogous quartz containing columns where they are not; b) there exists a reasonable correlation between the rate of Cu release and the effluent Fe3+ activity for all columns thus indicating that the leach rate is determined by solution speciation; and c) in all cases the proportion of Fe to Cu leached indicates preferential dissolution of Cu over Fe. However, the proportion of Fe to Cu is either the same or greater than expected on the basis of the single mineral columns. This indicates that chalcopyrite is not dissolving preferentially to pyrite.

Jan 1, 2014

"9.1. IntroductionThe application of the electron microprobe for in-situ analysis of platinum-group minerals (PGM), first reported by Borovskii et at. (1959) and Genkin (1959), has resulted in a large and growing number of analyses of unidentified PGM. Figure 9.1 charts the increase in the number of reported unnamed PGM from 3 in 1959 to a total of 140 by 1979. All these minerals, with the exception of UN1968-5 and UN1976-1, were analyzed with the electron microprobe.There has been no recognised or consistent manner for reporting these unidentified PGM. The term ""unnamed"" has been used for PGM that are considered to be new species without a name as well as for PGM which have not been exhaustively compared to named and unnamed PGM in order to determine whether they are unique. Many of the so-called unnamed PGM are only unnamed because of poor microprobe analyses or because larger than usual elemental substitutions (e.g. Bi for As) made it difficult to calculate the true stoichiometry. Some of these unnamed minerals have been unnecessarily reported with non-stoichiometric formulae, as discussed by Cabri (1980). Better analytical data should reduce the number of such proposals.Another unfortunate practice, possibly used to by-pass the scrutiny of the Commission for New Minerals and Mineral Names, I.M.A., is to report a ""new"" mineral by referring to it by a name made up of its constituent elements. This has been deplored on a number of occasions (cf. discussion in Fleischer et at., 1976, p. 179), but some journals continue to publish such material. The Commission on New Minerals and Mineral Names has not recognised such ""names"" and these PGM are therefore included in this chapter with other unnamed minerals.A further example of inconsistency in reporting unnamed minerals is the report of a specific ideal stoichiometry with no further details (not even an analysis!); sometimes with a note that this will be reported in greater detail in due course. I have also been guilty of similar optimistic statements and, unfortunately, it is not always possible to follow through. Some authors give data on unnamed minerals in such a discrete manner that their reports are sometimes missed in literature searches."

Jan 1, 1979

By Dongfeng Yun, Zhu Liu, Wendong Cheng, Zhendong Fan, Dongfang Wang, Yuanhao Zhang

"For the purpose of studying the strata behavior due to multislicing top coal caving longwall mining along-the-strike direction in steeply dipping extra thick coal seams, the shield support pressures of the upper and lower slices of Panel 37220 in Dongxia coal mine were monitored using the KJ513 dynamic monitoring system. The set up rooms adopted the ""horizontal line-arc segmentinclined line"" form and used different types of shield supports. The results showed that the strata pressure of upper slice Panel 37220- 1 changed slightly along the strike direction, while along the dip direction it exhibited strong to weak pressure from bottom to top. The first weighting interval oflower slice Panel 37220-2 was about 60.Sm, and the average periodic weighting interval were about 22.6m. The strata behavior of Panel 37330-2 exhibited a spatiotemporal characteristic in that periodic weighting occurred first in the middle-upper part, followed by the middle and upper parts, arc segment, and finally the lower part. During the periodic weighting, the weighting interval and intensity also exhibited strong space characteristics. The average dynamic load coefficient was 1.48 and the maximum lateral pressure of the side shield was 900Kn1000kN.INTRODUCTIONThe steeply dipping coal seams account for about 20% of proven reserves and 10% of output in China (Wu, et al, 2014). About 50% of coal mines in the western China have steeply dipping seams, and are mainly distributed in Sichuan, Gansu, and Chongqing. When the steeply dipping seam employs longwall mining along the strike direction is employed, the waste rocks fill the gob nonhomogeneously along the dip direction, the deformation and caving of surrounding rocks exhibit obvious asymmetry and spatio-temporal characteristics with particular and complex strata behavior. In recent years, many research on strata behavior due to multi-slicing top coal caving longwall mining along-the-strike direction in steeply dipping extra thick coal seams have been performed with significant results (Shi, 1999; Wu, Xie, and Ren, 2010; Wu, Xie, Wang, and Ren, 2010). However, there has been a complete lack of research on monitoring the strata behavior due to multi-slicing top coal caving longwall mining in steeply dipping extra thick coal seam. Therefore, the strata behavior of multi-slicing panels 37220-1 and 37220-2 was systematically investigated in this research."

