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By Samuel L. Hoyt

Cemented tungsten carbide, a product of a branch of metallurgy which has never possessed more than a relatively minor interest and importance, has recently commanded the attention of engineers, industrialists, financiers and men of business. New methods, new opportunities, new goals, never before considered, have suddenly been brought within the scope of practical realization. It is these results that attract the interest of the world, but we, as engineers and metallurgists, will have equal interest in the product itself and in that branch of metallurgy from which it comes. It seems appropriate, therefore, to speak today on Hard Metal Carbides and Cemented Tungsten Carbide, as the latter represents one of the outstanding developments of recent date in non-ferrous metallurgy. We are interested in the hard metal carbides as a group of related metallic substances and in cemented tungsten carbide as the leading representative of the metallurgy of this group. The literature of this general field is limited, and, bearing this in mind, it seems advisable to attempt to cover the field broadly and to correlate the subject matter, rather than to present a detailed account of any one of its phases. Our annual lecture lends itself admirably to this plan. The quickened interest in cemented tungsten carbide, once its potentialities were recognized, the many sides to the problem of developing the manufacture and use of this new product, and the desire to economize as much as possible in the time of development, have brought many into this work, both in this country and abroad, notably the Osram Lamp Co. and Friedrich Krupp in Germany and the General Electric Co. in this country. It would be presuming too much to attempt to speak too broadly of the present state of the art and science of the hard metal carbides and my remarks arc guided largely by the work done in the General Electric Co., and by our associates in the Carboloy Co. Even here the activities of many in these two groups have centered around the utilization of cemented tungsten carbide, which is outside the scope of this talk. To those whose work is represented in this lecture, it is my pleasant duty to acknowledge their assistance and advice. To Dr. Zay

Jan 1, 1930

By James H. Courtright, Kenyon Richard

TOQUEPALA is a porphyry copper deposit in which mineralization is localized by a large breccia pipe formed in close genetic relation to intrusive rocks. The deposit is in southern Peru, 55 airline miles north of the small city of Tacna and the same distance inland from the port of 110. Quellaveco and Cuajone, geologically similar deposits, lie 12 and 19 miles north of Toquepala. Chuquicamata is 400 miles to the south. The deposit is high on the southwestern slope about 20 miles from the crest of the Cordillera Occidental of the Andes Chain. It lies in a mountainous desert where the steep southwesterly slope of the Andes is dissected by a succession of rapidly downcutting, deep canyons. Local topography is moderately rugged with a dendritic drainage pattern and an elevation of 8000 to 14,000 ft. Volcanic peaks along the crest of the Cordillera rise over 19,000 ft. Local precipitation, including a little snow, amounts to about 10 in. during January and February, but general runoff in the region is slight. Throughout southern Peru the springs and streams are widely separated. Crude canals irrigate small farms on terraced slopes along the streams and provide sparse subsistence to the semi-nomadic inhabitants. During the past decade, engineering and geological explorations of the region, as well as the mineral deposits themselves, have required construction of a network of several hundred miles of roads. Before this, roads extended only a few miles inland. Many areas still can be reached only by trail. Toquepala was briefly described in 19th century geographical literature as a copper deposit, and it received desultory attention from Chilean prospectors early in the present century. It was first recognized as a mineralized zone of possible real importance by geologist O.C. Schmedeman during an exploration trip for Cerro de Paso Copper Corp. in 1937. The discovery was late as compared to earlier recognition of Chuquicamata, Potrerillos, and Braden of Chile and Cerro Verde of southern Peru. This was due partly to the region's difficult accessibility but principally to the obscure character of the outcrop evidence of copper. From 1938 until 1942 Cerro de Pasco Copper Corp. partially explored the deposit by adits and diamond drillholes. This campaign was supplied by a 60-mule pack train continuously shuttling over a 30-mile trail. Northern Peru Mining & Smelting Co., a wholly owned subsidiary of American Smelting & Refining Co., undertook regional engineering stud- ies in 1945 and drill exploration in 1949. According to published data1 the deposit contains 400 million tons of open pit ore averaging a little over 1 pct Cu. It is currently undergoing large-scale development by Southern Peru Copper Corp., which is owned by American Smelting & Refining, Phelps Dodge, Cerro de Pasco, and Newmont Mining. Summary of Geology: The deposit is situated in a terrane composed of Mesozoic(?) and Tertiary volcanic rocks intruded by dioritic apophyses of the Andean Batholith. These formations are exposed in a northwesterly trending belt about 15 miles wide. Along the northeast they are unconformably overlain by Plio-Pleistocene pyroclastic rocks, which occupy much of the crest of the Andes, and along the southwest they are covered by the Moquegua formation of Pliocene(?) age. The mineralized area, oblong in shape and about 2 miles long, has been a locus of intense igneous activity. Several small intrusive bodies having irregular forms occur within and adjacent to a centrally located, large breccia pipe. The mushroom-shaped orebody consists of a flat-lying enriched zone of predominant chalcocite with a stem-like extension of hypogene chalcopyrite ore in depth within and around the pipe. This breccia pipe is relatively large and has been formed by repeated episodes of brecciation. Small satellitic pipes occur at random within a 2-mile radius of this central pipe. These too were individual sourceways of mineralization, although not always of ore grade. Within and around the zone of breccia pipes and mineralization there are a few faults and veins, but these are discontinuous random structures of minor significance. There are no regional or local systems of faults or other planar structures recognized which could account either for the mechanical development of the breccia pipes or for their localization as a group or as individuals. Hydrothermal alteration is pervasive in the zone of mineralization. Clay minerals appear to be abundant in places, but their percentages are undetermined. Quartz and sericite are the principal alteration products, and in many instances original rock textures are obliterated. The principal sulfides, hypogene pyrite and chalcopyrite and supergene chalcocite, occur mainly as vug fillings in the breccia and as small discrete grains scattered through all the altered rocks. Sulfide veinlets are relatively scarce. Sulfides are more abundant and alteration is more intense in certain rock units, such as the diorite and most of the breccias. Although the Toquepala mineral deposit is similar in most respects to the porphyry copper deposits of southwestern U. S., it most closely resembles the Braden deposit of Chile, as described by Lindgren

