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By J. T. Williams

The equilibria in the V-Zr alloy system were investigated by solidus temperature determinations, thermal analysis, dilatometry, electrical resistance measurements, microscopic examination, and X-ray diffraction analysis. There is a eutectic reaction at 1230°C between a compound, V2Zr, and a solid solution containing 10 pct V in ß zirconium. V2Zr decomposes at 1300°C into liquid and a solid solution containing about 10 pct Zr in vanadium. The eutectic composition is probably about 30 pct V. A eutectoid reaction between V2 Zr and a zirconium takes place at 777°C at a very high rate. The eutectoid composition is 5 wt pct V. The limit of solubility of zirconium in vanadium was estimated to be 5 pct at 600°C. No attempt was made to determine the liquidus for the system. THE recent availability of large quantities of high purity zirconium has stimulated the study of zirconium binary systems. The equilibrium diagram for the V-Zr system has received little attention, however. Wallbauml appears to have made the first report concerning the equilibria in these alloys. He reported the existence of a compound, V2Zr, having the MgZn2 ((214) type of structure with a, = 5.277 kX and c° = 8.647 kX. Anderson, Hayes, Roberson, and Kroll2 made a survey of some potentially useful zirconium binary alloys and found that zirconium probably dissolves a small amount of vanadium. They reported the probable existence of a compound between the two elements and suggested that the zirconium-rich solid solution undergoes a eutectic reaction with this compound. Pfeil," in a critical review of the existing information, estimated that the solubility of vanadium in zirconium is less than 4.7 pct and probably less than 1.8 pct. Rostoker and Yamamoto' proposed a partial diagram for the V-Zr system in a survey paper on vanadium binary alloys. Their diagram indicates the compound, V,Zr, a eutectic reaction at 1360°C, a peritectic reaction at 1740°C, and a limit of solubility of zirconium in vanadium of about 3 pct. They obtained no information on the equilibria in the zirconium-rich alloys. In view of the potential utility of the V-Zr alloys and the incomplete knowledge concerning the equilibria in the system, an attempt was made to establish the constitutional diagram. Preparation of the Alloys Raw Materials: The vanadium for making up these alloys came from the Electro Metallurgical Corp. Zirconium came from two sources. In the beginning of the investigation, sponge zirconium from the Bureau of Mines was used in making some of the alloys. Later, iodide metal made at the Westinghouse Atomic Power Development Laboratories became available. This material was used in the preparation of all the dilatometric and resistance specimens and about two-thirds of the solidus temperature specimens. A typical manufacturer's analysis of the vanadium is shown in Table I. No other analysis of the vanadium was made. The metal contained a dispersed second phase and did not have a sharp melting point. Typical results of spectrographic analysis of the Westinghouse zirconium are shown in Table 11. These data indicate a very high purity. The Bureau of Mines sponge metal was probably less pure but had good ductility. Melting: All of the alloys used in the investigation were made by melting pieces of vanadium and zirconium together in a dc electric arc furnace similar to those of Geach and Summers-Smith, craighead, Simmons, and Eastwood," and others. Melting was done in an atmosphere of helium scavenged of residual air by the preliminary melting of a separate charge of zirconium. Each ingot was turned over and melted at least three more times before removal from the furnace to aid in the attainment of homogeneity. Alloys prepared for use in the investigation are listed with the results of solidus determinations in Table III with the exception of the following compositions upon which no solidus determinations were made: 0.29, 0.54, 4.57, and 5.55 pct V. Analysis: The weight of each ingot made from iodide zirconium was within 0.1 g of the total weight of the initial charge, about 90 g. Since each component of each charge was weighed to the nearest 20 mg for amounts less than 10 g and to the nearest 0.1 g otherwise, the gross composition of an ingot could be calculated accurately. Chemical analysis for the vanadium content of several alloys agreed

Jan 1, 1956

By Paul A. Farrar, Sanford Adler

The Tz-Zr system was reinvestigated using both metallographic and X-ray diffraction techniques. It mas found that titanium and zirconium are soluble in all proportions in both the a and 0 phases. The minimum in the transformation was found at 50 at. pct and 535°C. THE Ti-Zr system has previously been the subject of several investigations.1-6 A study by Hayes et al.5 using magnesium-reduced titanium and zirconium melted in graphite crucibles indicated that the system was a continuous series of solid solutions in both a and 0 phases with a minimum at approximately 545°C and 65 pct* Zr. These conclusions were substantiated by the work of Duwez4 as well as by the limited data of Craighead et al.,3 Fast,2 and deBoer.1 However, a more recent investigation of the titanium-rich region by Ence and Margolin6 indicated that the solubility of zirconium in a, titanium at 500°C is approximately 22 pct with the a + 0 field extending to approximately 47 pct Zr. Therefore in order to resolve this discrepancy the following investigation was initiated. EXPERIMENTAL PROCEDURES The alloys used in this investigation were prepared as 20- to 30-g buttons by nonconsumable electrode melting in a helium atmosphere. The starting materials were Bureau of Mines titanium BHN 71 and Foot Mineral iodide crystal bar zirconium. The titanium was premelted before the alloys were prepared to avoid excessive weight losses in the final melting. The compositions of the alloys prepared were as follows: Ti-12.0, 22.2, 29.0, 39.6, 50.0, 58.0, 72.4, 83.5. and 90.8 pct Zr. Prior to heat treatment the alloys were cold-worked 10 to 20 pct, stopping at the first signs of cracking. Specimens for heat treatment were wrapped in molybdenum sheet and annealed in argon-filled quartz capsules for the following times and temperatures: 750°C-40 days, 675°C-61 days, 600°C—109 days, 575°C-112 days, 550°C-112 days, 525°C—127 days, and 450°C —153 days. Quenching was accomplished by breaking the capsules in an iced-brine solution. The standard techniques used for polishing the specimens involved belt grinding, grinding on emery paper, polishing electrolytically, and etching with Remington "A" etch7 or "R" etch 8 Debye-Scherrer X-ray powder photograms were obtained for a number of samples using a 114.6-mm camera and CuK, radiation. Exposure times were from 19 to 23 hr. The powder samples were obtained by filing to 270 mesh size. After filing the samples were wrapped in molybdenum foil capsuled in quartz and reannealed at the temperature of original heat treatments; following the heat treatment the samples were quenched into iced brine without breaking the capsules. RESULTS AND CONCLUSIONS On the basis of the microstructural examinations of the heat-treated samples and of the X-ray diffraction data, the Ti-Zr system shown in Fig. 1 was constructed. The data indicate that a, and 0 titanium form a complete series of solid solutions with a, and ß zirconium. The minimum was found to occur at 65 pct (50 at. pct) Zr and 535°C, in excellent agreement with the earlier investigations,1-5 see Fig. 2. Typical microstructures are shown in Figs. 3 to 7. Figs. 3, 4, and 5 show the 22.2, 39.6, and 50.0 pct Zr alloys after heat treatment at 675°C for 61 days with the 22.2 pct Zr showing equiaxed a, and the 39.6 pct Zr alloy a + transformed ß, while the 50 pct Zr alloy shows only transformed B. Figs. 6 and 7 show the 29.0 and 39.6 pct Zr alloys after heat treatment for 109 days at 600°C. The 29 pct Zr alloy has an equiaxed a structure while the 39.6 pct alloy has a partially recrystallized a+ß structure. X-ray diffraction data obtained from the 50 pct Zr alloys after heat treatment at 525° and 450°C as well as from the as-cold-rolled material showed only those lineos which couid be indexed a: a titanium, good agreement with the previously published values. 2,4,5,9 Chemical analysis* of the 50 pct Zr alloy