Jan 1, 2016

By Z. T. Yuan, S. L. Qi, Q. Y. Meng, G. X. Zou, L. X. Li, J. W. Lu

"The selective separation of pentlandite from serpentine was investigated using the magnetic coating-magnetic separation by conditioning slurry pH value, stirring speed, magnetic field intensity and controlling dosages of sodium hexametaphosphate (SHMP) and sodium oleate (SO). The optimal conditions were obtained as follows: the natural pH of 6.8-7.2, stirring speed of 2300 rpm, magnetic field intensity of 200 kA m-1, SHMP of 60 mg L-1, and SO of 400 mg L-1. Results showed that the Ni grade was 21.4% with the Ni recovery of 81.4% and MgO content of 3.8%. FTIR indicated that coagulant of SO preferentially adsorbed onto pentlandite and magnetite particles while hardly adsorbing onto serpentine particles when dispersant of SHMP was used. Thus, the selective magnetic coating of magnetite on pentlandite surfaces could be formed due to the formation of hydrophobic surfaces on pentlandite and magnetite particles. Calculations of DLVO (Derjaguin-Landau-Verwey-Overbeek) confirmed that interactions between magnetite and pentlandite particles were much stronger than that between magnetite and serpentine due to the existence of hydrophobic interactions with the addition of SO. Consequently, the pentlandite was successfully separated from serpentine. Therefore, this study exhibits the potential to separate pentlandite from serpentine minerals for practical applications and thus deserves further study.INTRODUCTION It has been well known that the separation of pentlandite from serpentine is a worldwide challenging issue in the processing of sulphide nickel-copper ores. Because the serpentine mainly causes the high pulp viscosity and slime coating on the pentlandite, which interferes with the flotation process of pentlandite (Cao et al., 2016; Feng et al., 2012). The fibrous mineral of serpentine preferentially attaches to the air bubble, forming bubble-macrofibers aggregates, which are easily floated into the nickel concentrates, resulting in the decline of Ni grade (Alvarez-Silva et al., 2016; Kusuma et al., 2014). In addition, the serpentine adheres to the pentlandite as hetero-aggregation or slime coating via the electrical interaction between a positively charged serpentine and a negatively charged pentlandite, which decreases the hydrophobicity of pentlandite and reduces the collector adsorption (Feng et al., 2013). In other words, the presence of serpentine or the occurrence of aggregates (including slime coating and bubble-macrofibers aggregates) seriously affects the flotation process and deteriorates the flotation of pentlandite. Therefore, we propose magnetic separation to recover pentlandite from serpentine with the magnetic coating technology, which is different from forth flotation."