Jan 1, 1959

By R. L. Jones, H. Conrad

The effects of gain size on the room-temperature tensile properties of commercial(A- 70)-purity and iodide-purity titanium were investigated by concurrent tensile testing and thin-foil electron transmission studies. The mechanical tests showed that the effect of grain size on the yield, flow, and fracture stresses could be expressed by relations of the Hall-Petch type. The magnitude of the grain size effect on the yield stress was larger for commercial-purity than for high-purity titanium, but in both materials it was intermediate between that typically observed for fcc and for bcc metals. The effect of strain on the gain size sensitivity of the flow stress was found to reflect the grain size dependence of strain hardening. The electron transmission study indicated that the tensile behavior could be accounted for in terms of the dislocation densities and distributions formed during deformation. It was concluded that the effect of grain size on strength could be attributed to a relationship between grain size and the dislocation densities generated in initiating and sustaining plastic flow, the difference in stress levels being due to strain hardening. The relationships between recrystallized grain size, uniaxial tensile properties, and dislocation substructure have not been explored in any detail for a titanium. With titanium and its alloys finding increasing applications in aerospace systems, such information has become essential. The research described in this paper had two objectives: first, to document the effect of grain size on the yield, flow, and fracture stresses of unalloyed a titanium by the tensile testing of recrystalliied specimens at room temperature; second, to attempt an interpretation of the mechanical test behavior in terms of the dislocation structures formed on deformation, as observed by transmission electron microscopy of thin foils prepared from deformed tensile specimens. The observations reported here are restricted to room-temperature deformation; the effect of temperature on the yield and flow stresses of a titanium and the athermal nature of the contribution of grain size to the flow stress have been discussed elsewhere.1"3 EXPERIMENTAL DETAILS Two purity grades of unalloyed titanium have been investigated: Crucible Steel A-70, a typical commer- cial-purity product, and a high-purity iodide grade supplied by the Foote Mineral Co. Chemical analyses for interstitial impurity contents are given in Table I, together with values for the total interstitial atom fractions in terms of oxygen equivalents, which were calculated by the method outlined previously.4 ,5 According to the manufacturer's nominal analysis, the A-70 grade also contains about 3700 ppm of Fe. The A-70 grade was supplied as 1/4-in.-diam centerless-ground rod. The iodide grade was supplied as crystal bar which was consolidated to approximately 1/4-in.-diam rod by a single, rapid, molten zone pass, using electron beam heating. The 1/4-in. rods were cold-swaged to 0.078-in.-diam wires and 2-in. lengths were recrystallized in a vacuum of -5 x 10-5 torr, giving grain sizes in the range 0.8 to 30 . The thermal treatments and resulting grain sizes are summarized in Table 11. Since there was in no case any obvious sign of a bimodal dis tribution of grain sizes, the grain size values quoted in Table I1 are those obtained from mean linear intercept measurements involving >I00 grains. For the two coarse grain sizes the measurements were made using conventional optical techniques. It was noted during these determinations that the etchants used (10 pct HF; 30 pct HN03; 60 pct H20 and 2 pct HF; 2 pct H202; 96 pct H20 were both tried) etched grain boundaries very nonuniformly and optical polarization effects had to be employed to reveal boundaries which might otherwise have been missed. It was found that this problem became more acute with the finer grain sizes as the usefulness of polarization diminished.