Jan 1, 1967

By D. Symonds, A. J. Richins, M. G. Nelson

Mining activities at Kennecott and Park City are similar in that both districts began operation in the last half of the 18th century, and both operated on a large scale for many years. However, they are different for two important reasons. First, the mines in the Park City district were always underground. The city of Park City was above many of those mine workings. It remained in place after mining activity stopped, and as the town became an up-scale ski resort, property values increased and the tailings dumps became desirable building sites. In contrast, Kennecott eventually became an open pit mine, and the growing pit effectively consumed many of the small towns that had grown up around the underground workings. Even towns on the periphery of the open pit were razed to allow for the disposal of mine overburden. Second, almost all mining activity stopped in Park City in the early 1970s, just as the first rigorous environmental legislation in the United States was enacted. (The Ontario mine operated sporadically until 1982.) Many mine structures, waste dumps, tailings ponds and piles, and tunnels were simply abandoned as the mining companies ceased business or were bought out. Kennecott, on the other hand, has been in continuous operation for more than 120 years, so a closure plan hasn’t been finalized. Nonetheless, environmental and social problems have arisen, and the company has had to address them.

Mar 2, 2022

By G. R. Speich

The age-hardening of Fe-18 to 21 pct Ni marten-sites containing small amounts of titanium, aluminum, copper, or molybdenum has been studied by hardness measurements, transmission electron microscopy, extraction replicas, and X-ray and electron diffraction. Electyon-probe microanaly-sis of the extracted precipitates was also performed. The effects of temperature and composition on the aging kinetics have been studied in detail for Fe-20 pctNi-Ti martensites. All the martensites hardened on aging at 400" to 500 "C with a single aging peak. In the Fe-20 pct Ni-Ti martensites hardening was associated with the formation of a fine dispersion of NisTi. No precipitation was observed in an Fe-21 pct Ni-2.6 pct A1 alloy, although the hardness increased from 327 to 545 Dph during aging at 500°C. Copper was precgitated from an Fe-20 pct Ni-10.7pct Cu alloy, at least during the later stages of age-hardening. Hardening in an Fe-18 pctNi-5pct Mo alloy was associated with the precipitation of an intemzetallic compound tentatively identified as NisMo. The formation of austenite and the recovery of the defect structure of the martensite occur simultaneously with the normal precipitation processes &ring aging of these alloys. THE age-hardening of substitutional iron-base martensites is an important subject for study since age-hardening is one of the principal means of strengthening the substitutional iron-base martensites in precipitation-hardening stainless steels1 and maraging steels.a~4 In contrast to Fe-C martensites, these martensites are relatively soft and ductile in the as-quenched condition, due to the low solid-solution hardening effect of the substitutional elements as compared to carbon. Also, these martensites are hardened by aging at low temperatures in contrast to Fe-C martensites which generally soften on tempering. Titanium and aluminum, singly or in combination, as well as copper and molybdenum, have been the principal alloying additions to stainless steels to obtain an age-hardening reaction.' Ti-A1 or Co-Mo-Ti combinations have been used in the case of maraging steels.3"4 The present work has been confined to a study of the age-hardening characteristics of an Fe-18 to 21 pct Ni martensite to which small amounts of titanium, aluminum, copper, and molybdenum were added in order to study the effects of each alloying element separately. The age-hardening characteristics of these alloys were studied by hardness measurements, transmission electron microscopy, extraction replicas, and X-ray and electron diffraction. Electron-probe micro-analysis of the extracted precipitates was also performed. The types and dispersion of precipitates, the formation of austenite, and the recovery of the defect structure of the martensite all were studied. EXPERIMENTAL PROCEDURE Alloys. The compositions of the alloys studied are listed in Table I. The level of 18 to 21 pct Ni was chosen to assure complete transformation to martensite on quenching and is the same nickel level used in maraging steels. The concentrations of titanium, aluminum, copper, and molybdenum were chosen after a study of the and Fe-Ni-MO' ternary diagrams indicated how much of each alloying element could be dissolved in austenite. Alloys with 2 and 4 pct Cu were also studied, but these did not show precipitation hardening at the nickel levels studied and will not be discussed further. The alloys were prepared from electrolytic iron and high-purity* nickel, titanium, aluminum, copper, and molybdenum by vacuum melting of 20-lb heats. These were vacuum-cast into two 8-lb ingots. The ingots were hot-rolled at 1100°C to 1/2-in. plate and subsequently cut into 3/8-in. cubes. The cubes were sealed in evacuated quartz capsules and solution-treated for 16 hr at 1100°C, except for the Fe-20 pct Ni-10.7 pct Cu alloy and Fe-21 pct Ni-2.6 pct A1 alloys which were solution-treated for 16 hr at 1300" and at 1250"C, respectively. Metallographic examina-

