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By J. E. Wilson, F. P. Lebar

TYPICAL of southwestern Wyoming are coal structures that dip from 4° to 17°. Those at the Reliance and Winton mines of the Union Pacific Coal Co. average 91/2 and 15°, respectively, and dip almost directly to the west. By using a hoist for main haulage, a method of mining has been developed whereby four places are driven on the pitch of the seam. Slightly upgrade from the strike at intervals of 300 ft. a two-entry system is driven to the north and to the south. In both fields, which are nongassy, an exhaust system of ventilation is used. Slope, manway, and lower or haulage entries are placed on the intake. Two air-courses and upper or breaking entries are placed on the return. A retreat system of mining is used exclusively. From the opening of its first coal mine, in the year 1868, until 1923, when the Joy loader was introduced into The Union Pacific Coal Company's operations, slopes, aircourses, and manways were driven by hand. Development work was conducted on the night shift or off shift, and the main hoist was used for haulage. The slope was driven down the pitch with crosscuts turned at intervals of 75 ft. and entries at 230 ft. Connections to the workings above were made by driving up the pitch 75 ft. Mule haulage was used to furnish cars to the hand loaders driving planes up the pitch. A sheave was placed at the upper end of the plane with a I¼-in. hemp rope dou- bling around it and having me end attached to the car and the other to the mule. The mulc, pulling downhill, raised to the working face a small wooden car of one-ton capacity. With the introduction of mechanical loading, mule haulage and small wooden cars disappeared and auxiliary hoists, electric locomotives, and steel cars with a capacity of 4 tons were substituted. Development work was done during two or three shifts of the 24-hr. period. Driving of haulage slopes down the pitch by hand continued. Crosscut intervals were changed to Ioo ft. and the crosscuts were driven with shaking conveyors. With the use of a 90° swivel, aircourses and manways were driven up the pitch. Entry intervals were changed from 230 ft. to 300 feet. Mobile loaders of the Joy type were first employed by this company in the year 1923 at Hanna, Wyo. In 1928, two 4-BU Joy loaders were transferred from Hanna to Winton No. I mine, where they were used in driving slopes and entries. At the time of transfer, the machines were about worn out and were not suited to the steep pitch, therefore frequent breakdowns occurred. However, experience justified the belief that the mobile loader was fitted for driving slopes in this field. During 1937 a unit was installed in Reliance No. 7 mine, and during 1939 one in Winton No. 7½ mine. With the introduction of this type of mobile loader, a new era of development began. Speed undreamed of in hand-loading days became a reality. The cost of development down the pitch has been reduced until it

Jan 1, 1942

By William Romanoff

As you are no doubt aware, virgin copper, tin and zinc are strategic metals today and little, if any, is available in the manufacture of brass and bronze ingots. The source of alloys used in the manufacture of implements of war are, therefore, from ingots made almost exclusively from the refining of available scrap. It is amusing to hear of the unusual physical properties of various alloys because they mean nothing unless the foundryman is able to produce good sound castings. A discussion therefore means little unless something is said about the relative foundry practice of the various brass and bronze alloys. Among the factors that influence the quality of castings produced are: (I) improper gating, (2) wrong pouring temperatures, (3) poor sand conditions, (4) bad melting practice, (5) faulty pattern equipment, (6) metal. It seems to be the consensus of opinion that about 95 per cent of defective castings can be attributed to the first three causes. Metal while a source of trouble at times, is probably the least responsible for foundry difficulties. As you all know, a molder never makes a bad mold. His first thought is to blame the "damn" metal as this is the path of least resistance. Let's consider the "damn" metal and how it is made. In the manufacture of ingot, scrap is collected at various sources and is shipped to the refiner either directly or through the scrap dealer or broker. It is usually sorted, although not always. In working toward a specified chemical analysis, weighed amounts of various grades of scrap metals are melted, and then oxidized by blowing air through the molten bath to rid the metal of detrimental impurities or reduce them to within specification limits after which they are skimmed off with the aid of some added fluxes. The bath of metal is then poled with wood, chemically tested before and after additions are made to bring it up to required specifications, and finally poured into ingots, after which a final chemical and physical examination is made. As mentioned above, the impurities have either been eliminated or reduced to the point where they meet the specifications of all Government Agencies or Technical Societies. Therefore, why worry as to whether castings are to be made from strategic virgin metals or secondary ingots from available scrap. Tests have continuously been made and it is a proven fact that physical properties are no better from virgin metals than from secondary ingots of the same composition. It is definitely more fool-proof to make castings from secondary ingot than from virgin metals because if an error is made in weighing the copper, tin, lead or zinc of the latter, the analysis is off balance. If an error is made in weighing secondary ingots, you merely have more or less metal but the analysis remains constant. Furthermore, when pure copper is melted. it is susceptible to reducing gases as well as oxides. Secondary ingot is already alloyed with natural deoxidizers such as zinc, etc., and the volatilization of the latter also tends to eliminate reducing gases mechanically.