Jan 1, 2018

By C. J. Hodgson

"IntroductionThe Kitsault molybdenum mine is located 600 m above sea level on Lime Creek at the head of the inlet of Alice Arm, 140 km northeast of Prince Rupert, British Columbia (Fig. 1). The area of the mine is characterized by rugged topography and deeply incised creeks which drain into the Portland Canal - Observatory Inlet fjord system. Access is by air service or boat from Prince Rupert, or by secondary road from Terrace.HistoryThe deposit was first staked in 1911, at which time the main feature of interest was a polymetallic vein with values in silver located southeast of the molybdenum deposit (Woodcock and Carter, 1976). Turnbull (1916) was the first to describe the extensive zone of crisscross quartz stringers with Mo~ on the banks of Lime Creek which, although it clearly had the potential to contain a major deposit, was not then believed to be of commercial value.Following sporadic work through the 1920s and early 1930s, the ground lay idle unti11959 when Kennco Explorations (Western) Ltd. commenced a program of systematic diamond drilling on the stockwork-type mineralization adjacent to Lime Creek. In late 1964, Kennco announced the presence of an orebody of 36 million tonnes averaging 0.230Jo Mo~ which was to be put into production at a rate of 5450 tonnes per day under a new subsidiary, British Columbia Molybdenum Ltd., at a forecasted capital cost of $20 million (revised to $30 million).Between January 1968 and April 1972, 10.4 million kg of molybdenum were produced from 9.3 million tonnes of ore, at an average mill recovery of 90%, and with an average mill head grade of 0.206% Mo~. The ore was mined in an open pit with a waste to ore stip ratio of 1.5 to 1. The. operation was terminated in 1972 due to weak molybdenum prices and escalating costs.In May 1973, the deposit was purchased by Climax Molybdenum Corporation of British Columbia, Limited and in 1979 transferred to affiliate Amax of Canada Limited. In early 1979, following completion of detailed feasibility studies, Arnax commenced construction of the 10 900 tonne per day Kitsault mine and mill complex based on a mineable open pit reserve of 95 million tonnes at an average grade of 0.192% Mo~, sufficient for a mine life of 25 years. The original capital cost estimate was $143 million; it subsequently escalated to approximately $200 million."

Jan 1, 1995

By T. Watanabe, T. Fujiwara, S-Wan Song, Y. Nakagawa, R. Teranishi, M. Yoshimura

"We have tried a new method to fabricate cobalt oxide films onto porous films directly by an electrodeposition method at low temperature. The porous substrate films (membrane-filters and paper) were placed between a LiOH solution and a CoSO4 solution and treated under electricalcurrents of ~1 mA/cm2 using carbon cathode and anode. Effects of temperatures, flow rates and electrical currents have been investigated. As a result, we found that the LiCoO2 films and Co3O4 films were successfully fabricated on paper and/or porous PTFE filters at 120 °C, while CoOOH films were deposited at 80 °C. Furthermore, LiCoO2 films did not cover whole the substrates, but was fabricated just over the anode. The chemical reactions of electrodeposition were also considered from these results.IntroductionRecent agreements on global environment require rapid progress of low energy and low waste processings for the fabrication of thin films that succeed to the mass production with the high consumption of the fossil fuels (such as physical vapor deposition°, chemical vapor deposition2), dip coating with firing3) etc.). The total energy consumption in aqueous systems should be the lowest among all these processing routes, because a much lager excess of energy is necessary to create melts, vapor, gas, or plasma than to form an aqueous solution at the same temperature4)The ""Soft Solution-Process"" (SSP) was one of the best concepts to realize the low energy and low waste fabrication of inorganic materials. 5) Generally speaking, SSP gives similar results to any other processing using fluids (such as vapor, gas plasma) and/or beam/vacuum processing. However, the total energy consumption of all these possessing routes should be the lowest in aqueous systems, because a much larger excess of energy is necessary to create melts, vapor, gas or plasma than to form an aqueous solution at the same temperature (Fig. 1). This idea can be easily demonstrated, using the simple example of BaTiO3. The energy diagram for the formation of BaTiO3 from various precursors is shown in reference 4. The driving forces (AG) for representative synthesis reaction of (1), (2), (3), (4) and(5) are 38, 727, 3685, 17 and —14kcal/mol, respectively at room temperature. 6' 7"