Jan 1, 1963

By R. J. Wagner, Jim Douglas Jr., P. M. Blair

The calculation of the behavior of an oil reservoir during a water flood has long been an important problem to reservoir engineers. Buckley and Leverett derived the differential equation which describes the displacement of oil from a linear porous medium by an immiscible fluid, but this equation could not be solved by the methods of classic mathematics. Consequently, in order to integrate the equation over the length of the reservoir, they neglected the effects of capillary pressure. In the present paper, a numerical method has been developed for determining the behavior of a linear wafer flood with the inclusion of capillary pressure. The differential equation which was derived for the case of incompressible fluids is second order and non-linear. This differential equation was approximated by an implicit form of difference equation which is second order correct in both time and distance. An electronic digital computer was used to perform the numerical solution of the difference equation. INTRODUCTION The problem of calculating the flow and distribution of fluids in an oil reservoir subjected to a water flood has long challenged the reservoir engineer. The ability to solve this problem would provide a valuable tool for the design and study of field waterflooding programs. One of the first contributions in this field was made by Buckley and Leverett,'.' who developed a method of calculating waterflood performance in a linear reservoir. Their technique was limited by the practical necessity of excluding quantitative consideration of capillary pressure. It is the purpose of this paper to describe a method for calculating the behavior of a linear water flood with capillary pressure considered. This method, although limited to the linear case, should serve as a step toward the solution of the two- or three-dimensional waterflooding problem which would better describe actual reservoirs. The treatment of the problem has assumed that the reservoir is linear and homogeneous, that both oil and water -are incompressible, and that gravitational forces may be neglected, and that water is injected into one end of the reservoir and oil and water are produced from the other. Though these assumptions may not be directly applied to natural reservoirs, they are closely approached in many laboratory core tests. The physical problem has been represented mathematically by means of a differential equation and suitable boundary conditions. The inclusion of capillary pressure effects causes the differential equation to be second order and non-linear, and it is not amenable to solution in terms of known functions. Nevertheless, the problem may be solved by numerical methods. The authors have successfully calculated several cases from the start of water injection through to depletion of the hypothetical reservoir using realistic capillary pressure and relative permeability characteristics that were chosen so as to put the method to a test. DIFFERENTIAL EQUATION In two- or three-dimensional problems involving the two-phase flow of incompressible fluids through a porous medium, the equations describing the flow can be combined until only two dependent variables remain, the pressure in one phase, and the saturation of one phase. These variables appear in two simultaneous partial differential equations which are derived on the basis of the conservation of mass and Darcy's law as applied to two-phase flow. For a linear reservoir, this system of differential equations can be reduced to a single, nonlinear, second order, parabolic differential equation for one dependent variable, the saturation of one phase. For incompressible water and oil, it can be shown that the conservation of mass for the two phases requires that Where q, and q, are the water end oil flow rates per unit cross-section, s, and so their saturations, +, the porosity, and [ and T, dimensional space and time variables. Darcy's law, as extended to two-phase flow including the effect of capillary pressure,' is as follows:

By G. F. Revey

Most commercial explosives contain 70 to 94% ammonium nitrate, by weight. When portions of these explosives end up in shot rock and ore, through spillage or incomplete detonation, ammonia and nitrates leach out of them and into ground water. In recent years, State and Federal regulators have been applying more stringent water quality standards, particularly at new mines and development projects. Bulk ANFO, a mixture of ammonium nitrate and fuel oil, is the explosive of choice at most mines, when mining conditions allow it. ANFO use is desired because it less costly than other explosives, but when spilled it dissolves readily in water. Several case histories in the United States and Canada show a clear connection between uncontrolled losses of bulk explosives and high nitrate levels in mine effluents. Mining companies have tried several approaches to reduce ammonia and nitrate levels in groundwater. They either control explosive losses, or they treat mine effluents at the end-of-pipe. This paper addresses the former solution --controlling explosive losses. For both packaged and bulk explosives, guidelines designed to limit losses during storage, handling and use, are described in detail.

Jan 1, 1995

By M Paine

Work to date has shown that the use of Al2O3:SiO2 ratios is suitable for identifying potential gibbsite bearing samples. Subsequent confirmatory XRD analysis helps to refi ne this initial identification. Based on these methods, gibbsite has been mapped in two common settings:1. along steeply dipping faults, or2. in association with stratigraphic contacts where one unit is aluminous (eg dolerite dyke).On a microscale, the samples studied illustrate an association between haematite, and to a lesser extent goethite, and gibbsite. In these samples the dominant forms of haematite are discrete crystalline phases of either microplaty haematite or euhedral martite. Two dominant forms of gibbsite were observed:1. vein style gibbsite that cross-cuts (or infill’s cracks) in goethite ± martite iron oxide associations, or2. gibbsite intimately intergrown with and forming a matrix to microplaty haematite.Some of the gibbsite microstructures observed suggest that it may have formed during a very hydrous, possibly hypogene, stage of the deposit genesis. The presence of this crystalline gibbsite, and in particular its occasionally euhedral form, suggests that aluminium was at some stage remobilised, at least locally, and then precipitated from solution as gibbsite. These observations suggest that these has been a phase of gibbsite precipitation that may have had a close but complex relationship with regards to the timing of the mineralisation processes, rather than formation as a result of late stage lateritic weathering.