Jan 1, 1943

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 W. R. Opie, O. W. Mole

By confining the electrolytic reduction of TiCl4 to the interior of a porous basket-cathode the electrolyte between the anode and the cathode can be kept free of reduced chlorides of titanium eliminating undesirable reactions of reduced titanium chlorides with brickwork or oxygen in the cell atmosphere. Semiwalls are not needed. Operating conditions are given and discussed. Titanium yields based on TiCl4 introduced are 93 pct. Cell voltage and amperage characteristics are adaptable to multiple cathode-series cell urmngements. EXPERIMENTS conducted in the National Lead Co., Titanium Division, Research Laboratories showed that titanium metal could be deposited in an electrolytic cell in which titanium tetrachloride was introduced through a tubular cathode.' The reduced chlorides as formed were confined to the vicinity of the cathode. They were reduced to form a cocoon-shaped growth of sponge titanium metal at the end of the cathodic tube below the surface of a sodium chloride-strontium chloride electrolyte. No diaphragm was needed.' Although high-purity, coarse crystalline metal was produced in the cocoon growth, the yields of metal removed from the cell, based on the amount of titanium introduced as TiCl4, were usually only 40 to 60 pct. The growth of the cocoon deposits was away from the cathodic tube for introduction of TiC14 toward the anode following the flu lines. The reduced chlorides of titanium diffused through the deposit and were reduced to titanium electrolytically within the deposit or on the outer surfaces. Deposition within the pores of the cocoon was detrimental to further deposition because it plugged the passages used by the reduced chlorides to diffuse away from the point of titanium introduction. As plugging due to deposition in the pores proceeded, the reduced chlorides in the center of the cocoon would force their way out by forming a channel and would escape from the vicinity of the cathode without being reduced to metal. Some work had been done previously1, 2 which indicated that if the end of the cathodic tube for introducing TiC14 was surrounded by a screen connected to the tube so that it too was cathodic, growth could be made to occur within the screen. Direction of growth was observed to be from the screen toward the point of TiCl4 introduction. Through a series of design modifications a cathode evolved which consisted of an introduction tube surrounded by a perforated metal basket as shown in Fig. 1. Modifications of this design are described in Refs. 3, 4, and 5. This basic cathode designa" was incorporated in electrolytic cells containing graphite anodes and constructed of high-silica brick refractory for containing the electrolyte (see Fig. 2). The inner crucible was surrounded by insulating brick and the entire unit contained in a welded steel shell. Suitable cooling chambers with inert gas atmospheres were devised so that the cathode could be removed from the cell without exposing the deposit to air while hot. Various alkali and alkaline earth chlorides were used as electrolytes.

Jan 1, 1961

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 H. M. Giegerich

About Cominco Only the original Broken Hill zinc deposit in Australia is larger than Cominco's Red Dog deposit. Indeed, the third largest zinc deposit in the world is Cominco's own Sullivan mine in British Columbia. The production plan at Red Dog is based on mining 1.9 Mt/a (2.1 million stpy) or ore, to produce 635 kt (700,000 st) of zinc and lead concentrates. At full production, Red Dog will be the largest zinc mine in the Western world. It will also be one of the lowest cost producers in the Western world. And Red Dog will be one of the first major hardrock mines to operate in Alaska in modern times. The Red Dog project will be the third and largest of Cominco's Arctic mining developments in the past 13 years. The company's Black Angel mine in Greenland started up in 1973. And the Polaris mine, high in the Canadian Arctic, began production in 1982. Red Dog is one of the world's great mineral deposits. It is a unique combination of large size, high grade, and a waste-to-ore ratio less than 1:1. The present mineral reserve (Table 1) is estimated to be 77 Mt (85 million st), at a grade of 17.1% zinc, 5% lead, and 82 g/t (2.4 oz per st) of silver. Mine life is estimated at nearly 50 years. Cominco is the largest producer of lead and zinc in the Western world. Cominco's smelter in southern British Columbia is the largest lead and zinc plant in the west. Cominco also produces copper, steel, tin, gold and silver, along with potash, ammonia, urea, and phosphate fertilizers. But Cominco's major products are lead and zinc. At present, Cominco produces more than 10% of the Western world's needs for these metals. Background-Red Dog The Red Dog mineral deposit is located in northwestern Alaska, about 960 km (600 miles) by air from Anchorage, Alaska's largest city. The Red Dog region is home to more than 5000 people living in 11 villages. The largest village is Kotzebue, with a population of about 2500. The village is located about 144 km (90 miles) southwest of the mine. The country around Red Dog features rolling hills, leading into mountains to the north and east. The deposit, at an elevation of 305 m (1000 ft), is in permafrost. But the area is not nearly as cold as that encountered in the Canadian Arctic, or on Alaska's North Slope. At Red Dog, the permafrost is relatively fragile. It can be thawed fairly easily, a factor in engineering designs. Red Dog Creek flows through the mineral deposit. One of the natural phenomena at Red Dog is the presence of significant heavy metal contamination in the creek. This has eliminated all fish in the vicinity of the deposit. The contamination, and resulting mineral stain, led to the discovery of Red Dog. The staining is caused by rain and snowmelt. This leaches the heavy metals from the mineral zone, which comes right to the surface in many places. As mining progresses, this surface mantle will be removed. This will