Jan 1, 1970

By X. Xiang, W. Xia, J. Yin

"TiCl4 slurry containing valuable metals is an unavoidable by-product of the titanium ore chlorination process. The recovery of these valuable metals, which include titanium, niobium, tantalum, and aluminum, is an urgent issue to tackle in the titanium industry. The results of this investigation show that the valuable metallic elements can be recovered from the slurry by evaporation in a sealed container and leaching with dilute hydrochloric acid. After evaporation at 200°C for 60 minutes, nearly 99% of the titanium was recovered in the form of TiCl4, which was formed by the reaction of the TiO2 in the slurry with AlCl3. After evaporation, metals like niobium, aluminum, and tantalum remained in the residue. By leaching with 2.1 mol/L HCl at a L/S ratio of 6:1 mL/g at 80°C for 60 minutes, the soluble metals, such as aluminum, iron, and copper were all removed from the residue, and the niobium and tantalum were further enriched in the leach residue. A concentrate containing 53.40 wt% Nb and 5.57 wt% Ta was obtained by washing the leach residue with dilute aqueous ammonia under stirring. A potential waste water purifying agent containing 263.75 g/L AlCl3 was produced by purifying the leaching solution with Al(OH)3 and modified polyacrylamide. IntrodutionTitanium tetrachloride (TiCl4) is an important intermediate in titanium metallurgy that is widely used in the production of titania and titanium sponge (Akhtar, Xiong, and Pratsinis, 1991). In conventional titanium metallurgical processes, high-titanium materials such as synthetic rutile and high-titanium slag are reacted with chlorine to produce TiCl4 in a fluidized-bed furnace. Accompanying elements, such as Al, Nb, and Fe also react with chlorine to form chlorides, which are mixed with gaseous TiCl4 (Anderson, 1917). In the condensation process, gaseous TiCl4 is refrigerated and high-boiling-point chlorides are precipitated in the liquid TiCl4 to form a slurry containing 50~60 wt% TiCl4 (Wang, Xiang, and Wang, 2012).Various methods have been proposed to recover TiCl4 from the slurry, including microwave heating and spray drying (Wang, Xiang, and Wang, 2012, Wang et al., 2010a). To date, the only way to recover the TiCl4 is to return the slurry to the fluidized-bed furnace. However, this causes fluctuations in the temperature in the furnace, resulting in incomplete reaction between titanium-rich materials and chlorine. Therefore, the slurry is often rinsed with water and neutralized by adding lime, which caused serious environmental pollution and wastes a valuable resource (Roy, Bhatt, and Rajagopal, 2003)."

May 1, 2019

By Robert Mensah-Biney

The loading characteristics of gold-tetrabromide species onto a weak-base resin, PAZ-4, were evaluated using a completely mixed batch reaction system. The time. to attain equilibrium loading was substantial, on the order of several days. However, abnormally high loading capacity values of 15 to 20 meq Au/g dry resin (1000 to 1300 kgAu/tonne) were attainable in 24 hours. The loading of gold-bromo species by P AZ-4 was visualized as consisting of two processes: (i) the rapid adsorption of [AuBr4r species by simple ion exchange mechanisms and (ii) reduction of [AuBr4r to metallic gold by an electron exchange mechanism. Elution of gold from loaded resin was feasible provided the resin had not been loaded to high capacity values. If loading is carried to high levels, complete elution becomes impractical and the gold may be recovered by ashing the loaded resin. For the high loading capacities observed this option would appear to be cost effective.

Jan 1, 1991

By G. Y. Liu

The structure-activity relationships for four thionocarbamate collectors are discussed in this pa-per. These collector classes are alkyl alkyl thionocarbamates, allyl alkyl thionocarbamates, acyl al-kyl thionocarbamates and alkoxylcarbonyl alkyl thionocarbamates. The main emphasis will be to highlight the differences between collectors and their affinities for copper and iron sulfide minerals on the basis of data on electronic-structure properties obtained by vacuum and solution density functional calculations, flotation experiments and analysis measurements. The mechanism of interaction of these collectors with sulfide minerals is also discussed with perturbation theory. The theoretical study results indicate that acyl alkyl thionocarbamates and alkoxylcarbonyl alkyl thionocarbamates have best performance in these agents as flotation collectors with improved collecting power for copper sulfide minerals and excellent selectivity against iron sulfide minerals, which are consistent with the experimental data reported in the literature.

Jan 1, 2014

By J. W. Merks

"Simulation is the process of designing a computerized model, of a system or procedure, for the purpose of understanding its behaviour and developing strategies to control its operation. Simulation models are rapidly becoming an effective tool for mineral processing plants. Powerful software and fast hardware are more readily available than a basic understanding of the theory of mathematical probability and applied statistics that process simulation models demand.Simple spreadsheet software can be used to develop simulation models that take into account any set of conditions at a mineral processing plant. In addition to an unbiased estimate for each variable in the set on which two-and three-product formulas are based, reliable variance estimates for all variables are required. How to estimate variances, effectively and at the lowest possible cost, and how to compute variances for simple functions such as the metal content of a quantity of ore or concentrate, or more complex functions such as the percentage recovery at a mineral processing plant, are the key to useful simulation models.IntroductionSimulation models for systems and procedures have become one of the most powerful and effective applications for computers( 2.6). Simulations are applied to a wide range of scientific and engineering disciplines. Generally, a model is designed to investigate the effects of changes in and interactions between a set of variables before a system is in operation or a procedure is implemented, by simulating weeks, months, and even years of operation, in a few minutes of computer time. Along with practical and useful applications always comeabuse and misuse. For a simulation process cannot be more reliable than the methodology that was applied to design the model, to estimate variables and variances, to check performance, and to verify validity. A mineral processing plant is a dynamic system that changes continuously as a function of time but within probabilistic constraints that can be estimated by applying statistical tools and techniques to the variables and variances that interact in the system.A model should be designed to simulate not only the set of variables but also the variance for each variable. Models in which constant variances are used to generate normally distributed random numbers, are less realistic than models in which the variances are simulated first, and then used to simulate the variables."