Jan 1, 2000

By A. Wahab Khair, N. P. Reddy

"Abstract. This paper attempts to characterize coal breakage due to the action of a continuous miner. Coal blocks with an approximate dimension of 45.7 cm x 38.1 cm x 15.2 cm (18 in. x 15 in. x 6 in.) were first subjected to confining pressures equivalent to in-situ conditions. Then the blocks were cut along the face/butt cleat direction by the cutting head of a specially designed unique automated rotary coal cutting simulator (ARCCS), simulating the action of the continuous miner in underground coal mining. During the tests under a particular set of in-sit and operating parameters a number of other para¬meters such as 1) penetration of bit into coal, 2) penetration resistance (thrust and cutting pressures), 3) rotating velocity of cutting head, and 4) acoustic emission activity in the coal block were monitored. After each cutting cycle the fractured surfaces were photographed and a velocity survey was conducted by using a sonic technique. At the end of each experiment, the coal fragments were collected in bags using a vacuum cleaner. These fragments were then sieved for size distribution analysis.INTRODUCTIONThe Federal Coal Mine Health and Safety Act of 1969 was enacted to emphasize to coal operators that the airborne respirable dust should not exceed 2 mg/m3, thus reducing the incidence of coal workers pneumoconiosis.Dust control techniques such as conventional water sprays and dust collectors are only partially effective and require additional equipment expenditures. A more authentic approach would be to reduce respirable dust at the source the continuous mining machine cutting head, where the fragmentation process occurs. Improving the fragmentation process by characterizing the coal breakage will not only reduce respirable dust at the face, but it will also decrease the amount of respirable dust that is liberated during the secondary handling.Fragmentation of coal can be expressed as a function of three major groups of parameters; namely, 1) coal properties, 2) in-situ conditions and 3) operating parameters. This paper deals with the characterization of coal breakage along both face/butt cleat direction as a function of machine operating parameters such as rate of advance, angle of attack, bit configuration, bit lacing (spacing and pattern) and bit speed at various equivalent in-situ pressures."

Mar 1, 1989

By John Greenhalgh, Benoit De Rivaz

"1.INTRODUCTIONThis paper is based on the Lee Tunnel project, which is part of the Thames Water’s Tideway Project.The client Thames Water has employed a Joint Venture of Morgan Sindall/Vinci Grand Projets and Bachy Soletanche to undertake the design and construction of the project under the management of CH2M Hill at a value in excess of £635 million ($990 million). The 3 main elements of the tunnel project are the permanent lining to the shafts, the precast segmental lining, and the secondary tunnel lining. All 3 are reinforced with different types of Dramix steel wire fibers and each element will be described in this paper. This paper will explain the basic structural design of FRC (Fiber reinforced concrete) based on the post cracking residual strength and will describe all the steps of this project, the testing program realized to support innovative design and the main technical learning point. 2.BASIC CONCEPT OF STRUCTURAL DESIGN OF FRCStructural design of FRC elements is based on the post-cracking residual strength provided by fiberreinforcement. Other cases, like early age crack-control or fire resistance, are considered non-structural use of FRC. For structural use, a minimum mechanical performance of FRC must be guaranteed. Fibers can be used to improve the behavior at SLS (Service Limit State) since they can reduce crack spacing and crack width, thereby improving durability and they can be used to improve the behavior at ULS (Ultimate Limit State) where they can partially or totally substitute conventional reinforcement in most conditions. In the design process, the definition of the performance expected from the material is a key factor and the mix-design of the material has to be optimized for the required structural performance.2.1 Testing Method In order to ascertain the structural requirements requested for the FRC segments, preliminary testing on samples will be required and the effectiveness of the proposed solution will also be assured by the testing of full scale specimens. Structural design of FRC elements is based on the post-cracking residual strength provided by fiber reinforcement. For structural use, a minimum mechanical performance of FRC must be guaranteed. Nominal values of the material properties can be determined by performing a 3-point bending test on a notched beam according to EN14651 (figure 1) or by performing the ASTM C1609 beam test:"

