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By C. S. Yust, C. S. Morgan

ThIS note reports the observation of masses of dislocation etch pits around the weld necks of small single-crystal particles of calcium fluoride sintered to a cleaved face of a larger CaFz crystal. Crystal particles in the -50 +I00 mesh size range were placed on freshly cleaved CaFz surfaces and heated at 1250°C for 30 min in a tantalum container in an argon atmosphere. The sintered specimens were then etched at room temperature for 40 to 50 min in a 2 pct sulfamic acid solution which is a selective etch for dislocations in caF2.' Figs. 1 through 3 show the masses of dislocation etch pits observed around weld necks after etching. In some cases, particles were gone after removal of the specimens from the furnace although weld necks were readily visible. In other instances, no etch pits formed around a weld neck. A possible cause of concentrations of dislocations on the crystal surface is mechanical deformation of the base crystal when particles are dropped on it. If such a deformation does occur, it would be much smaller than that caused by the light punching of a crystal with a sharp instrument which results in large concentrations of etch pits. However, crystals deformed by punching with the sharp instrument and annealed at the conditions of these experiments (1250°C for 30 min) before etching showed no concentrations of dislocation etch pits. It is obvious that dislocations from the much smaller deformations, possibly created when particles are placed on the surface, would also anneal out. Etching of large crystals which had particles placed on them in the usual manner but which were not heated showed no dislocation concentrations such as shown in Fig. 1. The magnitude of dislocation etch-pit concentrations did not vary with extensive variation of the cooling rate. The appearance of many of the etch-pit clusters is typical of a well-annealed condition, indicating that the dislocation concentrations formed before the specimen temperature was lowered. Therefore, it is evident that thermal stresses in the neck area did not create the deformation indicated by the presence of the dislocations. The possibility exists that the etch-pit concentrations at places where particles had been removed may be the result of deformation accompanying the fracture of the sintered bond. That this is not so was demonstrated by breaking off particles around which no etch pits occurred. Re-etching never resulted in the large concentrations observed at sintering sites. The stress available to effect material movement during sintering of two particles will depend on the geometry at the contact point but can easily exceed the macroscopic yield stress in CaFz systems. It appears that the plastic deformation observed results from sintering or simply from the weight of the small particles. It has been previously argued that plastic flow makes an important contribution to the initial densification of oxides having the fluorite crystal structure.2p3 If plastic deformation occurs, it would begin as soon as the temperature is sufficiently high

Jan 1, 1967

By Sandy Clamage

"Seeing the elephant" was a term often heard from 1849 to 1859, the period of the great California Gold Rush. The phrase was used by those who dreamed of instant wealth that was believed could be found in California's virgin earth and mighty rivers. When news of fabulous gold strikes throughout the Mother Lode reached far away places, thousands of gold seekers left home and family to join in this incredible adventure. The barren, placid state of California soon became one of the most bustling, populous areas in the West. Before the Rush Before 1848, Mexico ruled California. It was then a remote region, isolated from the urbane world of the East by 2900 km (1800 miles) of scorching desert and impassible mountains. By sea, it was 29 000 km (18,000 miles) via Cape Horn. Spaniards and Mexicans settled the coastal lands on tracts donated to them by the Mexican government. The inland area was the domain of Indians and a few settlers who also had been granted estates by the government. Among those settlers was Captain John Sutter. He had been fortunate enough to obtain about 11 leagues of land (spanning 100 km or 60 miles in length) from Governor Alvarado of Mexico. Sutter was somewhat of a visionary and possessed a gifted tongue. He managed to receive most of what he desired with minimal trouble. Enchanted with Sutter's personality, Alvarado promised him Mexican citizenship. Alvarado also commissioned him to become "representative of the government and officer of justice on the Northern frontier." With uncanny foresight, Sutter began building a fortress on his land in the Sacramento Valley. It later became known as Sutter's Fort. It was more than just a fort, however. It also became a stopping place and home for many foreigners and Americans who traveled to California, and it eventually became the mecca for overland pilgrims. Sutter engaged in many projects. These included a cattle ranch, fur trade and trapping, and the construction of an irrigation system. With the influx of immigrants over the years, the community around the fort grew rapidly. Lumber was much in demand and Captain Sutter agreed to supply the necessary capital for a community sawmill. The site chosen was some 72 km (45 miles) northeast of the fort on the south fork of the American River in the town of Coloma. During the course of its construction, Sutter's carpenter, James Marshall, discovered a problem with the tailrace on the mill. First Strike Early in the morning on Jan. 24, 1848, Marshall was checking the tailrace. He noticed an object shining in a foot of water at the bottom of a ditch. He reached down and picked it up. His heart began to thump wildly as he realized he was holding a gold nugget. Sutter planned to keep the find a secret for as long as possible. He knew very well that he could become an extremely wealthy man. The ever loyal Marshall remained faithful to his employer and the pair did their best to obtain more