Jan 1, 1991

By X Dai

This paper investigated the loss of cyanide via oxidation in CIP or CIL processes by relating the cyanide oxidation to its major oxidation product, cyanate. The results showed that the oxidation of both free cyanide and copper cyanide complexes are substantially enhanced by the presence of carbon. The rate of cyanide oxidation was found to occur in the order: Cu(CN)32- > CN- > Cu(CN)2-. The cyanide oxidation rate in the presence of copper and carbon increased significantly with increasing dissolved oxygen concentration, typical of CIL vs. CIP conditions. Cyanide oxidation rate of iron, zinc or nickel cyanide complexes was minimal.

Jan 1, 2011

By Leonard B. Gulbransen, John R. Lewis, W. Martin Fassell, J. Hugh Hamilton

T. E. Leontis (The Dow Chemical Co., Midland, Mich.)—This paper is of particular interest to me because of my own work with F. N. Rhines on the oxidation of magnesium and magnesium alloys a few years ago. The authors are to be complimented on their development of an accurate and reproducible technique for measuring ignition temperatures and on their comprehensive study of the many variables that affect the ignition temperature of magnesium. It is indeed gratifying to see that they have obtained a good correlation between ignition temperatures and the oxidation rates reported by us. The correlation is valid not only with composition within one alloy system but also between alloy systems; that is, alloying elements which effect the greatest increase in oxidation rate also produce the greatest decrease in ignition temperature. There are a few points upon which I would like to comment. In attempting to correlate ignition-temperature data, one must be sure that the same definition of this quantity is used by all investigators. It does not appear to me that such is the case in the authors' comparison of their data with the theoretically calculated values of Eyring and Zwolinski. The equation derived by these investigators defines the ignition temperature, To, as the temperature at the gadoxide interface, whereas the present authors use the metal temperature as the criterion for ignition. The contradiction in the effect of oxide-scale thickness on ignition temperature between the predictions of the Eyring-Zwolinski equation and the observations reported in this paper indicate that some variable has not been taken into consideration. Could that be the geometry and size of the specimen? There is a marked difference in the type of specimen used in this investigation and that used in our work which formed the basis of Eyring and Zwol-inski's theoretical treatment. Another factor which plays an important role in ignition is the vapor pressure or the rate of vaporization. Combustion can safely be assumed to take place in the vapor phase by the reaction between vaporized magnesium and oxygen. Thus, a more accurate theoretical analysis may be made on the basis of the rate of vaporization which may be the controlling rate of the process. The effect of a large number of alloying elements on the ignition temperature has been reported in this paper, but beryllium was not included. Practical experience dictates that beryllium markedly decreases the burning tendency of magnesium. I was wondering if the authors plan to study the effect of beryllium in their future work. The authors predict that concentrations of sulphur dioxide in the furnace atmosphere greater than 5.8 pct would be expected to increase the ignition temperature to values still higher than those they measured. I would like to mention that large concentrations of sulphur dioxide markedly increase the rate of combustion of magnesium once ignition has started. Although it has been shown in the paper that the ignition temperature of magnesium in oxygen increases with increasing sulphur dioxide content up to about 1 to 2 pct, in practice relatively low-melting commercial cast alloys (AZ63A and AZ92A) are being continuously heat treated at temperatures just below the melting point in air containing 0.5 to 0.75 pct SO*. In regard to the change in color of the oxide scale observed on magnesium and magnesium alloys just prior to ignition, I would like to mention that in our work alloying elements were found to color the usually white magnesium oxide even though ignition did not occur. For example, the oxide formed on Mg-A1 alloys was gray, increasing in intensity with aluminum content in the alloy. Finally, I might suggest that the authors indicate their source of the value of 0.8 g per cc for the density of MgO as it is formed on magnesium upon oxidation at elevated temperatures. W. M. Fassell, Jr. (authors' reply)—The comments by Dr. Leontis are very excellent ones and I will attempt to answer them in order. First, the problem of ignition of magnesium is a rather difficult one since many factors are involved. Concerning the comparison of the To in the Eyring-Zwolinski equation, eq 4, with the experimentally determined values, it will be noted that the calculated and experimental values of the ignition temperature in Table I are not self-consistent. In the case of the 1.78 pct A1-Mg alloy the calculated value is 49°C below the experimental value; for the 3.81 pct A1-Mg alloy, 122°C below the experimental value; for Mg with 5x10-' cm film, 19°C above the experimental value; for Mg with 2x10-I cm film, 28 °C below the experimental value. Thus, if it were merely a matter of difference of location of temperature measurement the calculated ignition temperature would always be below the experimental value, the difference being due to the thermal gradient through the oxide film. The possibility of a thermal gradient in the magnesium metal must be considered. From Carslaw and Jaeger,'Y t can be shown that the maximum temperature gradient that could exist between the oxide-metal interface and the center of the sample is of the order of O.Ol°C. The geometry and size of the specimen could certainly have some effect on the ignition temperature. The equation for ignition that has been proposed in reference 14 is of the following type containing terms to account for this and other factors: M dT AHv(T) =Cp--------------\-J(.T—TB) + ZAHl-M A dx where AH is the heat of reaction, v(T) is the velocity of the reaction at temperature, Cp is the heat capacity of sample, M is the mass of sample, A is the area of sample, t is time, J is the total coefficient of heat transfer outward from the reaction zone, TR is the temperature of the bath or furnace, and AH,, is the heat associated with any phase change involved. Prior to the instant of ignition, the vapor pressure of magnesium is of no special significance. After ignition, neither eq 4 nor the above equation is applicable. The actual combination of magnesium cannot safely be assumed to take place in the vapor phase. While experimental data is lacking to support a hypothesis that ignition does or does not occur in the vapor phase, some observation on the pressure ignition experiments may be of interest. At high oxygen pressures, once ignition has occurred, the reaction of magnesium with oxygen approaches near explosive violence, the entire sample being consumed in probably less than 1 sec. At atmospheric pressure it usually requires 15 to 20 sec. Thus it appears that the oxygen concentration becomes the rate determining factor. Further, if burning magnesium is observed through darkened glass (Lincoln Super-visibility Shade No. 12) the magnesium sample is very much hotter than the "smoke" and the outline of the sample is retained perfectly. No "flame" is visible above the metal. No work was done on Mg-Be alloys. We do, however, intend to study this problem in the near future.