Jul 11, 2011

By J Beniscelli

Some words to honour Pedro:Pedro was among the few persons with a very high level in geology modelling, resource and reserve estimation, geometallurgy development, sampling theory applications and a deep understanding of the mining industry needs. He had brilliant ideas about new developments on economic geology. He was always looking at the long-term and proposed challenging questions a decade in advance on their solution.The fundamental aim of geometallurgy is to add economic value to the mining of mine resources by timely providing information and knowledge about the relevant metallurgical variables represented by their space models. In this manner, geometallurgy allows obtaining the optimal design of a project, mitigating processing risks, optimising and supporting production plans and identifying and assessing mining business opportunities.This paper describes the conceptual framework and geometallurgical definitions and some of Codelco’s experience during the last 15 years.

Sep 5, 2011

By Leonhard Weixler, Jean Wehbe, Gordian Ulrich, Franz-Werner Gerressen

"As alternative energy industries pick up speed internationally, and as government regulation and restriction limit the use of impact hammers due to environmental limitation from construction induced noise during pile driving operations, unconventional solutions for more complex offshore foundation installations have become a necessity. Manufacturers and equipment suppliers race to develop alternative techniques to accommodate for offshore foundation construction in various soil conditions and at greater water depth. The FLYDRILL, and the Seabed-Drill BSD are two innovative systems for the installation of offshore monopiles with reduced environmental impact.INTRODUCTIONThe FLYDRILL and the Seabed-Drill were born from techniques in the design and construction experiences gained from land based foundation drilling systems over the past 25 years. As the alternative energy industry ventures into uncharted territory of deeper water for offshore wind farms, fast current regions for tidal energy, and increasingly challenging soil conditions and project complexity, the advancement and development of new innovative equipment system and drilling techniques have become a necessity.The systems described herein show the development of the so called “FLY-Drill” system, and the all new “Seabed-Drilling” equipment applied at the first ever made job using a seabed drilling rig.FLY DRILLThe FLY DRILL method is a completely new concept for bored piling. The drilling unit with rotary head, Kelly bar, and drilling tool is suspended by the hoist rope of the base carrier crane. The hydraulic power for the drill unit is provided from the base crane.The drill head is fixed on the casing with a hydraulically operated clamping system during the drilling operation. When emptying the drilling tool, the clamp cylinders are opened and the complete unit is swung sideways."

Jan 1, 2017

By L. S. Richardson, N. J. Grant

Iodide-grade titanium, two oxygen alloys, and two aluminum alloys were studied by means of creep-rupture tests from 1000° to 2000°F. From the test information an evaluation was made of, 1) the relative creep-rupture strengths of the u (hexagonal-close-packed) and 0 (body-centered-cubic) forms of titanium and some of its alloys; 2) the effects of solid-solution strengthening by oxygen and aluminum on the u and 0 structures; and 3) the extent and effect of atmospheric oxygen alloying on rupture life in tests above about 1000°F. THE use of titanium alloys at temperatures above 1000° F has been considered unlikely by most investigators because of low creep strength (based on a requirement for room-temperature ductility, formability, and weldability) and low oxidation resistance. The results of this testing program do not entirely uphold this pessimism in terms of strength and oxidation resistance. The observations described below are based entirely on stress-rupture results, although some creep tests were also carried out on these alloys. Three items of primary importance are covered: 1) Stress-rupture data in the form of the Larson-Miller parameter curves for several titanium alloys at temperatures up to 2000°F; 2) The relative strengths of the hexagonal-close-packed a structure and the body-centered-cubic 0 structure; 3) The effects of solid-solution alloying on the creep-rupture properties of titaniu'm at elevated temperatures. Only two investigatons have reported the creep- rupture properties of titanium at temperatures above 1000° F. Lunsford and rant' tested iodide titanium; their values are also utilized in the paper. EXPERIMENTAL PROCEDURE After a few tests had been made in a conventional creep-rupture test machine, it became clear that some sort of atmospheric protection would be necessary for short-time tests conducted at temperatures above 1600°F, or for times greater than 100 hr at 1200°F or above. Several furnace units and test machines were constructed, the most successful of which is shown in Fig. 1. Certain experimental difficulties peculiar to titanium were encountered; for example, the use of molybdenum holders was dictated because titanium forms low melting-point eutectics with conventional high-temperature alloys containing appreciable amounts of iron, nickel, or cobalt. This same difficulty made the use of chrome l-alumel thermocouples in direct contact with a test specimen impossible, above about 1500" F. The high-temperature high-vacuum furnace shown in Fig. 1 was capable of maintaining a vacuum of better than 10"5 mm of mercury at 2000°F (pressure measured in cold part of system) and 5 X lo-' mm at room temperature. Most of the tests eonducted in vacuum were at temperatures