Jan 3, 1985

By Jon J. Connor, Richard J. Ebens

The trace element geochemistry of the cherty clay-rich carbonate residuum cover in much of southern Missouri and adjacent parts of Arkansas was investigated during the earl 1970's as part of a statewide geochemical survey of Missouri.1-8 The study was environmentally oriented and focused on the character of the "natural" geochemical environment of the state. A great deal of effort was made to collect samples that would represent the geochemical variability of the near-surface environment in an unbiased way. In general, this was accomplished by use of randomization procedures in both collection and analysis. Most of the sampling sites were selected randomly to prevent personal bias from entering into sample selection. In the laboratory, samples were analyzed in a randomized sequence to circumvent the effects of any systematic bias that may have entered into the laboratory procedures. Several residuum samples contained 1000 parts per million (ppm) or more of lead, zinc, and barium. All these samples were collected in Washington County in or near the barite district. The frequency distributions shown in [Fig. 1] indicate just how unusual the residuum in Washington County is with regard to these three metals and copper. The shaded parts of the histograms represent the samples from Washington County. The histograms are based on analyses of 141 samples from 38 localities scattered over the subcrop of the Cambrian and Ordovician formations of southern Missouri and northern Arkansas. Eight localities were selected randomly from over each of four bedrock units-Bonneterre dolomite; Potosi and Erninence dolomites; Roubidoux formation; and the Jefferson City, Powell, and Cotter dolomites. Six localities were selected randomly from over the Gasconade dolomite. The geochemical "background" for these elements in residuum is defined by samples from outside of Washington County (unshaded part of the histogram); lognormal distribution curves have been fitted to the corresponding parts of each histogram. Clearly, these lognormal models (curves) describe well the background distributions of zinc, barium, and copper. The background histogram of lead concentrations is notably skewed to the right compared to the model. Lognormal distribution curves were also fitted to the data on lead, copper, and zinc in the residuum of Washington County. No model was fitted for barium in Washington County because of the distinct bimodality in the data. All samples containing more than 1000 ppm were collected in the barite mining district and obviously reflect the elevated levels of barium in the district. The frequency distributions for Washington County are clearly anomalous with respect to the background data, those for lead, zinc, and barium being exceptionally so. The models for anomalous lead, zinc, and copper seem to fit reasonably well, although the scatters in the lead and zinc anomalies are considerably larger than the scatters in the corresponding background data. The copper anomaly, although probably real, is greatly subdued compared to that of lead and zinc. Also, it has a scatter similar to that of the copper background, rather than being enhanced like the scatters of the lead and zinc anomalies. On the basis of techniques described by Miesch,4 the chemical abundance of copper, lead, and zinc in residuum in

Jan 1, 1980

By G W. Wilson

Rehabilitation of tailings impoundments is one of the most challenging aspects in mine closure, as not only does the potential for producing leachate pose a challenge to the rehabilitation designer, but also other aspects such as stability and settlement must be considered. The water balance of a tailings impoundment is unique in the sense that it usually hosts a pond that in turn causes a phreatic surface in the impoundment. The position of the phreatic surface defines the saturated and unsaturated zones in the impoundment, which of course varies spatially and temporally. Predictive modelling for this hydrologic system becomes difficult, as numerical models capable of analysing the combined saturated/unsaturated zones are not adequately refined to accurately solve the flux boundary problem for infiltration at the surface of the tailings. This paper describes the development of a flux boundary model that enabled accurate modelling of the unsaturated zone in the tailings impoundment at Kidston Gold Mine, Queensland, Australia. The technique made it possible to accurately predict the spatial variation of infiltration to the tailings as a result of the presence of the phreatic table. The conceptual model defines the boundaries for the Kidston tailings dam, based on a generalised cross-section through the dam. Firstly the shape of the top boundary is described as a function of the particle size of the tailings associated with particle segregation that occurs due to the hydraulic placement of tailings. This function was rigorously tested using in situ infiltration tests and laboratory materials testing. Defining the bottom boundary of the section, ie the phreatic level was done by physically monitoring the phreatic level with 42 piezometers strategically located on the tailings dam. Analysis of this data led to the fixing of this bottom boundary through the introduction of a function that mimicked the shape of the phreatic level. The material properties of the surface layer, specifically the saturated permeability that would govern the infiltration/exfiltration rates of moisture through the top boundary layer were defined on the basis of extensive field and laboratory material testing. A function for the saturated permeability was developed on the principles of particle segregation. Finally a procedure is described for the application of the new conceptual model for the evaluation of the spatial infiltration to a tailings surface using the rigorous one dimensional surface flux boundary code SoilCover. The solution of the calculation described above was then used as the top boundary condition in conventional three dimensional saturated/ unsaturated flow codes, and verified against measured seepage rates from the tailings impoundment under drains. It was concluded that the developed spatial flux function was in deed a most rigorous representation of reality. Although the methodology presented here was applied to the Kidston tailings impoundment only, the authors believe that the principles are directly transferable to any tailings impoundment that has a varying phreatic zone thickness due to the presence of a pond. The mathematical equations describing the typical cross-section would be site specific, but the approach and ultimate solution methodology would be directly transferable.