Jan 1, 1952

By U. Erb, L. Monaco

"Electrodeposited nanomaterials are increasingly implemented in advanced industrial applications in the aerospace, automotive, defence, electronic and energy sectors due to their outstanding physical, chemical and mechanical properties, including increased strength and wear resistance, and superb magnetic properties. As with the development of any class of new materials, the corrosion behaviour of these electrodeposits requires considerable attention. This presentation reviews major developments in the understanding of the corrosion behaviour of nanocrystalline metals, alloys and metal-matrix composites over the last 20 years, with emphasis on how their intrinsically high density of intercrystalline defects does not compromise their corrosion resistance as initially expected.INTRODUCTIONOver the past 20 years there have been considerable advances in the understanding of the corrosion behaviour of nanocrystalline metals, alloys and metal-matrix composites. In this paper, the corrosion properties of nanocrystalline nickel, cobalt, zinc, and copper and some of their alloys are reviewed with emphasis on how a reduction in grain size affects their corrosion behaviour in different environments.It is possible to synthesize nanocrystalline material with microstructural refinement down to the range of 1 to 100 nm using a large number of processing methods. There are five basic synthesis methods that can be used to classify the different techniques: vapor-phase processing, liquid-phase processing, solid-phase processing, chemical synthesis, and electrochemical synthesis. Depending on the synthesis method, different microstructures (equiaxed, textured), shapes, and forms (films, particles, bulk material) may be produced. In addition, other factors may be introduced (eg. residual porosity or amorphous phases), aside from microstructural refinement, which can contribute to the observed properties (Erb, 2010). In order to interpret and categorize the resulting microstructures, a dimensional structure modulation scheme was introduced by Siegel (1996) in which zero-dimensional nanomaterials refer to individual particles, clusters, etc. One-dimensionality results from a layered (less than 100 nm thick) structure. Two-dimensional nanomaterials are thin, layered structures containing grains (less than 100 nm) within each layer. Finally, three-dimensional nanomaterials are bulk structures with an average grain size less than 100 nm."

Jan 1, 2012

By A. Yoshida, H. Miura, S. Yamaguchi, Y. Oyama

A discussion is given on the ionic equilibria and precipitation reactions in seawater based on thermodynamic analysis, in order to understand the solubility of nutrient salts from dephosphorization slag for marine phytoplankton. Solubility of constituent ions in aqueous solution in equilibrium with solid phases are determined using thermodynamic data. It is proposed that CO2 potential defined as log Pco2 should be considered as an additional parameter to the Eh- pH diagram for the ionic equilibrium of slag components in seawater, since magnesium and calcium carbonate precipitation reactions determine the solubility of ligand ions in nutrient salts, such as phosphates, silicate, and sulfate.

Jan 1, 2004

By K Baker, J Purtill

One of the key challenges for the mining industry globally is the rehabilitation of mine waste, and Queensland is no exception. Out of pit waste structures including waste rock dumps and tailings storage facilities, create novel landforms which require ongoing maintenance and pose risks to the environment and sometimes to communities. Where multiple mines co-exist, such as parts of Queensland’s Bowen Basin, individual mines can often be constrained by adjoining operations. This presents opportunities for broader rehabilitation strategies to be considered, more than the current single mine focus, that may provide better environmental outcomes for a region. Given recent global disasters involving tailings storage facilities, and our team’s endeavours to drive Queensland to achieve leading practice in mined land rehabilitation, we are exploring the potential to maximise in-pit tailings disposal opportunities for coalmines through shared waste management strategies. A shared waste management strategy is where two (or more) mines, owned by different companies and operated under different environmental permits, transport tailings from one mine to the other in close proximity, for disposal in a disused pit. This would reduce the number of waste structures to be built, operated, rehabilitated, and managed in perpetuity. Our objectives are to identify and remove barriers to tailings transfers between sites to:  minimise environmental risk  minimise long-term liability of managing mining waste and associated structures  support improved outcomes for the local community. To determine the scale of coal waste disposal practices in Queensland, internal data and publicly available imagery was analysed by the Department of Environment and Science in 2022. Approximately 6500 ha of land in Queensland is utilised for coalmine tailings storage facilities or codisposal areas, with ≈ 3700 ha of this being out of pit.