Jan 1, 1960

By J. F. Radavich

Many heats of steel of low carbon value have been known to produce brittle pieces of steel. The brittleness is believed to be due to the impurities located within the grain boundaries. Such brittle steels have been examined with an optical microscope to ascertain the nature and the amount of the impurities present at the grain boundaries. Due to the relatively low resolving power of the optical microscope, the impurities are not visible in fine detail. The writer obtained some sheet steel and proceeded to determine the location of the impurities and to show the application of the electron microscope to the study of grain boundaries. One sample was known to be capable of becoming embrittled, whereas another sample was believed to be much less susceptible to embrittlement. Treatment of Specimens The specimens were embrittled by annealing above the A3 point under mildly oxidizing conditions. One piece of ingot iron could not withstand a 90" bend, whereas another piece of ingot iron was not affected and could withstand a 90" bend. The brittle piece was then annealed at a high temperature in a hydrogen atmosphere. The annealed ingot iron was termed cured and could withstand a 90" bend very easily. The three specimens examined will be designated as brittle, good. and cured in the discussion that follows. Procedure The sizes of the specimens were as follows: one piece of brittle ingot iron-3/8 by 35 in.; one piece of good ingot iron-96 by 1/8 in.; one piece of cured ingot iron-36 by 54 in. The specimens were too small to be polished by hand and therefore were mounted in bakelite. The polishing procedure was carried out in the conventional manner with the use of 1/0 through 3/0 papers, and the final polish was done with alumina on a billiard cloth. The specimens were then etched in a 4 pct solution of picral in alcohol, and then they were examined through an optical microscope. An area was chosen that showed distinct grain boundaries, and an effort was made to keep near this area when pulling the replicas REPLICA TECHNIQIJE The replica technique used in the preparation of the replicas for examination under the electron microscope is described in Electron Metallography.' It consists essentially of the following steps: 1. Obtaining a suitably etched specimen. 2. Applying a swab of ethylene di-chloride on the surface. 3. Applying a formvar solution on the surface. 4. Placing a screen on any desired spot. 5. Breathing on the fornivar layer. 6. Applying scotch tape on the screen and film. 7. Pulling the film and the screen up with the Scotch tape. 8. Separating the screen from the Scotch tape. This replica technique is very similar to the one described by Harker and Shaefer. However, with the added step, the percentage of replicas removed is very much higher regardless of the length of the time from the etching of the specimen to the actual pulling of the replica. The replicas were then shadow cast with manganese at a filament height to replica distance ratio of 1 1/2:7. This produced a very high contrast replica for use in the electron microscope. One of the dificulties encountered with this study was the restricted area of the specimen. The width of the specimens was the same as that of the 200 mesh nickel supporting screen. In order to increase the effective area, the screens were cut down as shown in Fig 1. The arrow indicates the direction in which the replica was pulled. This operation made it possible to obtain a large percentage of good replicas. Fig 3 shows an electron micrograph of a brittle piece of ingot iron and a grain boundary that was polished mechanically. The surface is very rough probably due to the incomplete removal of the flowed layer by the picral etchant. The grain boundary does show evidence of impurities. It was decided to electropolish the specimens to obtain a much smoother surface than the one obtained by mechanical polishing. ELECTROPOLISHING The specimens were cut in half to expose the metal on the back side. The exposed metal had sufficient area to make good electrical contact and electropolishing was carried out easily. The conditions for electropolishing were 0.9 amp, 35 volts, and 25 sec. in an electrolyte composed of 850 cc of ethyl alcohol, 100 cc distilled water, and 50 cc of perchloric acid. The polished specimens were then etched in the 4 pct picral solution for a shorter time than was necessary for

Jan 1, 1950

By Alla D. Mah

Heat content and entropy data at high temperatures were obtained for cuprous oxide (Cu2O) and cupric oxide (CuO). Experimental results and smooth values are reported. Equations are given for heat content increments and heat capacities above 2980 K. Low-temperature heat capacity data were obtained for cuprous oxide. Evaluation of the 2980 K entropy gave, S°298.15(Cu2O) = 22.08 ± 0.10 cal/deg mole. Heat of formation values were determined for cuprous and cupric oxides. The values are: AH°298.15(Cu2O) = -40,830 ± 300 cal/mole, and AH°298.15(CuO) = -37,230 ± 150 cal/mole.

Jan 1, 1967

Technological advances evolve in different ways; many are the result of years of application of scientific method. i.e. Observation Deduction Hypothesis Proof In the mining industry, better explosives, drill bits or reagents are likely to evolve in this way. The scientific method is suited to the application of conventional management structures. Other advances are simply the result of a "brainstorm", or building a better mousetrap. Such inspiration probably resulted in "Split Set" rock bolts, or the first use of a vacuum cleaner principle for shaft mucking. Individual consultants can apply lateral thinking to generate such inspiration. In gathering material for the presentation of this paper, I sent a.letter to all members of the Mineral Industry Consultants Association asking them to nominate problems in the industry requiring investigation. The answers have been divided into two classifications - contract research underway and identified problems requiring solutions.

Jan 1, 1986

By J. S. Pascover, J. K. Abraham

The influence of the early stages of precipitation on the kinetics and structure of martensite formation in Fe-27Ni and Fe-29.5Ni alloys containing from 0 to 10 pct Ti was examined with X-ray and electron microscopy techniques. The formation of a coherent, ordered preprecipitate had a profound stabilizing effect on the austenite. The Ms was decreased by increased titanium content and aging time up to a critical time. When the critical aging time was exceeded, the Ms was observed to increase markedly. The formation of the clusters was insuppressible and the volume fraction of clusters formed during the quench was a function of the titanium content. Martensite resulting from transformation of the clustered austenite is tetragonal with the c/a ratio increasing with titanium content. A model for the tetragonality is suggested. The morphology and substructure of the m artensite is inter-preted in terms of the above information and the cur-rent models of twinned martensite. ThE ramifications of precipitation in austenite to the properties of austenite have been the subject of numerous investigations. The current research is concerned with the influence of precipitation in austenite on the kinetics and structure of subsequent marten-site formation. In a previous investigation, Abraham et al.1 followed the aging reaction in an Fe-29.5Ni-4.2Ti* (at. pct) alloy using an X-ray diffraction technique. This technique, employing a Guinier camera, provided kinetic measurements through observation of the side band position as a function of aging time. The salient results of this work were: 1) The initiation of precipitation was not suppressed by quenching, i.e., there was a finite cluster zone size at zero aging time; and 2) The hardness of the aged austenite correlated extremely well with the zone size. During the previous work it was noted that the mar-tensite formed after aging was tetragonal, substanti-ating an earlier observation.2 Systematic investiga-tion revealed that the martensite was tetragonal in both the solution-treated then quenched, and the solu-tion-treated, aged, and quenched condition, and, furthermore, that a marked stabilization of the austenite occurred as a function of aging time. The present work is concerned with documenting the tetragonality and the stabilization phenomena as well as the ob- served microstruction with a suggested rationale for the behavior noted. EXPERIMENTAL PROCEDURE The compositions of the alloys are listed in Table I. The analyses were performed after the solution treatment of the strip material. Nickel was determined using the standard dimethy1-glyoxime procedure whereas titanium was determined colorimetrically with hydrogen peroxide and volumetrically by titrating with ferric iron. The materials were melted in a 5-lb vacuum induction furnace, cast into 2-in.-diam ingots, and forged in a temperature range of 950. to 1200°C to 1/2-in. slabs. The three higher titanium containing materials cracked during forging; therefore, to get the alloys into strip form, slices 1/8 in. thick were cut from the slab, homogenized 4 hr at 1150°C, then cold-rolled to a 0.04-in. thickness. The remaining slabs were hot-rolled, homogenized 4 hr at 1065"C, then cold-rolled to a final thickness of 0.03 in. All of the heat treatments were performed under a protective atmosphere of argon. The Ms temperature for most of the alloys is below room temperature; therefore, it was possible to solution treat, quench to room temperature, polish, and then observe the transformation optically on a cold-stage microscope. To determine the effect of aus-tenitizing temperature on Ms, eight of the alloys were treated at two temperatures, 1025" and 1120°C. No measurable variations in Ms were noticed. The remaining alloys were treated at 1025°C. The specimens, : by 5/16 by 0.03 in., were austeni-tized in a vertical tube furnace under a dynamic argon atmosphere. The bottom of the tube was submerged in water for quenching purposes. The question of stabilization that may be operating at room temperature was investigated and found to be negligible. Many of the specimens were held at room