Jan 1, 2003

By Nelson R. Shaffer

Implementation of the Clean Air Act Amendments of 1990 is negatively affecting production of coal used for generating electric power in Indiana and elsewhere in the Midwest. Required reduction of sulfur compounds from gas streams, or flue gas desulfurization (FGD), as mandated by these laws has encouraged installation of wet limestone scrubbers at many generating plants. Increasingly stringent requirements that will begin in 2000 are expected to necessitate even more extensive use of such scrubbers. Indiana has large deposits of limestone-the major consumable alkaline raw material for scrubbers-that can meet an expected demand of several million tons per year. Limestones are not, however, all the same. Different limestone units behave quite differently when used as scrubbing agents under differing conditions in actual practice. Decisions about FGD scrubber details would be more meaningful if made with thorough knowledge of which rock units contain the most efficient FGD limestones and which sites of suitable stone are located nearest to power plants. Even though abundant carbonate rocks occur in Indiana, their suitability in scrubbing systems has received little attention, and few attempts have been made to match details of stone resources to the needs of various utility scrubbing systems. Optimization of the fit between geologic characteristics of the existing limestones and prospective scrubber systems should allow for greatly improved planning and operation of FGD systems with attendant lower overall costs. Electric utilities have developed computer models to predict FGD system operations (i.e., the FGD-PRISM model of the Electric Power Research Institute). These models require geologic data such as reactivity that are commonly unavailable. We generated data needed for the model during a cooperative effort among the Indiana Department of Commerce, the Indiana Geological Survey (IGS), The Indianapolis Power & Light Company (IP&L), and the Indiana Mineral Aggregates Association. Our goal was to provide missing geologic details of Indiana scrubber-stone raw materials and apply them to the FGD-PRISM model to provide general guidance to determine which geologic units can provide stone for most efficient sulfur removal. Results of the model for selected limestones and parameters specific to one IP&L generating plant were calculated. Geologic data generated may be applied to other generating operations, and our examples should help other utilities optimize their FGD strategies. Several hundred samples from more than 30 active quarries were collected and their physical properties determined. Grindability values ranged from 7.2 to 21.7 for 50 selected samples. Acid-insoluble residues of less than 1% to 13% were found. Dissolution rates as an index of reactivity varied over two orders of magnitude for the selected set of samples. These data plus chemical analyses and Ca/Mg ratios were entered into a generic scrubber model to determine rates of stone use and scrubber efficiency for various stone. Removal of SO2 was projected by this model to range from 89% to 98% with stone utilization rates from 91% to 93%. Excellent scrubber stone sources were found in Mississippian rocks in central and southern Indiana. Rocks of the Ste. Genevieve and Salem Limestones were especially efficient. Even small amounts of dolomite, clay minerals, and quartz diminished a stone's usefulness for FGD scrubbing.

Jan 1, 1999

By F. F. Roberto

Sulfuric acid has been demonstrated to be an efficient lixiviant for in situ extraction of copper in the past. It has also been proposed that microbial production (by iron-oxidizing acidophilic bacteria) of acidic ferric solution either in situ or ex situ could provide an economic strategy to generate an effective lixiviant for secondary copper sulfides like chalcocite and covellite. Copper heap bioleaching also takes advantage of this approach. One of the major environmental challenges for in situ copper recovery projects is the containment of acidic solution within the ore deposit, and ultimate restoration of the depleted ore body to its original state. A perhaps overlooked aspect of projected site operations is the possibility that transient acidic conditions and pumping of solutions from surface reservoirs into the subsurface could stimulate the growth of iron- or sulfur-oxidizing species of acidophilic microbes in the ore body, leading to additional generation of sulfate, acid, and ferric iron. This outcome would benefit extraction of copper in mixed oxide/sulfide deposits in the near term, but could complicate restoration activities. Possible impacts of pumping acidic solutions into the subsurface on microbes resident in, or introduced into the extraction zone will be discussed in the context of our understanding of copper heap bioleaching and acid rock drainage. INTRODUCTION As high-grade deposits of metals such as copper continue to decline, conventional mining methods have evolved to mine and process larger and larger quantities of ore. What is currently defined as ore today often contains less metal than the tailings of past operations. The throughput of current-day mines also leads to tremendous amounts of waste rock and tailings that may threaten the environment and communities around the mines, as has recently been experienced by 3 major tailings disposal facility (TSF) dam failures in the past 5 years. Community acceptance of large mine operations thus becomes more difficult to achieve—loss of the social license to operate. In response, the mining industry urgently pursues innovations that can fundamentally change the way in which metals are mined. In situ recovery (ISR) of metals represents one approach to a more sustainable mining industry, by virtually eliminating the need for a pit or underground excavation, eliminating the excavation and transportation of ore to a process plant and the ensuing noise and dust that is generated, and any need for waste rock dumps or TSFs. An ISR mine could have very little visible footprint beyond a network of wells, pumps, and piping to transport pregnant solution to a recovery facility, such as a solvent extraction- electrowinning (SX-EW) plant for copper. Some solid waste streams and/or storage ponds may still be necessary to treat and store bleed streams, PLS and raffinate, or as part of groundwater restoration activities. Thus capital and operating investments would be potentially much lower than that required for a conventional mine. In theory production could also be controlled to adjust to subsurface ore grade and match market demand. A not insignificant challenge of in situ recovery is the requirement for an efficient, low-toxicity, and ideally selective lixiviant for the metal that is targeted.