Jan 12, 1986

By Armine F. Banfield

Selection of a plant site is becoming increasingly more difficult and important, mainly due to advances in technology, antipollution laws and changing economic, social and community conditions. Discussion falls naturally under three headings: economic considerations, terrain considerations and environmental considerations. Aside from acquisition of surface rights, problems having to do with terrain, or topography, usually can be dealt with by management alone. Economic considerations largely involve decision-making by management plus negotiations with government regulatory agencies acting under set rules. Environmental considerations, such as air and water pollution, townsites, housing, schools, noise abatement, appearance, surface rehabilitation and the like usually are under the control of federal, state or municipal bodies with a large degree of decision-making. Successful solution of environmental problems often requires lengthy negotiations, during which public relations by the mining company play an important part. Existing union regulations or customs may confine decision- making and the establishment of wage rates, working conditions and regulations. These usually require considerable planning and negotiation. Initial production often is delayed by formal negotiations involving environmental considerations. Surface facilities usually can be divided into two broad categories: those for concentrating ores--such as gold, silver, base metals, low-grade iron ore and potash-which should be close to the mine to reduce ore-transportation costs to a minimum, and those where the ore is shipped elsewhere for treatment, such as direct-shipping iron ores, complex high-value ores requiring specialized treatment and ores from which special products are derived for local markets. The long-term aspects of plant-site selection too often are given no or insufficient consideration because of immediate financial limitations. As a result, additional capital expenditures may be required later for relocation of townsites, roads, railroads and plants to provide additional room for extension of open-pit operations, tailings disposal, storage, or danger to foundations through underground subsidence or open-pit mining. In the past, hillside sites for concentrating plants were preferred to take advantage of gravity-feed methods The modern trend, however, is to place as much equipment as possible on approximately the same level. Some costs relating to plant sites are included in this section. They are ranges as of the year 1970, and are offered only as general guides. Such costs will vary widely with geographical location and, of course, are subject to escalation due to increases in wage rates and cost of materials. Items in the bibliography contain more detailed and comprehensive data on specific subjects. Most contain bibliographies of their own.

Jan 1, 1973

By S. Timbillah, R. LaDouceur, C. A. Young

A low carbon-chain trithiocarbonate ((NaOOC)CH2SCSSNa), known as Orfom® D8, was tested as an organic depressant in the differential flotation separation of chalcopyrite and molybdenite in the presence of sodium isopropyl xanthate (SIPX) or potassium ethyl xanthate (KEX). Flotation studies show that Orfom® D8 depressed chalcopyrite while molybdenite remained floatable. A study of the mechanism of adsorption indicated that Orfom® D8 adsorbs on chalcopyrite through its trithiol (CS -) functional group via both chemisorption and Orfom® D8-metal complexation. The Orfom® D8 depressant is selective and does not form any Orfom® D8-molybdenite bonds. Thus, the combination of the depressant in the presence of either collectors depressed the chalcopyrite and floated the molybdenite. Fundamental parameters of the Orfom® D8 including pKa determination as well as possible decomposition products of the depressant are reported. This study includes an economic case illustrating why using Orfom® D8 can be beneficial. INTRODUCTION Depressants are incorporated in mineral flotation units in the minerals processing industry to make or keep certain minerals selectively hydrophilic. Most reagents used are inorganic and include, for instance, sodium hydrosulphide (NaHS) which is the major reagent used for depressing sulphide minerals, particularly those containing copper, iron and lead. However, the issues with its use include high toxicity, tendency to explode, strong odour, and high consumption due to required high dosages as well as its tendency to oxidise in air. The latter can be overcome by conducting flotation in inert atmosphere such as nitrogen and recycling of the flotation process water. The use of more novel depressants that are essentially organic including homopolar reagents with hydrophilic functional groups on both ends of its structure adsorb onto a target mineral surface with one polar end, the other polar end orients into the bulk solution making the surface hydrophilic and thus, ensuring that the desired mineral does not float. The stringent environmental and safety regulations being implemented in recent times mean that the search for more environmentally friendly chemical reagents is a must have. The use of organic depressants for flotation systems such as Cu-Mo sulfide bulk concentrates has become increasingly important because of the health, safety and environmental issues surrounding the use of inorganic depressants. Nagaraj et al. (1983) investigated the Cu-Mo sulphide flotation system using a structure-activity study of depressants. They found that a major advantage of organic depressants is the ability to tailor-make organic molecules to suit many requirements and thus, eliminate the guess work approach that was predominant in the industry.

Jan 1, 2019

By Roswell P. Angier, Herbert S. Kalish, Henry H. Hausner

POWDER metallurgical methods as applied to zirconium are of great interest because they permit not only the fabrication of parts directly to shape with a minimum loss of material but also the utilization of scrap, and the metal may be processed at temperatures well below the melting point. There are two difficulties involved in the melting of zirconium: 1—High vacuum is necessary for melting, and 2—zirconium reduces nearly all oxides so that it is difficult to find a suitable crucible material. The purest material, zirconium crystal bar, is expensive, and the methods of subsequent fabrication are limited. Cold rolling or extrusion of zirconium crystal bar is suitable for making certain shaped parts, but where melting is involved, the purity of the metal is invariably decreased, and a method of casting to a definite shape does not seem to be feasible. Where it is necessary to hot roll, zirconium must be covered with a steel or copper jacket to protect it from the air. This is necessarily a costly operation. Where annealing of zirconium is necessary it must be conducted in a high vacuum. In an investigation of powder metallurgical methods for zirconium, the zirconium powder used should be characterized by the following properties: 1— Highest degree of purity, 2—uniformity in particle shape and particle size distribution, 3—low compression ratio, and 4—satisfactory strength of the compacted powder (green strength). Zirconium powders have been made by two methods: By the reduction of zirconium tetrachloride (ZrC1,) with sodium,2' *•B "nd by the reduction of zirconium oxide with calcium and calcium chloride." The first method represents one of the earliest attempts to make pure zirconium metal, and the sintered product from this powder is extremely brittle. The second method of making zirconium powder is far more successful, but the powder made in this way has an extremely low apparent density and, therefore, a high compression ratio which makes it difficult to use it for powder metallurgical purposes. For these reasons the above-mentioned zirconium powders were not used in this investigation. To obtain sintered zirconium compacts of highest 'purity and good ductility, it was first necessary to produce a suitable powder. Pure zirconium has been made chiefly by two methods. The iodide method, from which the product is "crystal bar" zirconium, was developed first.'." By this purification process very soft, pure, and ductile zirconium metal can be produced. The second method, developed by the U. S. Bureau of Mines, is the reduction of zirconium chloride with magnesium." The sponge zirconium thus produced is also soft, pure, and ductile.