Jan 1, 1970

By M. T. Simnad, G. Derge, I. George

Measurements of current efficiency on iron-silicate slags in iron crucibles showed that conduction is about 10 pct ionic in slags with less than 10 pct silica and about 90 pct ionic in slags with more than 34 pct silica, increasing linearly in the intermediate range. The balance of the conduction is electronic in character. Silicate ions are discharged at the anode with the evolution of gaseous oxygen. Transport experiments show that the ionic current is carried almost entirely by ferrous ions, which may be assigned a transport number of one. THERE has been increased evidence in recent years that the constitution of liquid-oxide systems (slags) is ionic.1-3 The principal studies designed to establish the structure of liquid slags have been by electrochemical methods', " and conductivity measurements1,6,7 which also have indicated the presence of semiconduction in several silicate systems1,4-0 and in pure iron oxide.' It is well known that many slag-forming metallic oxides have an ionic lattice type in the solid state, and their properties are determined to a large extent by the lattice defects and ion sizes. As Richardson8 as pointed out, the detailed models of liquid slags cannot be found on thermodynamic data only but "must rest on a proper foundation of compatible structural and thermodynamic knowledge, combined by statistical mechanics." A careful thermodynamic study of the iron-silicate slags has been carried out by Schuhmann with Ensio9 and with Michal.10 They obtained experimental data relating equilibrium CO2: CO ratios to slag composition and made thermodynamic calculations of the activities of FeO and SiO, and of the partial molal heats of solution of FeO and SiO2 in the slags. It was found that the activity-composition relationships deviate considerably from those to be expected from an ideal binary solution of FeO and SiO2. However, the partial molal heat of solution of FeO into the slags was estimated to be zero. Their experimental results were correlated with the constitution diagram for FeO-SiO2 of Bowen and Schairer,11 with the results of Darken and Gurry" on the Fe-O system, and with the work of Darken"' on the Fe-Si-O system. All these studies were found to be consistent with one another. The variation of the mechanism of conduction with composition in the liquid iron-oxide-silica system in the range from pure iron oxide to silica saturation (42 pct SiO2) in iron crucibles was reported in a preliminary note." The current efficiency, or conformance to Faraday's law, showed some ionic conductance at all compositions, the proportion increasing with the concentration of silica. The current-efficiency experiments since have been extended. Furthermore, transport-number measurements have been completed in silica-saturated iron silicates to determine the nature of the conducting ions. Experimental Current Efficiency in Liquid Iron Oxide and Iron Silicates using Iron Anodes: This study was carried out by passing direct current through slags in the range from pure iron oxide to iron oxide saturated with silica (42 pct silica), using pure iron rods as anodes and the iron container as the cathode. A copper coulometer was included in the circuit to indicate the quantity of current passed during electrolysis. Assuming that the cation involved is Fe-+, the theoretical quantity of iron lost from the anode according to Faraday's law may be calculated and when compared with the actual loss observed, gives an indication of the extent to which Faraday's law has been obeyed. It also gives an indication of the presence and extent of ionic conduction in the melt. Preparation of the Slags: About 100 g of chemically pure Fe,O, powder is placed in an iron pot which is heated by induction until the contents liquefy. In this way, FeO is produced according to the reaction Fe2O3 + Fe = 3 FeO. More Fe2O3 or SiO, powder is added and, when a sufficient quantity of molten slag is obtained, the induction unit is turned off, the pot withdrawn, and the molten slag poured on to an iron plate. Homogenization and Electrolysis of the Slag: Apparatus—After considerable development, the setup illustrated in Fig. 1 proved to be quite satisfactory. A is an Armco iron cylinder, 1 in. ID and 1/8 in. wall, consisting of three sections placed one on top of the other. The bottom section is a pot about 5 in. long with a small hole drilled in its bottom to allow withdrawal of gases during evacuation of the apparatus. The middle section is 6 in. long and consists of a pot which serves as the slag container, while the top section is a hollow-cylinder continuation of the slag-container pot. The height of this latter section is about 5 in., giving an overall length of approximately 16 in. The iron cylinder is constructed in this way for ease of fabrication, the individual sections becoming welded together after the