Jan 1, 2019

By J. Allen Hennessy

INTRODUCTION The increasing price of gold and silver over the past 10 years has caused increasing activity in the exploration, evaluation and development of precious metals projects. Many of these deposits are relatively low grade and short lived which makes them economically sensitive to the major project variables such as metal price, capital cost, operating cost and of course, tonnage and grade. These deposits often have a high variability (a coefficient of variation -- o/u -- of 2 to 5) and an intermixing of ore and waste, making ac- curate ore reserve estimation difficult, especially when using traditional estimation techniques. Often, the estimation errors are not the 10-20% that we see in base metal deposits, but the errors in grade and tonnage can be as high as 50%, which can obviously have a very substantial impact on project Recent developments in geostatistics when coupled with a sound geologic under- standing of the deposit can significantly improve reserve estimates of these variable precious metals deposits. Examples will be given of actual deposits where the traditional estimates were significantly different from the mining reality. An attempt will be made to explain the applicable geostatistical tools with a minimum of geostatistical jargon and without complicated formulas. The key to improving the reserve estimates is a clear understanding of the deposit geology and especially the structural controls, the use of local estimation techniques that recognize the intermixing of ore and waste and the effect of selective mining on recoverable reserves, and the use of a global reserve estimation technique to check the results of local estimation. Geologic Controls The vast majority of gold and silver deposits being mined or developed in the western United States would be classified as epithermal deposits. These are formed by hydrothermal fluids (hot water) flowing upward through the rocks and depositing the precious metals as the fluids cool or reach areas of lower pressure or favorable chemical sites. These environments are similar to present day hot springs-type environments with some deposits being clearly associated with surface hot springs such as the Homestake McLaughlin deposit in northern California. These hydrothermal deposits can occur in sedimentary rocks resulting in deposits such as those of the Carlin district or the Jerritt Canyon deposit or the hydrothermal systems can occur in volcanic rocks forming deposits such as Round Mountain, Manhattan or Paradise Peak. The mechanisms of formation were similar in either case with the solution flow controlled by geologic structure and precious metals deposited when the conditions of temperataure, pressure and chemical environment were favorable. The gold and silver hearing solutions travel through permeable faults and fracture systems depositing the metals along these channels. The grades are often highest in the central parts of these conduits (veins) with lower grades in the surrounding rocks. An understanding of the nature, distribution and orientation of these plumbing systems is important to a realistic reserve estimate. The point to be emphasized here is that, good, sound geology in the form of geologic mapping, logging of rotary chip samples and a few carefully logged core holes can signifi- cantly improve an ore reserve estimate.

Jan 1, 1987

By E. R. Johnson

THE copper concentrator of the Braden Copper Co. is at Sewell, Chile, on the western flank of the main Cordillera of the Andes, at an air distance of approximately 50 miles southeast of Santiago, the capital of Chile. The region is mountainous and space is at a premium because of the rough terrain and the danger of snowslides. The ore, a porphyry copper, is mined by a block-caving method in the El Teniente mine. It is hauled to the crushing plant, a distance of 2.6 miles, through a tunnel because of the severe climatic conditions. The usual large, coarsely set, primary reduction machines found in many ore-treating plants are not required because of the mining method employed and the limiting 12-in. spacing of the grizzlies at the head of each ore pass in the mine. It is the purpose of this paper to describe the numerous improvements made in the crushing section of this large flotation concentrator during the past 15 years to increase the efficiency of the plant and to raise the daily output from around 20,000 to 30,000 tons. The reader should keep in mind that all changes were made within the existing building because of the limited space and the high cost of the heavy mill construction required to withstand the snow loads. The present plant, as well as the former, operates on a 24-hr, 7 days a week basis. FORMER PLANT The crushing plant commenced operations at its present site in 1921, the main equipment being No. 7 1/2 gyratories, followed by 48-in. horizontal Symons disk crushers. Fines were removed by 1-in. stationary grizzlies ahead of each unit. In 1924 a third stage of crushing was provided by the addition of 55 1/2-in. rolls in open circuit. During 1925, the Symons disks were replaced by 73 1/2-in. rolls. All the rolls were changed to closed circuit, with elevators and screens to give a final product of 9 pct plus 3-mesh. The flowsheet of a typical crushing section of this period is shown in Fig I. The ore was withdrawn from a common bin serving the five crushing sections in the plant by means of any number up to six 36-in. wide-apron feeders, which discharged onto a common conveyor belt. This conveyor was equipped with a magnetic head pulley and the crusher attendant was stationed near the discharge end to pick out pieces of mine timber that might be in the ore. The ore was discharged over a 1 3/4-in. spaced bar grizzly and the oversize of this grizzly was crushed by a No. 7 1/2 McCully gyratory crusher set at 1 3/4-in. opening on the closed side. The grizzly undersize joined the gyratory discharge and was conveyed to a set of 73 1/2 by 20-in. Garfield rolls set at I in. These rolls operated in closed circuit with a bucket elevator and two vibrating screens placed in parallel.

Jan 1, 1947

By S. F. Reiter, You-Chao Liu, W. R. Hibbard

Cyril Stanley Smith in his classic 1948 Institute of Metals Lecture&apos; noted in the case of annealing twins that: "It is well known that approximately the same number of twins per grain and the same general geometry are maintained in large grains as in small ones, a fact that certainly suggests a mechanism such as the above where boundaries and twins grow proportionately together. If this view is correct, a spherical grain should have no mechanism for keeping its twins." Such a constancy of twins per grain would be of particular interest as a tool for the study of the origin of annealing twins. It has been proposed by Mathewson2 and later by Barrett3 that in face-centered cubic metals slip occurs in the {111} Plane by <112> shears which combine in pairs to produce the macroscopically observed <110> movement. This proposal is schematically inviting since <112> slip permits the atoms of one plane to move along the interatomic valleys of its neighbor rather than up the hills across the atom sites. There is strong experimental evidence? that annealing twins result from the growth of nuclei in the form of "twin faults" produced by single <112> shears. It is possible that more twin faults would form at low deformation temperatures than at high ones, because the more rigid lattice may increase the tendency to "freeze" the twin faults and prevent their dissipation in slip. This is suggested by (1) the greater resistance to slip at low temperatures and (2) greater tendency toward mechanical type twinning at low temperatures (a summary of these factors is presented by Seigle and Brick5). These considerations would suggest the following: a. Constant twin/grain ratio for specimens deformed constant amounts at a constant temperature, but annealed at various temperatures above the recrystalliza-tion temperature. b. Larger twin/grain ratios for re-crystallized specimens previously deformed constant amounts at lower temperatures. The purpose of this investigation is to test these considerations. One half in. rods of phosphorous deoxidized electrolytic copper and 70-30 alpha brass (68.36 pct Cu, 0.005 pct Pb, 0.01 pct Fe, 31.62 pct Zn remainder) were rolled to a thickness of 0.350 in., milled to specimens 0.350 X 0.437 X 2 in., and annealed at 700°C to an average grain size of 0.065 and 0.090 mm, respectively. These specimens of each metal were then rolled 50 pct reduction in thickness and annealed one hour under conditions which were all combinations of the following: Rolling Annealing Temperatures. Temperatures. °C °C -70 400 25 500 200 600 Standard metallographic techniques&apos; were used to prepare specimens. By the method of lineal analysis7 using a minimum of ten samples of fifty grains each for each specimen, the ratios of twins* per total grains, twins per twinned grain, twin families* per total grains and twins per twin family were determined. Recrystallized grain size was determined by A.S.T.M. comparison methods.8 All observations were made on the rolling plane. The results are tabulated in Table 1.