Jan 1, 1952

In process industries, time-dependent random fluctuations of raw material characteristics and mass flow rates in bulk material or pulp flows cannot be always prevented for different reasons, eg because of the geological situation (no uniform formation of the deposit) or/and because of technological reasons (segregation in bunkers or on stockpiles). Since a long time, there has been practical experiences that such fluctuations can have negative consequences, eg: ò fluctuations of the final products characteristics, ò higher expenditure for process control, and ò worsening of technological characteristics (recovery, sharpness of separation etc). In the literature, there can be found a lot of details as to, eg the improvement of the sharpness of separation in a jig in dependence on the raw material blending (Kubitza, 1981; Gathen, 1983) or the decreasing flotation recovery with increasing raw material characteristics fluctuations in iron and potash ore processing plants (Bubnova, 1978). Up to now, however, these consequences of random fluctuations, eg on the separation efficiency were not predictable theoretically. It could be shown by means of the general model for mechanical macroprocesses (Graichen, 1989), that the reduction of input fluctuations passing a blending facility can be calculated in this way. In this example, the steady state of the process is the wanted one, but, in general, the model is able to describe also non-steady states of the process regarded. On the other hand, the general model of mechanical macroprocesses has been used successfully for modelling separation processes, eg cross-flow and air classification (Molerus, 1969; Neesse, 1991; Senden, 1979; Tichonov, 1973). That is why it was tried to model the influence of input fluctuations on the separation efficiency, that means, the impact of non-steady states - as a consequence of random fluctuations - on a separation process. The models discussed in the literature are related nearly exclusive to the steady state of the process. The reasons for that lie in the (relative) simplicity of the models and the fact that the steady state is aimed at by means of blending processes under practical conditions - because of the advantages mentioned above. These models cited have proved to be superior to other models, since the basic general model for mechanical macropocesses makes possible the consideration of turbulence acting in the process unit and influencing the separation process essentially. This is valid also for modelling the turbulent cross flow classification, applied to mechanical classifiers and hydrocyclones, where the correspondence between the separation characteristics (cut size, separation sharpness) calculated and determined experimentally is much better than for other models.

Jan 1, 1993

By G. L. Boxer, R. A. Metz, C. B. Smith, W. J. Atkinson

Since the inception of the Ashton (formerly Kalumburu) Joint Venture in 1972 the parties have expended $24 million in exploration for diamonds in the Kimberley region. This has resulted in the discovery of the currently subeconomic Ellendale deposits in 1976/77 and the economic Argyle - Smoke Creek Deposits in 1979/80. In addition to Argyle and Ellendale, some 90 kimberlitic and lamproitic bodies have been discovered grouped broadly into three provinces, on the west, east and north margins of the Kimberley block. These bodies range in size from dyke-like features less than 1. m (3 ft) in width to pipes with a surface area of 128 ha (316 acres). The bodies with larger surface area are volcanic crater deposits, infilled with air-fall and water deposited tuffs and epiclastic sediments. Diamond content ranges from trace to economic concentrations. The exploration discoveries were facilitated by early recognition that the petrography, mineralogy and chemistry of the diatremes varied from the classical kimberlites of Kimberley, South Africa, to unusual more highly fractionated, alkaline, silicic varieties having affinities with leucitelamproite. Indirect methods of exploration were used, primary reconnaissance work being stream gravel sampling for kimberlitic indicator minerals. Evaluation showed Ellendale diamonds are 90% gem and near gem, but the low grades and tonnage render the deposits unpayable at today's diamond prices. After discovery in late 1979 and completion of a feasibility study by the end of 1982, evaluation studies proved Argyle to be economic. Alluvial diamond mining commenced in 1983 producing 6.2 million carats (Mct) that year, 60% being of industrial quality. Open-pit mining of the AK1 lamproite should start in 1985/1986 at a rate of 3 million t/a (33 million stpy) and produce 25 Mct/a, making this mine, and Australia, the world's largest producer of diamonds by weight. The revenue from diamond deposits is a function of the grade and the value of the diamonds. The grade is the product of the frequency of occurrence of the diamonds (stone density) and the mean stone size; the value of the stones is strongly related to their size distribution. During evaluation of the West Australian deposits, the diamond distribution was found to vary markedly in stone density and stone size. These variations corresponded with changes in the rock type hosting the mineralization. As a consequence, geological mapping of individual lamproites within the multiple vents and careful monitoring of the diamond distribution became an important part of the evaluation work. Similar variations in stone size and frequency were found to accord with separate alluvial terraces and fan gravels derived by erosion from the Argyle AM lamproite.