Jan 1, 1955

By A. J. Kroha, N. Wesis

Most copper mines produce both ore and low-grade "leach" rock or acid waters that contain recoverable copper. The sulfide ores pre¬dominate, and a portion that is too low grade for milling to produce concentrates for smelting, but has to be mined and trucked away anyhow, may be leached successfully with acid in dumps. Most of this leach material consists of sulfides and silicates or carbonates, and if the gangue is such that it consumes a high quantity of acid, this factor may rule out a leach operation. There are also valuable deposits that contain mostly acid-soluble copper, or occasional sulfide ores from which a sulfide concentrate can be roasted and acid-leached to produce a copper-bearing solution. Finally, there are milling ores in which the lesser part of the copper is acid-soluble and can be precipitated with iron or synthetic inorganic precipitants that produce metallic copper or copper sulfides that will float with the sulfides. Ordinarily, ores that contain copper associated with the sulfur ion, such as in the minerals chalcopyrite, chalcocite, bornite (and others), are milled to produce a 25-30% Cu concentrate for smelting, while a lesser amount of acid-soluble copper may be converted from solution to cement copper on iron scrap. A fast-growing percentage of such copper, however, is removed from solution with exchange resins or organic compounds in organic carriers such as kerosene (solvent extraction), then eluted with strong acid and subjected to electrolytic precipitation either in marketable form or as anodes that can be refined further. From the point of view of conventional copper smelting, copper flotation concentrates and cement copper are of chief interest in this chapter. Table I is a condensed open schedule for concentrates that generally run between 25 and 35% copper, and much less frequently as low as 12-15% or as high as 65-75% copper, the former being due to intimate relationship with pyrite (like the former United Verde Extension), and the latter representing such ores as the Bolivian Coro¬coro ore in which the copper is in the form of chalcocite in sandstone. These extremes are no longer common. When they occur, a special purchase schedule has to be negotiated. Included in Table 1, copper precipitates (cement copper) generally run from 70-85%a copper, and the same basic purchase schedule is used as with flotation concentrates. Sulfide Flotation Concentrates The sulfide copper concentrate produced in the mill as a flotation froth, with water then added for transportation of the heavy mineral particles from the flotation cells to thickeners, may run 60-80% water by weight, and the removal of water down to 25-50% by weight by means of thickeners, followed by further dewatering by continuous vacuum filters to 7-18% moisture by weight (depending on size of solids by screen analysis and also by content of clay) is a critical operation. Mill operators would like to produce a filter cake with 7-9% moisture content, but even with the help of steam on the filter this desirable condition is seldom realized when the concentrate is as fine as 80% -325 mesh. More commonly, the final concentrate is reground in pro- to produce best copper recovery and grade of concentrate (or molybdenite separation). In those cases, increasingly frequent, the filter product may not be a cake at all, but a mud that is hard to handle-even requiring a thermal dryer. Greater difficulty of form¬ing a manageable cake often comes from the copper-molybdenum separation by flotation, because the alkaline sulfides and hydrosulfides, or cyanide, or other similar reagents used for the separation, may leave the now relatively molybdenite-free copper concentrate even more difficult to filter. Handling a wet filter cake is difficult enough when its destination is only a short distance away-a matter of yards rather than miles. In those cases the filter cake may be thermally dried near the point of production, using rotary or multiple hearth, or fluidized-bed dryers. Alternatively, the concentrate may be pumped or carried in slurry form to the smelter and filtered there, or it may be spray-dried and compacted. For transportation to a smelter just a few miles to a few thousand miles away by ship or railroad other factors may be important, such as: in shipping by sea, avoidance of spontaneous combustion; in shipping by rail, losses by leakage if too wet or by wind and sun if too dry. It is the responsibility of the millman-usually the mill superinten¬dent-to make sure that his concentrates are in satisfactory condition when they leave the mill so that they meet these requirements: 1) They must have been accurately sampled and dry-weighed, the latter meaning that a moisture determination and gross weight must have been taken. 2) They must be dried sufficiently when necessary to prepare them for safe transportation. 3) They must arrive at the smelter with reasonable likelihood that they can be check-weighed and sampled fairly and equitably,

Jan 1, 1985

By Chang-Dong Li

Cadmium ingot and nickel powder were prepared by high-temperature distillation, solvent extraction and liquid phase reducing method with spent Cd-Ni batteries as raw material. The impacts of the temperature in the distilling stage, the concentration of H2SO4 in the leaching stage, the pH value in the extracting stage and the dosage of hydrazine in the reducing stage were discussed respectively. The as- prepared samples were characterized by the means of X-ray diffraction (XRD), scanning electron microscopy (SEM), atomic absorption spectrometry (AAS) and inductively coupled plasma-atomic emission spectrometry (ICP-AES). The results showed that the recovering efficiencies of Cd and Ni reach 99.5% and 99.3%, and the purities of Cd ingot and nickel powder were higher than 99.90% and 99.80%, respectively.