Jan 1, 1950

By Peter J. Huck, Bruce G. Stegman, Richard E. Thill

Introduction Backfilling of mine voids is used to prevent or control the effects of subsidence on surface structures. The pumped-slurry, blind backfilling process has been successful for stabilizing ground over abandoned and inaccessible underground coal mines. This article describes the systematic evaluation of potential monitoring systems that would improve backfill monitoring technology. It also details field proof-of-concept tests that were conducted on two systems and the difficulties encountered in implementing them. Ground subsidence over abandoned mines can have devastating effects on overlying urban areas decades after the closing of room-and-pillar mines. The result may be cracked foundations, ruptured gas and water lines, broken sewers, distortion or cracking in superstructures, and sinks and potholes in the ground surface. About 32,000 km2 (8 million acres) of land in the US has been undermined for coal. According to the US General Accounting Office, 8,100 km2 (2 million acres) have undergone subsidence and another 8,100 km2 (2 million acres) are expected to subside by the year 2000. Actual figures on land affected or threatened by subsidence are probably much higher, since these estimates were based on late 1960s studies. For more than two decades, the US Bureau of Mines has been engaged in stabilizing ground in urban areas having high risk potential for subsidence costing millions of dollars. Thousands of acres have been stabilized, protecting property worth hundreds of millions of dollars. In earlier years, backfilling was by in-mine stowing in accessible mines or by gravity-feed sluicing methods in inaccessible mines. Pumped-slurry methods have been used extensively in the past 10 years. More slurry can be injected for fill from fewer injection holes over extensive areas in abandoned mines. Since this method requires fewer boreholes, there is less disruption of surface facilities. An investigation was conducted to determine the feasibility of using advanced, remote sensing or monitoring technology for application in monitoring fill placement. Specifically, the objective was to develop conceptual systems for blind backfill monitoring that require fewer boreholes and give better definition than is now possible. Blind Backfilling Methods Blind backfilling operations are conducted from the surface and do not require personnel and equipment underground. These methods are applicable to abandoned, inaccessible workings. Two categories of blind back-filling are point support, practiced by civil engineers to protect individual structures or surface facilities, and areal backfilling methods, used to protect large areas against subsidence. Point Support Methods Point support methods are usually gravity-feed systems. In general, point support methods use small volumes of expensive fill material and require a large number of boreholes within a constructed site. Because of the processed material used for back-filling and the close spacing of the injection boreholes, control and monitoring often may be accomplished by borehole cameras. Areal Backfilling Areal backfilling is conducted mainly by the pumped-slurry injection process (Fig. 1). In eastern and interior coalfields, material used is often mine refuse or flyash, both being undesirable on the surface and costly to dispose of in an environmentally acceptable manner. Hence, the use of

Jan 12, 1983

By C. F. Vos, J. D. Miller, J. C. Davidtz

"This work presents the potential for improving the flotation recovery of slow-floating sulphide minerals with the use of starvation dosages of a normal dodecyl (n-C12) trithiocarbonate (TTC) co-collector, together with a sodium isobutyl xanthate (SiBX) and dithiophosphate (DTP) collector mixture.At potentials below –150 mV (SHE), addition of nC12-TTC with SiBX improves the hydrophobicity of pyrrhotite, yielding captive bubble contact angles greater than those measured for SiBX or nC12-TTC alone, suggesting a low potential synergistic effect. This synergistic effect is further studied using Fourier transform infrared (FTIR) spectroscopy, the results indicating an increase in the surface concentration of the collector species when in a mixture. Thus, nC12-TTC with SiBX may act as an immobile surface anchor to which SiBX/SiBX2 molecules bond, increasing the localized concentration of collector species.Bench-scale flotation tests using mixtures of SiBX/DTP/nC12-TTC on a platinum group element (PGE)-bearing ore from the Bushveld Complex in South Africa confirm an improved metallurgical performance at very low substitutions (approx. 5 molar per cent) of SiBX. The improved recoveries for PGE, Cu, and Ni are correlated with improvements in the flotation kinetics of their slow-floating components. IntroductionShort-chain (less than C6) xanthates are very effective for the bulk flotation of sulphides (Fuerstenau, 1982), and early work (Leja, 1968; Gaudin, 1957; Plaksin and Bessonov, 1957; Taggart, Giudice, and Ziehl, 1934) indicated that the presence of oxygen in flotation slurries is necessary for the oxidation of the xanthate collector at the sulphide mineral surface to induce hydrophobicity. Fundamental research into the chemistry and adsorption mechanisms of xanthates over the years (Finkelstein and Poling, 1977; Woods, 1976; Winter and Woods, 1973) has revealed that for iron-bearing sulphide minerals, adsorption occurs predominantly through a charge transfer process. Oxidation of the adsorbed collector to the corresponding dimer takes place if the mixed potential of the system is greater than the reversible potential for dimer formation. In oxygenated slurries the rate of adsorption and oxidation of the collector depends greatly on the substrate surfaces. For chalcopyrite (Guler et al., 2005; Leppinen, 1990; Roos, Celis, and Sudrassono, 1990) and pentlandite (Hodgson and Agar, 1989) a two-step interaction with xanthate is proposed in which the chemisorbed xanthate is further oxidized to the dimer. The initial interaction of xanthate with pyrrhotite is through physisorption where xanthate physisorbs onto positive surface sites followed by oxidation to dixanthogen (Khan and Kellebek, 2004; Bozkurt, Xu, and Finch., 1998). This process is, however, slow and increased reaction time is needed to improve flotation recovery (Buswell and Nicol, 2002)."