Aug 2, 2023

By E. J. Roberts

F. C. Bond (Allis-Chalmers Mfg. Corp., Milwaukee) —This paper considers comminution as a first order process, with the reduction rate depending directly upon the amount of oversize material present. The data show that other factors should be taken into account, and it is possible that in time these may be evaluated as simultaneous or consecutive reactions: Development of the theory of comminution has been retarded for many years by the assumption that surface area measurements constitute the sine qua non of the work done in crushing and grinding, and it is encouraging to note the belated growth of other ideas. In the Abstract the term "net power" should be changed to "net energy." Throughout the paper the term "hp per ton" should be changed to "hp hrs per ton", or "hp hr t." The term "Probability Theory" in the title does not seem appropriate, since it is not clear how the probability theory is used in developing the ideas in the paper. There seems to be a contradiction between the large calculated advantages of closed circuit operation and the statement following that the closed circuit test results showed no significant change in grinding behavior, when compared with the batch grind curves. Tables I and II show that between 75 pct and 50 pct solids the energy input required decreases with increasing moisture content and may indicate the advisability of grinding at higher dilutions in certain cases. The calculation of the hp-hr per ton factor indicates an input in the laboratory mill of only 7.32 gross hp per ton of balls; this casts some doubt upon the accuracy of the factor used, since the power input in commercial mills at 80 pct critical speed is customarily much higher. The tests show that within fairly wide limits the amount of ore in the laboratory mill may be varied and a product of constant fineness obtained, provided that the grinding time is varied in the same proportion. This has often been assumed, and confirmation by actual testing is of value. The Cavg corrections for differences between the plant and laboratory size distributions do not seem very satisfactory, since in many cases the plant/laboratory ratio is farther from unity after correction than before. The following equation has been derived from the data in Table VI: Relative Energy (log new ball diam in in. + 0.410) Input = --------------—--------------- from which the relative energy inputs for balls of different sizes can be calculated and compared. The relative energy input is unity for balls of 2.715 in. diam. The equation indicates that the work accomplished by a ton of grinding balls per unit of energy input is roughly proportional to the square root of the total ball surface area; provided, of course, that the balls are sufficiently large to break the material. The data in support of this statement are admittedly meager, but are fairly consistent when plotted. The relative grindability values listed in Table VI for 200 mesh multiplied by 4/5 apparently correspond approximately to the A-C grindability at 200 mesh.' It would seem that for open circuit tests comparable accuracy could be obtained much more simply by the old method' of plotting the test grind, extending the mesh grinds to the left of zero time if necessary, and determining from the plot the equivalent time required to grind from the plant feed size to the plant product size, using the average of several mesh sizes. The en- ergy input value of one time interval could be determined by tests on materials of known grinding resistance, and this multiplied by the interval required should give the desired energy input value. The relative grindabilities would be the relative time intervals required for a specified feed and product size. When the plotted mesh size lines of a homogeneous material are extended to the left beyond zero time they meet at one point at zero pct passing. The horizontal distance of this point from zero time indicates the equivalent energy input required to prepare the mill feed. The author's results show that the closed circuit grinding tests give about the same K values as open circuit tests, from which he concludes that open circuit tests are satisfactory in many cases. The value of the closed circuit test is its ability accurately to predict energy requirements in closed circuit grinding for both homogeneous and heterogeneous materials. If the material is homogeneous, the open circuit test gives satisfactory results; but if the material contains appreciable fractions of hard and soft grinding ore, the open circuit tests will not be accurate because of the accumulation of hard grinding material in the circulating load. Since in most cases it is not possible to determine a priori whether the material contains hard and soft fractions, the closed circuit tests are preferable and more reliable. B. S. Crocker (Lake Shore Mines, Ontario)—Dr. Roberts probability theory of grinding is very similar to our log pct reduced vs. log tonnage method of plotting and evaluating grinding tests at Lake Shore. However, although we both seem to start at the same point we finish with different end results. Shortly after publishing our grinding paper (referred to by Dr. Roberts) in 1939, we did pursue the subject of the "constant pct reduction in the pct +28 micron material for each constant interval of time. We ran innumerable tonnage tests on the plant ball mills, rod mills, tube mills with 11/4 and 3/4 balls, and lastly pebble mills, with tonnage variations from 180 tons per day to 950 tons per day. We found that when we plotted the log of the tonnage against the log of the pct reduced of any reliable mesh, we had a straight line up until 90 pct of the mesh is reduced. We have also tested this in our 12-in. laboratory mill with the same results. We have used this method of evaluating grinds for the past 8 years and developed the recent four stage pebble plant on this basis. By pct reduced we mean the percentage of any given mesh that is reduced in one pass through a mill at a given tonnage (or time). For example, if the feed to a rod mill is 90 pct +35 mesh and the discharge at 500 tons per day is 54 pct +35, the pct reduced is 90 — 54/90 = 40 pct. If the feed had been 80 pct +35 the discharge would have been 48 pct +35 or pct re- duced 80-48/80 = 40 pct as long as the tonnage re- mained constant at 500 tons per day. Thus we can easily correct for normal variation of mill feeds. This log — log relationship derived from the tonnage tests of all our operating mills has proved of tremendous help in checking laboratory work and in designing alternate layouts or new plants. The difference between the log — log and the semi-log plot is only shown up when the extremes in tonnages are plotted. When the relationship between the pct reduced and the tonnage was first investigated, we used semilog