Jan 1, 2009

By M. A. Goudie

The deposits occur in a large salt basin of Middle Devonian age. The potash, the final deposit in the salt basin, results from several interrupted cycles of evaporation and dessication. The deposits are extensive, and, at first glance, relatively undisturbed. With more and more wells being drilled, it has now become evident that salt solution has played a large part in changing the original deposits, resulting in some cases in partial to complete removal of the potash and the underlying halite. The most dominant factor in the removal of salt by solution appears to have been tectonic movement and consequent faulting, probably of relatively minor dimensions but of major importance. Evidence which indicates the tilting of the evaporite basin to the north and northwest is shown by the changing pattern of the basin during succeeding eras of potash deposition. The potash minerals of most importance economically are sylvite and carnallite. Reserve calculations indicate that 6.4 billion tons of recoverable high grade potash in K2O equivalent exist in the basin. The Devonian salt basin, which contains the Saskatchewan potash deposits, extends from just east of the foothills in Alberta, north as far as the Peace River area, across Saskatchewan and into Manitoba as far east as Range 10 west of the First Meridian and south into Montana and North Dakota (Fig. 1). The basin is closed everywhere except to the northwest. The known potash deposits are confined almost entirely to the Province of Saskatchewan, with the exception of a small area in western Manitoba bordering the Saskatchewan boundary. The following discussion will concern only the Saskatchewan part of the basin. The evaporite series in the basin is defined as the Prairie Evaporite Formation of the Elk Point Group, of Middle Devonian age. Recent work done by potassium-argon dating methods has indicated an Upper Middle Devonian (Givetian) age of from 285 to 347 million years for the potash. The Elk Point Group consists in ascending order of the Ashern, Winnipegosis, and Prairie Evaporite Formations. The Ashern formation, with an average thickness of 30 ft, sometimes called the Third Red Bed, consists of dolomitic shales and shaly dolomites. The Winnipegosis, is a reef-type dolomite, usually with good porosity, and in many cases oil-staining, although to date no production has been obtained. The thickness varies from 50 to 250 ft. The Prairie Evaporite formation, varying from 0 to 600 ft in thickness, consists of halite with interbedded anhydrite and shale, with considerable amounts of potassium salts in the upper part of the formation. The potassium salts are chiefly chlorides, although very minor occurrences of sulfates have been re- ported. The anhydrite beds do not appear to be continuous, although generally one or two bands of anhydrite underlie the lowest potash zone and are used as marker horizons. The shale occurs as seams interbedded with the salts, as large irregular inclusions in the salts and as very fine particles in intimate mixture with the salts. The Prairie Evaporite Formation is overlain by the Second Red Bed member, the Dawson Bay Formation and the First Red Bed Member of the Manitoba Group, listed in ascending order. The Red Beds are shales which vary in color from red to green, maroon, grey, grey-black, and reddish purples. They serve as marker horizons for coring the potash. The Second Red Bed averages 14 ft in thickness, the First Red Bed 35 ft. The Dawson Bay Formation, which everywhere overlies the First Red Bed and the Prairie Evaporite Formation in the area under discussion, is a reef type of carbonate, in some places limestone, in others limestone and dolomite, with vugular to pinpoint porosity averaging 130 ft in thickness. In some parts of the area, it has a salt section near the top of the formation, usually with interbedded shales and limestones. In other parts of the area, it is waterbearing and the salt is absent. Detailed mapping has indicated that the areas in which the Dawson Bay is water-bearing are areas which have been disturbed by faulting. Where the Dawson Bay is salt-bearing, the porosity has been plugged by salt. The total thickness of the salt varies from between 600 to 700 ft in the center of the basin to zero at the northern edge of the basin (Fig. 2).* The salt-free area in the center of the Province is believed to have resulted from removal of salt by solution. Evidence from several wells suggests that salt removal has been a continuing process from the time of deposition to the present day. One well drilled between the Quill Lakes for potash information encountered

Jan 1, 1961

The process design of a metallurgical plant follows logical steps, one built on another, to define the circuit requirements. The methodology used and documentation produced is the same for a pre-feasibility study, feasibility study or a complete design, regardless of the process complexity or the commodity. The only difference is the amount of detail developed at each stage. SRK Consulting, an independent, international consulting practice that provides focused advice and solutions to clients, mainly from earth and water resource industries, has consulted on many mining projects around the world. The company was formed in 1974 and now employs more than 1,600 professionals in more than 50 offices on six continents. According to SRK experts, the first element in plant design is collecting metallurgical testwork or operating data from an existing plant. The testwork identifies the process, the metallurgical recoveries and product qualities and establishes the process parameters for economic recovery of the commodity. Test samples must represent both the whole orebody and the feed over the life of the mine, showing the mineral?s variability. With this data available, the process design can be developed.

May 1, 2013

By T. L. Joseph

S. T. Killian (Johnstown, Pa.)—This is one of the finest papers I have read. Tying in stoichiometric calculations with furnace performance and practice is a step which had to be taken sooner or later. The noteworthy difference between Dr. Joseph's type of calculations and regular blast furnace calculations is that the Ib mol system is used as a basis. With the Ib mol system, weights, volumes, and chemical reactions can all be expressed in the same equation. In the paper, wind, ore, flux, and fuel are all expressed as lb or lb mols. Probably Dr. Joseph does not realize it, but the vague word coke appears only twice in the entire paper. Lb of C and Ib mols of C are followed through reactions but the word coke appears only as 1560 lb coke per ton of pig and 1700 lb of coke per ton of iron. Obviously in order to understand furnace reactions, the coke should be expressed as lb or lb mols of C. Furnaces can also be compared more easily. In some respects the paper is too thorough and too complete. The effects of the metalloids reduced into the iron upon the top gases represents a difference of less than 3 pct of the CO formed in the bosh. Due to the completeness of the calculations in relation to the CO/CO2 ratio, this was included and was necessary. However, the exclusion might have enabled more furnace men to follow the lb mol system of calculations through the blast furnace to a better degree. By considering only irons of similar analysis, this part of the calculations might have been omitted. However, if this had been done, total rewriting of the paper later would have been necessary in order to make the work complete as it is now. It also would not have been nearly as authoritative. In the paper, there appears the reaction: H2O + CO ? CO2 + H2 [I] Dr. Joseph states that he did not take this reaction into consideration in any calculations pertaining directly to the paper. The piobable reason- for this is that although it contains all the main reacting top gases except N2, it is rather inflexible since it is monomolecular in relation to each of the reacting gases and does not tie in with the gasification of C. Actually the reactions: H2O + C ? CO + H2 [2] 2H2O + C ? CO2 + H2 [31 and the solution loss reaction: CO2 + c ? 2c0 [41 tend to assume an equilibrium through the reaction: H2O + CO ? CO2 + H2 [I] which should be considered a balancing or equilibrium reaction. Reactions 2, 3 and 4 permit furnace conditions to balance with the CO/CO2 ratio and H2 formation. They tie in with the solution loss. Reaction 1 unites them chemically. Probably the best calculation to make at this time would be to try to find the relative importance of the CO2 from the flux and the H2 in the dilution on an actual furnace gas analysis. For this purpose the Dob-scha-Carnegie-Illinois paper—"Effect of Sized and Sintered Mesabi Iron Ores On Blast Furnace Performance" is chosen. This paper was presented before the blast furnace section of the AIME in 1948. This represents the best large scale furnace operation available to me. Unfortunately the changes were brought about by beneficiation of the burden and not by changes on one burden. In choosing the basis for the calculation in relation to the furnace, 100 mols of dry top sgas is chosen. This leaves something to be desired inasmuch as the nitrogen basis is changing but I believe it will be better understood than any other type and it is the easiest to use.

Jan 1, 1952