Jan 1, 2016

By S. Dixon

The Peñasquito deposit in Mexico is one of the largest grassroots precious metal discoveries to be developed in recent years Minera Peñasquito, S.A. de C.V., a wholly owned subsidiary of Goldcorp Inc. has developed the Peñasquito property to produce an average annual production of 120 kilo tonnes of lead containing 0.5 million oz gold and 30 million oz silver and 250 kilo tonnes of zinc containing 20 thousand oz gold and 1.5 million oz silver.. The Peñasquito property covers a 39,000 hectare area in the western half of the Concepción del Oro district in the northeast corner of Zacatecas State, Mexico, approximately 24°45'N latitude/101°30'W longitude. The sulfide reserves are 1,009 million tonnes containing 3.33 million tonnes of lead, 7.26 million tonnes of zinc, 16.5 million ounces of gold and 1,000 million ounces of silver. This paper describes the work to justify the installation of HPGR to process the crushed SAG mill pebbles. The SAG mill-Ball mill and Cone crusher (SABC) grinding circuit is an accepted grinding circuit to optimize power consumption relative to throughput. The installation of a HPGR to process the Cone crusher product is the next step in optimizing the power consumption relative to throughput. The concern expressed by some on not moving on this change in grinding circuit was the capital and operating cost of a HPGR. The pebbles from the SAG mill can be the more abrasive and are the harder particles of the ore. Goldcorp investigated the wear and processing costs of installing a HPGR to treat the crushed pebbles. The conclusion was the installation would be a very positive economic factor in the operation of Peñasquito. This is the first application of the use of HPGR to process crushed pebbles. Minera Peñasquito and M3 Engineering selected Polysius to provide a Polycom 2417 with twin 3,000 kW drives. A comprehensive test program at Polysius has shown for 7% higher circuit power it is possible to get approx 23% higher tonnage. This coupled with full utilization of the SAG and Ball Mill available energies and with proper control of the transfer size from the SAG to Ball mill circuit there is the potential to achieve 65,000 mtpd per line or a 30% total capacity increase. The cost of power for this increase in grinding capacity is 4.8 kWh/tonne.

Jan 1, 2010

By H. J. Morel-Seytoux

The influence of patten, geometry on assisted oil recovery for a particular dispIacement mechanism is the object of investigation in this paper. The displacement is assumed to be of unit mobility ratio and piston-like. Fluids are assumed incompressible and gravity and capillary effects are neglected. With these assumptions it is possible to calculate by analytical methods the quantities of interest to the reservoir engineer for a great variety of patterns. Specifically, this paper presents (1) very brieily, the methods and mathematical derivations required to obtain the results of engineering concern, and (2) typical results in the dorm of graphs or formulae that can be used readily without prior study of the methods. Results of this work provide checks for solutions obtained from programmed numerical techniques. They also reveal the effect of pattern geometry and, even though the assumptions of piston- like displacement and of unit mobility ratio are restrictive, they can nevertheless be used for rather crude but quick, cheap estimates. ,These estimates can be refined to account for non-unit mobility ratio and two-phase flow by correlating analytical results in the case M = 1 and the numerical results for non-piston, non-unit mobility ratio displacements. In an earlier paper1 it was also shown that from the knowledge of closed form solutions for unit mobility ratio, quantities called "scale factors" could be readily calculated, increasing considerably the flexibility of the numerical techniques. Many new closed farm solutions are given in this paper. INTRODUCTION BACKGROUND Pattern geometry is a major factor in making water-fIood recovery predictions. For this reason many numerical schemes have been devised to predict oil recovery in either regular patterns or arbitrary configurations. The numerical solutions, based on the method of finite difference approximation, are subject to errors often difficult to evaluate. An estimate of the error is possible by comparison with exact solutions. To provide a variety of checks on numerical solutions, a thorough study of the unit mobility ratio displacement process was undertaken. To calculate quantities of interest to the reservoir engineer (oil recovery, sweep efficiency, etc.), it is necessary to first know the pressure distribution in the pattern. Then analytical procedures are used to calculate, in order of increasing difficulty: injectivity, breakthrough areal sweep efficiency, normalized oil recovery and water-oil ratio as a function of normalized PV injected.

Jan 1, 1967