Oct 1, 2018

By R. E. Ogilvie, N. L. Peterson

DURING the past few years there has been considerable effort devoted to the understanding of radiation effects on materials. The concept of the Frenkel defect and other lattice defects suggests that radiation may alter the rate of many metallurgical phenomena, especially those which are diffusion controlled. In 1953 R. R. Coltman and T. H. ~lewitt&apos; found an appreciable effect on the ordering rate of Cu,Au due to fast neutron bombardment. Since then several rather comprehensive theories have been proposed on radiation enhanced diffusion in solids.2"4 If the energy of the radiation is above the threshold energy for the removal of atoms from their lattice sites, Frenkel defects will be produced. These defects can be produced uniformly throughout a relatively thick specimen by high-energy X-rays permitting bulk property studies without the limitation of sample dimensions. In this work high-purity arsenic containing 30 ppm of impurities was allowed to diffuse from the gase- ous phase at 700°C into single crystals of p-type germanium containing 0.1 ppm of indium. As the arsenic diffuses into the P-type germanium, an n-type layer will form in the diffusion zone. At the p-n junction the concentration ratio of the arsenic to the indium is constant. Experimental Procedures—The germanium specimens, 10 mm by 10 mm by 5 mm, were polished and etched to a mirror finish. The specimens were then sealed off in a vycor glass capsule at 3 x 10"5 mm Hg pressure with a sufficient amount of arsenic powder to produce a one-half atm pressure of arsenic at 700°C. Specimens were annealed at 700°C without radiation. The diffusion coefficient determined from these samples agrees within 8 pct of previous work done by Fuller.&apos; Three specimens were given diffusion anneals at 700°C in a small tube furnace centered beneath the target of a 2-mev Van de Graaff generator. The specimens when in the furnace were 10 cm directly below the target of the generator. These specimens were irradiated continuously during the diffusion run with the Van de Graaff operating at 2 mev with a tube current of 225 to 250 pamp. The intensity of the x-radiation at the surface of the specimen inside the furnace was approximately 4.5 x l05 roentgens per hr. The temperature was controlled within m2°C by a chromel-alumel thermocouple throughout this work with no observed temperature change due to the radiation. By careful lapping, the samples were polished at an angle of 5 deg to the diffusion zone, giving a XI0 increase in the visible diffusion zone. The location

Jan 1, 1960

By C. A. Snyders

The research presented in this paper investigated the practical aspects of the recovery of platinum group metals (PGMs) from a dilute cyanide leach solution containing base metals, in a manner similar to that used for gold extraction in a typical CIP process, and focuses on both the adsorption and elution stages. The carrier-phase extraction of precious metals using activated carbon offers significant advantages over other processes in terms of simplicity, the high pre-concentration factor, rapid phase separation, and relatively low capital and operating costs. As a sorbent, activated carbon is still by far the most important material because of its large surface area, high adsorption capacity, porous structure, negligible environmental toxicity, low cost, and high purity standards. Adsorption tests were conducted on a pregnant alkaline leach solution (0.15 ppm Pt, 0.38 ppm Pd, 0.1 ppm Au) resulting from cyanide extraction performed in column leach tests. The initial adsorption rates of Pt, Pd, and Au were very fast and recoveries of these three metals were approximately 90 per cent after 2 hours, and 100 per cent for Pt, 97.4 per cent for Pd, and 99.9 per cent for Au after 72 hours. The parameters that influence the extraction of PGMs and Au were examined to assess their relative importance during the adsorption process in order to provide the basis for process optimization. The concentration of thiocyanate was not identified as significant factor for PGMs adsorption, while Ni concentration was the most significant extraction process parameter. Base metal cyanide complexes adsorb and compete with the PGM complexes for sites on activated carbon, and while copper adsorption can be minimized by adjusting the residence time, Ni adsorbs at approximately the same rate as the PGMs, influencing the loading capacity and adsorption kinetics of the PGMs. The feasibility of eluting platinum and palladium cyanide complexes from activated carbon using an analogue of the AARL process was investigated. Platinum and palladium elute from activated carbon almost to completion in 4 to 5 bed volumes at 80°C, while the elution of gold at this temperature is slow, with a significant amount of gold still to be eluted after 16 bed volumes. The equilibrium loading of gold is exothermic in nature (Fleming and Nicol, 1984) which will result in an increase in gold elution kinetics with an increase in temperature at similar pre-treatment conditions. A similar result was found for the elution of Pt and Pd. Cyanide pre-treatment was found to have a significant influence on PGM elution. Higher cyanide concentration in the pre-treatment step results in more efficient elution up to a point, and results suggest the possibility of an optimum cyanide concentration, beyond which elution efficiency starts decreasing due to increased ionic strength.