Jan 1, 1992

By G. L. Boxer, C. B. Smith, W. J. Atkinson

Since the inception of the Ashton (formerly Kalumburu) Joint Venture in 1972 the parties have expended $24 million in exploration for diamonds in the Kimberley region. This has resulted in the discovery of the currently subeconomic Ellendale deposits in 1976/77 and the economic Argyle - Smoke Creek Deposits in 1979/80. In addition to Argyle and Ellendale, some 90 kimberlitic and lamproitic bodies have been discovered grouped broadly into three provinces, on the west, east and north margins of the Kimberley block. These bodies range in size from dyke-like features less than 1 m (3 ft) in width to pipes with a surface area of 128 ha (316 acres). The bodies with larger surface area are volcanic crater deposits, infilled with air-fall and water deposited tuffs and epiclastic sediments. Diamond content ranges from trace to economic concentrations. The exploration discoveries were facilitated by early recognition that the petrography, mineralogy and chemistry of the diatremes varied from the classical kimberlites of Kimberley, South Africa, to unusual more highly fractionated, alkaline, silicic varieties having affinities with leucite-lamproite. Indirect methods of exploration were used, primary reconnaissance work being stream gravel sampling for kimberlitic indicator minerals. Evaluation showed Ellendale diamonds are 90% gem and near gem, but the low grades and tonnage render the deposits unpayable at today's diamond prices. After discovery in late 1979 and completion of a feasibility study by the end of 1982, evaluation studies proved Argyle to be economic. Alluvial diamond mining commenced in 1983 producing 6.2 million carats (Mct) that year, 60% being of industrial quality. Open-pit mining of the AK1 lamproite should start in 198511986 at a rate of 3 million t/a (33 million stpy) and produce 25 Mct/a, making this mine, and Australia, the world's largest producer of diamonds by weight. The revenue from diamond deposits is a function of the grade and the value of the diamonds. The grade is the product of the frequency of occurrence of the diamonds (stone density) and the mean stone size; the value of the stones is strongly related to their size distribution. During evaluation of the West Australian deposits, the diamond distribution was found to vary markedly in stone density and stone size. These variations corresponded with changes in the rock type hosting the mineralization. As a consequence, geological mapping of individual lamproites within the multiple vents and careful monitoring of the diamond distribution became an important part of the evaluation work. Similar variations in stone size and frequency were found to accord with separate alluvial terraces and fan gravels derived by erosion from the Argyle AK1 lamproite.

Jan 1, 1985

By W. E. Hess

AT the Vesta mines of Jones & Laughlin Steel Corp. on the Monongahela River, 35 miles south of Pittsburgh, JCM Joy continuous miners and 6-SC shuttle cars are used for pillar extraction in the Pittsburgh seam. The entire output of the mine goes into coke production for blast furnace operation. Because this coal, which has high metallurgical qualities, is close to the place where it is ultimately used and because supplies with these properties are being rapidly depleted, complete recovery is essential wherever possible. With the introduction of the continuous miner into the Vesta mines, therefore, it was necessary to plan a project which would facilitate development and provide for future total recovery. Use of continuous miners under varying conditions has necessitated changes with regard to roof action, method of development, method of pillar attack, and innovations in mining. In the following pages these changes will be discussed. Prior to the introduction of continuous miners the Pittsburgh seam at the Vesta mines averaged 72 in. of mineable coal, overlain with 12 to 16 in. of draw-slate, which weathers on exposure to air and is not firm enough to be held in place. Immediately overlying the drawslate is a wild or rooster coal, very irregular in thickness, varying from 12 to 30 in. Above the rooster coal the strata consists predominately of shales or lime shales and sandstone. Just before the introduction of continuous miners in the Vesta mines, however, a definite change took place in the character of the roof structure overlying the coal seam, see Fig. 1. Thickness of laminated roof immediately overlying the rooster coal varies with localities and introduces a problem in roof control. In fact, a set of eight main entries being developed in Vesta NO. 4 mine entered an area with so difficult a roof structure that management considered abandoning the section altogether. In this locality the roof consisted of thin alternate layers of roof coals and shales, with an occasional irregular seam of slick slate, the whole being laced with slips, and the coal itself being cut by clay veins and spars. In an effort to overcome some of the difficulties, roof bolting was introduced in 1949. This made it possible to recover coal from an area which otherwise would have been economically unmineable. There have been very few roof bolt failures. When conventional machines were replaced by continuous miners there were changes in the development of butt entries. By these changes it was hoped: 1—to speed development, 2—to provide as much solid coal as possible for the machine on retreat, 3—to avoid the necessity of crossing entries already driven, 4—to facilitate ventilation, 5—to provide a roadway for shuttle car haulage which would eliminate the necessity of running through doors or curtains, 6—to provide better roof control by increasing the distance between the center lines of entries, and 7—to provide more coal when the machine was in position to load.