Jan 1, 1952

By Giacobbe Braccio, Alessandro Blasi, Assunta Romanelli, Massimo Morgana, Corradino Sposato

The solvent extraction technique is the only industrial way used to separate and purificate yttrium from other rare earths. In order to improve this process several extractants have been tested during last years. In this work, the behavior of sec-octylphenoxy acetic acid (CA-12) in Yttrium recovery from high concentrated heavy rare earth mixture was investigated. Sodium hydroxide solution was used in order to presaponify organic phase composed by CA-12 and Tributyl phosphate (TBP) diluted in kerosene. In the investigated condition, TBP confirmed its role of phase modifier not significantly altering the extraction behavior of CA-12. The CA-12-TBP system showed a high affinity in extraction for lighter rare earths such as Sm, Eu, and Gd, leaving yttrium in aqueous phase. Using a feed concentration (P [RE]) of 1 M, organic mixture is capable to extract around the 70% of metals in a single extraction test showing this extraction sequence Y < Lu < Yb < Tm < Er < Ho < Dy < Tb < Gd < Eu < Sm. Moreover for [CA-12] = 1.79 M and [RE] = 2 M the instability of extraction system occurs.

Mar 1, 2017

By M. D. G. Salamon

In view of an urgent need for a simple means of simulating the behaviour of stratified rock masses, a long-dormant linear model is revived, generalized, and discussed in some depth. The model is piecewise homogeneous, consisting of a pile of homogeneous isotropic laminae in which the interfaces between beds are parallel and free of shear stresses and cohesion. The continuity of stresses and normal displacements is assured across the bedding planes and, to maintain simplicity, the theory of thin plates is used as the kernel of the model. The paper covers the case where the rock mass consists of discrete beds in which the layers are arranged either heterogeneously or systematically. The stratification is regarded as uniform when all the layers are of the same thickness and their Young&apos;s modulus changes exponentially with depth; and as homogeneous when all the laminae have identical thickness and properties. The case of uniform stratification is also treated on the basis of an equivalent quasi-continuum derived from the basic model. Attention is focused on the derivation of influence functions. It is demonstrated that the quasi-continuous models, when they are available, can replace their distinct-layer counterparts. The homogeneous quasi-continuum, which permits the derivation of its influence functions in closed form, facilitates the determination of the roof and floor convergence. In the case of a significant class of problems, the convergence distribution is defined by a second-order differential equation in the xy plane, which simplifies to a Poisson&apos;s equation if the mining excavations are unsupported. With this result, the paper prepares for the solution of a host of practical problems. These solutions are presented in a series of papers that are in the course of publication.

Jan 1, 1991

By R. Schuhmann, J. Th. Overbeek, P. L. De Bruyn

W. E. Ewers (Commonwealth Scientific and Industrial Research organization, Melbourne, Australia)— Any attempt to elucidate further the meaning of the contact angle, particularly if it deals with the magnitudes of those mysterious quantities rsa and YRL, is certain to receive attention from those interested in the fundamentals of flotation. It is imperative, therefore, that attention should be drawn to a serious defect in the treatment by de Bruyn, Overbeek, and Schuhmann. Their Eq. 5 dy = — ld/u — Th2odiH2n states correctly the relationship between the change in surface free energy dy, the surface excesses Tx and Th2o of the components x and H2O and the changes du, and dph2o in the chemical potentials of these components. This equation applies for constant pressure and con- stant temperature and for the interface of the solid in contact with the solution of X. The relation remains valid when 1.1 is defined so that TH2O becomes zero and the equation becomes dr = — r:dur [17] From this equation de Bruyn et al. correctly concluded that increased adsorption of the collector at the liquid solid interface would reduce the interfacial energy ysl. Their second conclusion, however, that an increase in the contact angle, corresponding to a greater decrease in rsa than in rsL can be attributed to greater adsorption of X at the solid-air interface than at the solid-liquid interface is quite unjustified. The portion of the solid surface from which the collector solution has been displaced by air must be considered in terms of the equilibria between the gaseous phase and the surface. If the pressure and temperature

Jan 1, 1955

By José Sancho, Julia Ayala, Purificación García, Begoña Fernández

"A study is carried out to see the adequacy of using flying ashes coming out from a Power Station to eliminate the copper present in waste cyanide waters that are generated in gold mines.Tests were carried out to see how effective the use of ashes taken from two Power Stations could become. Later, the tests are made using ashes coming from Narcea Power Station as this station is near the mine under study.The work is centred on looking for an economically adequate solution to close the waste decanting ponds of the mine. After several tests, a solution is reached through filtration in static layers.IntroductionThe use of coal to adsorb some water-soluble substances has been widely studied [1,2,3,4,] both in the treatment of drinking waters and in determined stages of selective detachment of cyanide gold complexes in cyanide lixiviation processes [5]. Flying ashes sent to the atmosphere by Power Stations as a sub-product of the combustion of this combustible are generally made of more of a 70 % of a vitreous aluminosilicated material and of an active residual material which may be changeable making them potentially useable as adsorbents. Nevertheless, the properties of each kind of ashes depend on the characteristics of the Power Station, of the combustibles being used, of the combustion system, etc.The aim of this work is to demonstrate the adequacy of flying ashes to eliminate the copper present in cyanide waters coming from gold mines either by adsorption or precipitation of the metal. For this study, tests are carried out with Asturian ashes which have a low quantity of lime. The results obtained in the tests demonstrate the effectivity of the method used. To be absolutely concrete, the work is carried out to establish the use of these ashes in the closing of Asturian installations of cyanide lixiviation"