Jan 1, 2013

By Jamal Rostami, Ataollah Rahbar, Jürgen Paulzen

"Uma Oya Multipurpose Development Project is located in South West Sri Lanka. This project, among many components, involves a 14 km headrace tunnel in very hard and abrasive igneous rocks. The tunnel is under construction by a 4.4 m diameter double shield TBM. The headrace tunnel has already reached 5km milestone and has passed through some reaches of tunnel in very strong intact rocks with no joints. Low penetration at high thrust load has caused frequent stoppages for cutter change leading to low utilization and advance rates due to low cutter life. The original machine was dressed with 27, each 432 mm (17 inch) diameter, 19 mm (3/4 inch) tipped disc cutters. However, various cutter diameters of 432-457 mm (17-18 inch) and cutter tip widths 13 to 19 mm (0.5-0.75 in), have been tried to see if cutter life and penetration rates could be increased. This paper will review the history of the project, the geology of the site, along with TBM performance and cutter life. It will focus on frequency of different cutter failures encountered in various ground conditions. Also, results of using variation of cutter diameter and tip width to cope with the hard rock conditions and their results in increasing machine productivity will be discussed. This allows for better understanding of parameters controlling disc wear/life and leads to better cutter management for more efficient operation of hard rock TBMs in similar conditions. INTRODUCTION The Uma Oya Multipurpose Development Project (UOMDP) lies in the south-eastern part of the central highland region of Sri Lanka. The Multi-Purpose Project Of Uma Oya is planned to convey water from Uma Oya to Kirindi Oya Basin for hydroelectric energy generation includes Puhulpola RCC Dam on the Uma Oya river, Dyraaba RCC Dam on the Mahatotilla Oya river, a conveyance tunnel from Uma Oya to Mahatotilla Oya (L=3750m), a long feeder tunnel from Mahatotilla Oya to Kirindi Oya river. The feeder (or power) tunnel Lot consists of a 15554 m long headrace tunnel (HRT), the surge shaft and tank on the upstream of the vertical shaft, a 650 m vertical shaft with steel lining, and an underground powerhouse and a 3650m tailrace tunnel (TRT)."

Jan 1, 2016

By E. Lehtila, J. J. Talja, R. Jahila

Kopar Oy is acknowledged for engineering and manufacturing of robust systems for bulk material handling, especially at power plants and non-ferrous smelters. We have over thirty years of experience with systems specifically designed for hot and abrasive conditions, aiming at low cost of ownership, low operational expenses and high availability. Although Kopar has numerous applications in mechanical conveying, there is an increasing number of pneumatic conveying applications at copper smelters. They relate to conveying of concentrate, and primary and secondary flue dusts. Challenges of pneumatic conveying for concentrate are mechanical and do relate to broad particle size distribution and high abrasivity of the concentrate and the silica. Therefore, special attention shall be paid to system design and ceramically lined bends in the pipeline. Challenges of pneumatic conveying of primary flue dust relate predominantly to material stickiness and poor flowability which are mainly due to fine particle size and formation of sulphates and furthermore, crystalline water. Additional challenges of secondary dust, the white converter dust, are primarily due to the extremely fine 0.1–1.5 µm particle size and problems which come with this. Effect of material characteristics like size distribution, morphology and chemical composition in pneumatic conveying are discussed in detail. INTRODUCTION Fine materials like concentrate, feed mix, flue dusts (from waste heat boiler and electrostatic precipitator) and secondary dusts need to be conveyed at smelters. Both horizontal and vertical conveying and combination of those are used. For the purpose, both mechanical and pneumatic conveying are applied. Most widely used equipment are drag chain conveyors and various combination of pneumatic systems, e.g. air lift, dense-phase and dilute phase pneumatic conveyors. Kopar has 30 years’ experience with multiple material handling systems. Furthermore, acquired companies like Elmomet Oy and SITE Oy (both currently fully merged and integrated into Kopar Oy) have even longer history, especially in pneumatic conveying. It is commonly known that the finer the dust the more difficult it is to be conveyed pneumatically. Sulfates with crystal water makes it all even more difficult, if not ineligible. Challenges related to pneumatic conveying of various fine materials is discussed in detail in the paper.

Jan 1, 2019

By Vladimir Adamek

Due to the variability of subsidence characteristics across the U.S. coalfields, it was concluded that it would be practically impossible to develop a universal predictive model for mining-induced subsidence based on theoretical assumptions. Therefore, an effort was made to find a procedure to develop an empirical subsidence predictive model based on a sufficient Mount of field data from one mining area (in this case the Northern Appalachian Coal Region). It was also thought that this procedure, if successful, could be used as the template for developing predictive capabilities for other coalfields with different subsidence characteristics given a reasonable amount of field data. It has been found, in the Northern Appalachian Coal Region, that the variability of subsidence characteristics can be expressed by a polynomial equation developed through regression analysis of the variable subsidence coefficient and derived directly from the field data. In this study, field data were obtained from 11 Bureau longwall panel studies (16 half profiles) for static subsidence and 14 panels for dynamic subsidence. The effects of lithology, expressed in the form of a variable subsidence coefficient, have been separated for each test site by introducing a correlation between hythetically homogeneous overburden and existing ithological conditions. For each longwall panel, the characteristic of the variable subsidence coefficient was defined along individual static profiles. The definition of mean values with acceptable deviations is the substance of the predictive method. The same procedure was used to develop a predictive model for dynamic subsidence, since it war discovered that the rate of longwall face advance is not a necessary functional parameter. This paper presents the theory, development, and application of the static and dynamic subsidence prediction models.