Jan 1, 1956

By M. E. Wadsworth, R. C. Peterson, W. M. Fassell

The temperature and pressure dependence of the reaction of tantalum in oxygen were investigated from 500° to 1000°C at pressures from 10 mm Hg to 600 psi total oxygen pressure. Tantalum was found to oxidize linearly under the above conditions. Three distinct regions of temperature dependence were found with different energies of activation. From 500° to 600°C the rate of oxidation of tantalum was found to be essentially independent of the oxygen pressure at the pressure investigated. The oxidation rate increases rapidly with an increase in pressure from 600' to 800°C. The dependence of the oxidation rate on the bulk concentration may be expressed by V = k'?, where k' is the specific rate constant and ?=k1Co2/(l + K1Co2), where k, is the equilibrium constant for the adsorption of oxygen on tantalum. ONLY a limited amount of data has been presented in the literature on the oxidation behavior of tantalum metal. Some values of its oxidation rate were reported in the Corroston Handbook' and by McAdams and Geil.' The first extensive investigation of tantalum was the work of Gulbransen and Andrews." They determined the reaction rates of tantalum in O2 and N2. Their results indicate that this metal follows the parabolic law in the temperature range of 250" to 450°C in O2 at pressures of 7.6 cm Hg with an activation energy of 27,400 cal per mol. In nitrogen, the metal follows the parabolic law from 500" to 800°C at 7.6 cm pressure. All experimental determinations were of 2 hr or less duration and deviations from the parabolic rate conceivably could occur at longer times. A spectrophotometric study of the oxidation of tantalum by Waber, Sturdy, Wise, and Tipton' indicated good agreement between this method and the microbalance technique. The experiments were conducted in air from 220" to 350°C. The metal was found to follow the logarithmic law up to 320°C and the parabolic law at higher temperatures. The activation energies in the logarithmic region were reported as 12.57 kcal per mol and 12.0 kcal per mol, respectively, for the microgravimetric and spectrophotometric methods. In the parabolic region the activation energy was found to be 27.2 kcal per mol, in good agreement with the value of Gulbransen and Andrews." The present paper is concerned with the behavior of tantalum from 500" to 1000°C in oxygen at pressures ranging from 1/2 to 40.8 atm pressure. Earlier preliminary data by McKewan5,6 indicated the following facts concerning the behavior of tantalum: 1—At temperatures from 500" to 600°C, the metal followed the linear law. 2—At 575°C the rate of oxidation was found to increase with O2 pressure. 3—The Arrhenius plot in this temperature region was nonlinear. The nonlinearity of the Arrhenius plot was indicative of a more complex behavior than originally suspected for this metal.

Jan 1, 1955

By Laxman M. Amaratunga, David N. Yaschyshyn

"In the mining industry several types of fills are used for backfill operations. They are broadly categorized into three basic types: rock fill, hydraulic/slurry fill, and paste fill. Recently, paste fill has become the popular alternative for backfilling because of its many environmental and operational advantages including its ability to use total tailings, both fine and coarse fractions.Amaratunga and Annor (1989) first presented the concept of agglomeration of fine tailings, which are otherwise excluded from backfilling, for utilization in backfill as a readily available supplementary fill material. Previous experimental work investigated the use of agglomerated fine gold mill tailings in low density slurry fills (60 -70% pulp density), consisting of varying agglomerate and tailings compositions.This paper presents the ongoing research into utilizing these agglomerates in a high density paste fill (70 -85% pulp density). Various binder dosages, curing period and combinations of pellet to tailings ratios were studied. Slump of the raw fill and compressive strength of the cured fill were examined. Results from the above study indicate that agglomerated tailings paste fill (ATPF) has superior strength characteristics when compared to total tailings paste fills.IntroductionThe Canadian mining industry produces more than 500 million tonnes of tailings and waste rock annually. These tailings consist of a coarse fraction and a fine fraction and can be either reactive (acid generating) or non-reactive. Most mining operations utilize the coarse fraction for backfilling while the fines have been traditionally disposed of in surface tailings ponds. Recently, paste fill has been developed which uses a fraction of the fine tailings (10 - 30% w/w of >45 µ,m) in the fill. This development allows a mining company to use a portion of the fines for backfill, but still requires surface impoundment of the unused portion of fine tailings. Reactive tailings have the added problem of generating acid after being oxidized in the presence of air and water. Treating acid mine drainage (AMD) seepage from these ponds has proven to be a costly environmental problem and long term liabilty.One practical method to reduce the environmental impacts associated with the tailings pond and to modify the physical and chemical properties of the fine mill tailings so that they can be utilized and disposed of in an environmentally safe manner through surface or underground applications is through agglomeration using the Cold-Bond Tailings Agglomeration technology (CBTA). Amaratunga and Annor (1989) first presented the concept of agglomeration of fine tailings for use as coarse aggregates for underground backfill, gravel or surface covering material. CBT A reduces the necessity of impounding tailings on surface while providing a suitable coarse aggregate resource material for practical use. Although the CBT A technology has been mainly limited to pelletization agglomeration alone, other forms of agglomeration like pressure agglomeration which include the techniques of briquetting and extrusion have also proven successful (Amaratunga, 1995)."

Jan 1, 1996

By James J. Wise

INTRODUCTION One basic design feature utilized in all caving and sublevel stoping techniques is the drawpoint. These draw- points may be located on sublevels when using sublevel caving methods, or may be located at the floor pillar when using block caving, sublevel stoping, or vertical crater retreat methods. How the drawpoint is designed and developed, how the ore is withdrawn from it, and how the ore is hauled to the shaft or portal encompasses the loading and hauling function. Historically, predominate loading and hauling equipment for caving and sublevel stoping has been scraper slushes, air powered overshot loaders, and rail mounted train haulage. All of this equipment has two common constraints. 1. The equipment locked into specific work sites and is not easily shifted to another work site. This constraint can often lead to low utilization and low productivity of the equipment. 2. The equipment is designed to work in the horizontal plane, which requires that the various horizontally-oriented working levels be connected by vertical raises. Vertical development by means of raises is not only increasingly expensive, but also is not well adapted to movement of large equipment into the s topes. In the last 20 years, two revolutionary labor saving concepts have become common in caving and sublevel stoping. 1. Long-hole drilling, which has mechanized the development of the drawpoint extraction structures as well as the actual stoping and undercutting. 2. Vertical access by means of trackless ramps between levels, and the consequent use of rubber-tired vehicles not only for loading and hauling ore, but also for the rapid movement of men, material, and equipment into the stopes. The introduction of trackless mining systems for loading and hauling has not been easy. Mine planners have had to re-orient their thinking to the larger operating space required by trackless equipment. Larger access drifts have necessitated reconsideration of mine layouts to insure proper ground support. Spiral ramp developments were often laid out on slopes steeper than the equipment could negotiate economically. Mine operators had to accept the added responsibility of roadway maintenance, water removal, increased ventilation, and more technical mechanical maintenance. As the mine operators enthusiastically embraced the superior productivity of trackless mining systems, they some- times disregarded opportunities to combine the older types of equipment with trackless vehicles to maximize productivity at lowest cost. In order to gain a proper perspective on this problem, it is necessary to review the various techniques of loading and hauling and to become more aware of the advantages or disadvantages that each system or combination contributes. LOADING Loading consists of two functions, (1) primary loading from a drawpoint and (2) secondary loading into long- haul system, a truck, train, or possibly a belt conveyor. Mine planners have four basic loading systems available to fit various operating conditions. These are, (1) gravity, (2) scrapers, (3) rail-mounted loaders, and (4) trackless loaders. Each system can produce optimum results under specific conditions. Gravity Loading Gravity loading is probably the oldest system in general use today.