Jan 1, 2008

By Walter Harvey Weed

INTRODUCTION. IN a previous paper upon the enrichment of mineral veins by later metallic sulphides,† the writer has shown that certain masses of rich ores, such as are found in many mines, either near the water-line or as bonanzas in depth, are of secondary origin, and are due to a leaching of lean ore and the concentration of the material by reaction between the solution and the unaltered ore below. The geological and mineralogical evidence is believed to form an adequate basis for a chemical and physical explanation of the phenomenon. In the present paper the writer will give a brief synopsis of this theory, and will apply it more particularly to deposits of the precious metals, laying special emphasis upon the dependence of such enrichments upon the presence of iron sulphide (as pyrite, etc.) in the primary ore, and upon structural features which control the circulation of the enriching solutions below the water-level. It is believed that many, though not all, of the bonanzas and pay-shoots of rich sulphide ores, especially those carrying gold and silver, which are encountered in ore-deposits, are of such secondary origin. Apparently it is essential that the occurrence and structural relations of such ore-masses should be understood, as the success of the mine is often dependent upon the finding and extraction of these ores. A legitimate deduction, too, is that such ore-deposits decrease in value with depth. As my own studies have been mainly in Montana, my illustrations must be drawn from the ore-deposits of this and adjacent States, with such as I have noted in hurried visits elsewhere. The literature of ore-deposits doubtless affords also

Jan 1, 1902

By R. R. Rogers

Abstract The degree of cathodic protection of steel and the anode corrosion rate were determined when anodes and waters of different compositions were used at temperatures similar to those in domestic hot water tanks. Twenty different anode materials, sixteen of them magnesium-zinc-aluminium alloys, were immersed in six different synthetic waters resembling those to be found in various parts of Canada. The results showed that (1) both the degree of cathode protection and the average anode corrosion rate are affected materially by the composition of the anode material and of the water; (2) poor cathode protection is associated with low anode corrosion rate; (3) when the cathode is fully protected, the anode corrosion rate may be comparatively low or comparatively high, depending upon the composttion of the water; (4) an important relationship exists between the magnesium content of the preferred anode and the composition of the water; (5) the degree of protection is increased as the SO4 + Cl concentration of the water is increased and thus as its specific resistance is decreased. Introduction Fairly recently, the use of the active metal magnesium for the cathodic protection of less active metals has received a considerable amount of attention. This interest has been stimulated by the fact that the wartime demand for magnesium was greatly in excess of the present peacetime demand and there is a natural desire to continue operating the large magnesium plants which were built during the war. Information regarding the use of zinc and aluminium for this purpose also has been published. There is a great demand for domestic hot-water tanks which will resist corrosion satisfactorily. This problem is rather complex in view of the fact that the composition of the water used in such tanks varies greatly and depends upon the locality. The present investigation deals with the effect of water composition and anode composition on the cathodic protection of steel under conditions approaching those which exist in hot water tanks.

Jan 1, 1948

By D. K. Crampton, H. l. Burghoff, J. T. Stacy

The study of copper alloys of the age-hardening type has received considerable attention, and, among the alloys which the authors have, considered, those containing small amounts of nickel and phosphorus as essential constituents seem especially interesting. Such alloys have attractively great capacity for being hardened and also have relatively high electrical and thermal conductivity in the hardened state. In addition, it has been found that the increase in hardness and strength imparted by aging can be greatly augmented by cold-working, so that excellent combinations of conductivity and strength are obtainnble. In view of the possible importance of the alloys, it was considered desirable to investigate the solubility relationships in the copper-rich portion of the copper-nickel-phosphorus system. This paper present&apos;s some of the results obtained in this investigation to date and also gives propertics of typical alloys that appear to be of commercial interest,. Review of literature Copper-nickel System,.—The binary coppcr-nickcl system has long been known to consist of a series of simple solicl solutions, copper and nickel being miscible at all tempcrat&apos;ures and concentrations. This was establisl~cd by Guertler arid Tammannl in 1907 and confirmed by Tafel2 in 1908. Copper-phosphorus System.—The copper phosporus system has been studied by numerous investigators, so that the copper-rich portion is now definitely known. As most recently determined by Lindlief," his system has a cutectic at 8.38 per ccnt P, cuteetic temperature being 714° C. Components of the cutectic are Cu3P and alpha solid solution. Solid solubility of phosphorus in copper, according to Mertz arid Mathewson decreases from a maximum of about 1.75 per cent at 714&apos; C. to about 0.6 per cent at 300&apos; C.

Jan 1, 1940