Jan 1, 1986

By Hugh Rose

The properties of the company lie within the Pachuca district, State of Hidalgo, Mexico, connected by three railway lines with Mexico City, 55 miles southwest, and by two lines with Vera Cruz, 250 miles southeast. The ores were formerly divided by sorting into two classes, smelting and milling, the former averaging about 2 oz. gold and 335 oz. silver, the latter 0.12 oz. gold and 23 oz. silver. The smelting ores were sold to custom plants, principally the American Smelting & Refining Co. at Aguascalientes. The milling ores were treated by the patio process at the Guadalupe Hacienda, at Pachuca. This patio was probably the largest in existence at the time and continued in active operation up to March, 1910. In January, 1910, the mines and patio were sold to English interests, represented by Camp Bird Limited. Two new companies were formed, the Compania de Santa Gertrudis, S. A., to operate the mines, and the Compania Beneficiadora de Pachuca, S. A., to build and operate a custom cyanide milling plant. (Pig. 1.) Geology The historical geology of the Pachuca district is described by Fred J. Pope as follows: 1. The most recent sedimentary rooks in the district belong to the Cretaceous period. 2. The Cretaceous rocks were penetrated by large intrusions of ande-site (referred to hereafter as primary andesite) which occur as elliptical dome-shaped masses, striking northwest-southeast. These andesite bodies are from 8 to 15 miles long and from 6 to 10 miles wide. 3. A period of erosion during which the contour of the andesite domes was changed. 4. Andesite flows (secondary andesite) which rest unconformably on the primary andesites.

Jan 1, 1917

By G. Ladouceur

"Northgate Patino Mines produces a copper concentrate assaying 22% Cu. Moisture content of the product averages 12.5% H20 when dewatering with conventional vacuum disk filters. Recent smelter contracts reflect higher energy costs and increased moisutre penalty charges. This situation forced Northgate to investigate possible alternatives such as rotary dryers and pressure filtration. After laboratory testing, it was found that pressure filtration could economically meet the smelter&apos;s requirements. Pilot plant testing was then undertaken and confirmed the laboratory test results. Based on these findings, a model I.S.F. 1500 x 28/22 I.R. Lasta filter press was purchased.This paper describes the Copper Rand dewatering flowsheet the operation of the press and the economics of the process.IntroductionNorthgate Patino Inc. operates 2 und erground copper mines and a 3300 ton per day concentrator located in Chibougamau, Quebec, some 425 miles north of Montreal and 300 miles northeast of Noranda, Quebec.Copper concentrate is produced at a rate of 250 tons per day and trucked to a C.N. siding 10 miles from the mill site and then railed to the Noranda Smelter.Until early 1981, moisture content of the concentrate at 12.5070 was acceptable since smelter penalty charges were relatively low ($0.10 per ton per unit over 10070 moisutre). In January 1981, a new contract was signed in which the penalty was raised to $1.10 per ton per unit over 8070 . This and the increasing truck and rail costs justified the change-over from vacuum to pressure filtration . The Lasta I.S .F. 1500 x 28/22 was then purchased and installed, and operation of the plant started in December 1981."

Jan 1, 1987

By Robert J. Ramsey, N. Finkelstein, Robert D. Jeppson

Introduction The Utah Copper Division of Kennecott Copper Corporation will present its contribution to the A. M. Gaudin Flotation Symposium in four parts. The first two segments will discuss briefly the historical development of the Utah Copper Division and describe the ore body. The third segment will review the facilities and operating procedures for the Utah Copper Division&apos;s newest plant, the Bonneville Concentrator. The fourth segment describes flotation process circuitry, including molybdenite recovery procedures, at the Arthur and Magna concentrators. (The Bonneville concentrator is a crushing and grinding unit and does not have its own flotation plant). 1. Historical Development Progress through change has been the key to the successful operation of the Utah Copper Division for more than 70 years. As the original low-grade copper operation, the pressures to remain competitive and profitable have dictated the timing and extent of each change. Much of the change has occurred through technological and economic advancements in the area of ore concentration. The Utah Copper Division of Kennecott Copper Corporation had its origin June 4, 1903, when the Utah Copper Company was organized. The initial mill of 300 tpd (tons per day) capacity was built at Copperton a few miles below the present Bingham Canyon mine site. From 1904 to 1910, the Copperton mill was expanded to a capacity of 1,000 tpd, and all of the known apparatus for gravity concentration were evaluated.

Jan 1, 1976

By M. Kamali

In this research, in order to predict the peak particle velocity (PPV)(as vibration indicator) caused by blasting projects in the excavations of the Karoun III power plant and dam, three techniques including statistical, empirical, and neural network were used and their results were interpreted and compared. First, multivariate regression analysis (MVRA) was used as statistical approach. Next, PPV was predicted using some widely used empirical models. Lastly, an artificial neural network was used. In the ANN model, maximum charge per delay, total charge per round, distance from blast site, direction of firing, blasthole length, number of blastholes, total delay in milliseconds, number of delay intervals, and average specific charge were taken into consideration as input parameters and consequently the PPV as output parameter. The results of the techniques were interpreted from two points of view. Firstly, the correlation between the observed data and predicted ones, secondly the total error between observed data and predicted ones. The MVRA had a satisfactory correlation but its error of estimation was comparatively very high. The empirical model had reliable correlation and a small error of estimation; in total the results of empirical method were more reliable than those of MVRA. Generally, the ANN approach showed very high correlation and a very small error. The results of this research indicated that the ANN model is the best predicting model for PPV in comparison with other approaches.

Jan 1, 2010