Jan 1, 1981

By Jacques H. Nantel, James F. Archibald

INTRODUCTION The use of backfill materials in underground mines has been practiced for many years and the trend indicates an increased utilization of backfill materials in the future. The authors believe that underground uranium mines will follow this trend. Classified mill tailings, approximately 65 to 70 percent solids by weight, are the most common fill materials. Cement or other pozzolanic agents can be added to the fill materials to improve their strength characteristics. Dry fill, in the form of development muck, can also be introduced in the stopes to fill the voids left by the action of mining the ore. This paper deals mainly with the assessment of the radiation properties of backfill materials originating from uranium ores processed in the milling facilities of uranium mines. The paper outlines the results of studies conducted in the laboratory and in the field on radon emanating characteristics of backfill materials utilized in uranium mines. ROLE OF BACKFILL IN UNDERGROUND MINES Fill is a structural component of any underground mining system. In general terms, the purpose of filling is to provide ground support, to make recovery of pillars possible to provide a working platform for men and equipment, and to allow for waste disposal. There are other inherent advantages to the use of backfill underground, the most important being to: a) permit mining at greater depth b) allow the adoption of highly mechanized mining techniques c) reduce dilution from wall sloughing d) minimize roof and wall failures e) allow the mining of irregular orebodies, and f) to provide better control of ventilation. The last item is of great importance to uranium mine operators. The introduction of radioactive backfill underground represents an additional source of radioactive contamination; on the other hand the introduction of backfill in the mine reduces the volumes of workings to be ventilated. As a consequence the general mine ventilation is more effective and in most cases, better air quality can be maintained in the mine. There are of course some disadvantages associated with the use of hydraulic fills in underground mines, notably: a) the handling of large additional quantities of water from the underground workings b) delays caused by the cyclical nature of the backfilling operations c) additional costs associated with stope preparation and placement of backfill in stopes, and d) "spills" caused by bursting pipes or fill escaping from fill retaining structures. ASSESSMENT OF NON-RADIOACTIVE PROPERTIES Although the radioactive properties of mine backfills are discussed in this paper many other physical and chemical properties of backfills must also be considered. These properties include: a) percolation rate and permeability parameters b) viscosity versus pulp density relationships c) size analyses d) settling times e) settled bulk density f) solids specific gravity g) void ratio and porosity h) optimal percentage solids by weight i) specific gravity of slurry j) addition of cement:cement retention and loss k) unconfined compressive strength 1) triaxial compressive strength m) shear strength (non-cemented)

Jan 1, 1981

By R. N. Orava, L. A. Rice

One of the difficulties encountered in the preparation of titanium specimens for transmission electron microscopy is the formation of a surface hydride phase during the thinning operation at ambient temperatures. This necessarily destroys contrast, obscures features of interest, and often leads to misleading results. By cooling the polishing solution to temperatures below 0°C this problem has been overcome in the past.' However, since the cooling procedure is somewhat inconvenient, a technique was developed at this laboratory for thinning titanium at room temperature. The material examined was A-70 grade commercial-purity titanium purchased from the Crucible Steel Co. of America. The mechanical and thermal treatments, described elsewhere,* yielded a recrystallized grain size of 4 to 5 p. A slice, 0.5 mm thick and 1.5 mm wide, was cut from a bulk specimen2 using a diamond cut-off wheel. The sample was then coated with paraffin wax. Two concentric holes, about 0.03 in. in diam, one at each broad face, were scribed through the wax to the surface of the metal. Subsequent polishing for several minutes resulted in a preferentially thinned or "dimpled" region. Thereafter, the wax was removed and the entire specimen polished to the point of breakthrough at the base of the dimple. Polishing was effected chemically in a solution comprised of 30 pct HF and 70 pct HN03 by volume, kept at room temperature (20" to 25OC). The specimens were mounted in a stereo holder and examined at 100 kv in a Philips l00 B Electron Microscope. A micrograph from a foil prepared in this way is shown in Fig. 1. Several points are noteworthy. This electron micrograph shows neither evidence for the formation of a hydride phase due to specimen preparation nor, moreover, any obvious surface contamination in electron-transparent areas of a freshly prepared foil. Localized variations in foil thickness observed in some specimens seem to be a characteristic of this chemical polish but do not seriously detract from foil quality. With care, the lifetime of a good specimen exceeded 2 hr in the beam. On the other hand, the exposure of a sample to air at atmospheric pressure, when still warm from the beam, clearly resulted in the deterioration of a formerly clean area just previously in the field of view. An example in Fig. 2 of a contaminated region reveals features, and "spots", not unlike those observed in much purer material by Yamane and Ueda. Consequently, one can conclude that: 1) titanium can be thinned satisfactorily in a solution at room temperature for transmission electron microscopy; 2) large-scale precipitates, >50A in diam, are absent in commercial-purity, recrystallized, furnace-cooled, A-70 titanium containing about 3800 wt ppm interstitials and the same amount of iron; 3) care must be taken not to attribute surface contaminants to precipitates. The authors are grateful to the Franklin Institute for financial support and permission to publish.

Jan 1, 1967

By Esau Arinaitwe

The most extensively used methods and approaches for characterizing and selecting frothers for plant application are based on measurements in clean water in the absence of solids and without consideration of physicochemical factors that are known to influence the properties and behavior of practical flotation froths, and hence metallurgical performance. These factors include aquatic chemistry, collectors, modifiers, particle shape and size, and pH. Understanding interactions between these factors and the role and relevance that they play in characterizing frothers and froth zone properties will greatly aid in the frother selection and evaluation process. Such an understanding will also greatly enhance the ability to create frother solutions that are sufficiently robust to plant variability (such as mineralogical, comminution, operational, physical and chemical). This paper reports results of a “systems approach” study designed to probe ways in which physicochemical factors a) influence frothing and bubble size in 2- and 3-phase systems, and b) confound correlation of fundamental measurements with flotation outcome. The concepts and insights gleaned from fundamental frother studies, together with empirical laboratory flotation tests and a thorough understanding of plant needs and challenges were used to develop and implement customized frother formulation at a Cu-Mo flotation plant. Plant metallurgical data and qualitative assessment of froth zone properties showed that a frother formulation developed using the approach described here provided marked improvements in froth characteristics (mineralization, mobility, stability, persistence). These improvements resulted in better circuit control and significant performance advantages, such as improved coarse particle recovery, without detrimental downstream effects. INTRODUCTION Successful recovery of value minerals of sufficient grade via the flotation process requires selective manipulation of the value minerals to render them hydrophobic using collectors and modifiers. The hydrophobized minerals then attach to air bubbles and the mineral-bubble aggregates are selectively transported through the froth zone into the launder. Frothers play a critical role in froth formation and more importantly in achieving and maintaining desirable froth characteristics (for example, stability, mobility, persistence) that are essential in facilitating the selective transport and recovery of value minerals. Fundamentally, frothers contribute to the transient stability of thin liquid films and so aid in reducing bubble coalescence rates in the pulp and froth zones, thus maintaining the small bubbles required for efficient bubble-particle attachment. Invariably, the action of frothers at the liquid/gas interface in the pulp and froth zones impacts bubble stability, water and solids transport to the froth, which in turn impacts froth mobility, froth carrying capacity, and the concentrate grade. Consequently, selection and evaluation of frothers for flotation plants requires careful consideration of the frother type, though the choice is limited to only two main classes of compounds in terms of fundamental chemistry (short chain alcohols and polyglycols). Even with the selection of what is thought to be the best frother for a given application, practical knowledge of the froth zone and experience in controlling its properties is essential for maximizing metallurgical performance.

Jan 1, 2019

By C. M. Shigley

The record of the largest magnesium plant in the country utilizing sea water as a primary raw material stands as another victory in the struggle for large-scale production of pure chemical elements from the oceans. Bromine was the first element to be economically and commercially recovered when, in 1933, a six million pound per year unit was built on the Atlantic Coast near Wilmington, North Carolina, by the Ethyl-Dow Chemical Co.' The bromine was reacted with ethylene to produce ethylene dibromide, a necessary associate of tetraethyl lead in ethyl antiknock fluid. Within eight years that plant was doubled and redoubled, and another large unit was put into operation at Freeport, Texas, taking bromine from the Gulf of Mexico. In 1940, in fortunate anticipation of the impending demand for magnesium and the limitations of the production from natural brines, the Dow Chemical Co. turned again to the sea, and designed and erected a plant at Freeport to produce 18,000,000 lb. of magnesium per year from sea water.2 In March 1941, two months after the new plant had been put into successful operation, construction of an 18,000,000-lb. Government-financed addition was started. Late in 1941, as a part of the $400,000,000 defense program intended to produce 600,000,000 to 700,000,000 lb. of magnesium per year, Dow was asked to build and operate for the Government a 72,000- ooo-lb. plant near its original plant at Freeport. Within a few days after the Japanese made their attack at Pearl Harbor, preparation of a site was started at Velasco, Texas, for the construction of the Texas plant of the Dow Magnesium Corporation—the largest known magnesium plant utilizing sea water as its raw matcrial. Less than six months after the first piling was driven, urgently needed magnesium was in steady production. Five months later, the plant had exceeded its rated capacity. It remained above rated production until the general magnesium curtailment was applied to Government-owned plants in September 1944. Credit for the construction and operation record of the Dow magnesium plant must be divided among the efforts of a very determined group of men, an extremely favorable location, and a simple process. It is with the two latter aspects of the plant that this article will deal. The site for the new enterprise was chosen near Velasco, Texas, on property bordering the Brazos River at a point approximately 6 miles inland from the original magnesium plant and 7 miles inland from the Freeport Harbor on the Gulf of Mexico. This location was far enough away from the operating units to reduce both interference from construction and the vulnerability of closely bunched manufacturing operations to the then feared air attack. Yet the distance inland was not too great for the economical construction of a barge canal from the harbor to serve both the transportation of raw

Jan 1, 1949