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By L. C. IlsLey

The Burrell methane indicator developed by Col. G. A. Burrell in 1915 was modified during 1918 to the " insulated type " Burrell indicator . Though this instrument has been used in mines since its first introduction , it was not formally approved as safe and efficient for use in gaseous mines until March , 1922. The indicator is built by the Mine Safety Appliances Co. , Pittsburgh , Pa . Approval No. 800 has been assigned to this company covering the device in the form approved . The principle of the Burrell indicator is fairly well known . It works on the same principle as the Haldane apparatus - measuring by means of a graduated scale , the contraction resulting from burning the methane content of an air and gas sample . A given aniount of the mine atmosphere to be tested for methane is taken into the indicator . This sample is then thoroughly shaken with the water in the indicator for one minute . The methane in the sample is now burned for two minutes by a heated platinum coil , which is energized by a small storage battery containing two Edison cells connected in series . After the glower or coil has burned for two minutes , it is disconnected from the battery , and the indicator is again shaken for The amount of contraction observed by reading the height the water rises in the U- tube of the indicator , which is calibrated to register the percentage of gas in the sample . The indicator as approved by the Bureau of Mines is equipped with an Edison M- 8 mine- lamp battery , and with a glass and scale graduated from zerɔ tɔ four per cent methane . The approval of a corbination service that is , an attachment for giving illumination from the same battery from which the platinum glower is operated has been withheld pending the development of a battery with larger capacity . - Under laboratory operating conditions , a single glower was capable of making at least 20 determinations if proper precautions were observed with respect to the use of freshly charged batteries . A battery should be allowed to stand off charge several hours before using in an indicator , as the peak voltage of a freshly charged Edison cell is abnormally high and is likely to send enough current through the glower to melt the platinum wire . Fifty-seven satisfactory operations were obtained from a completely charged battery . The point where the voltage of the cell fell below 2.3 volts with the glower attached was taken as the lower limit of successful operation . A voltage less than 2.3 does not permit the lower to burn at a temperature that will completely burn the gas in the air , thus giving a false indication .

Jun 1, 1922

By L. C. IlsLey

It is most unfortunate that a few must often suffer , and perhaps die , that many may benefit from a fuller appreciation of the dangers surrounding mining operations , and the proper use of safeguards designed to offset these dangers . This has been the case during the last few months in the anthracite field , where several explosions , resulting in burnings and fatalities , were undoubtedly caused by the use of defective or improperly assembled flame safety lamps . Had this same lack of understanding regarding the proper assembly of this safety device , and had the same indifference on the part of the user in keeping the lamps locked at all times and in proper condition , been permitted in a region where the coal dust was explosive , a terrible disaster would surely have resulted . In nearly every coal -mining country , including the United States , the Davy and Clanny type of lamp has been superseded , and rightly so , by some kind of bonneted lamp . Bonneted lamps , especially of the naphtha type , while much safer than the unbonneted , have more parts and need more care in assembling . For this reason the burden of responsibility for their being properly assembled rests with the lamp man ; therefore , it is extremely important that only competent , conscientious men be given this work .

Dec 1, 1921

By F. W. Jessen, R. W. Durie

The development in the laboratory of mall salt cavities illustrated that the rate at which salt dissolved was affected by the rate at which fresh water was injected into the cavity. Experiments carried out with small salt cores indicated, however, that the rate of salt removal with forced convection was not significantly different from that with natural convection, provided the flow was in the laminar range. The rate of salt removal was apparently controlled by the water salinity, which was indirectly determined by the rate at which fresh water was injected. With turbulent flow at the salt surface, the rate of salt removal could be increased several-fold with the same salinity in the cavity. A quantitative evaluation of the rate of salt removal under natural convection was attempted by using techniques developed for the analogous heat convection system. The values of the rate of salt removal obtained experimentally were considerably greater than the calculated values due to surface irregularities. To determine the effect of these irregularities, the rate of salt removal was found for a smooth surface by experimentally determining the rate of salt removal as an irregular surface developed and then extrapolating this trend back to initial time. Excellent agreement between the experimental and calculated values then were obtained. INTRODUCTION Underground cavities leached from massive salt formations have been used extensively for many years for storing LPG products. The development techniques for storing methane gases in the liquid state at a low temperature leads to speculation that underground salt cavities of controlled geometry may be quite important in this field. In any case, the increasing number of salt cavities used for storage indicates an impending need for a better knowledge of the washing process to efficiently develop maximum storage volumes within specified regions. It is the purpose of this paper to discuss some aspects of -the mechanism of the dissolution of salt in the formation of cavities of controlled geometry. Although no attempt is made to resolve the problems that may be associated with the storage of methane in the liquid state in salt formations, the fact that salt cavities can be formed at reasonable cost would appear to justify investigation in this direction. Further, underground cavities have the advantage of allowing storage at higher pressures (and hence higher temperatures) than surface installations. Although the thermal conductivity of salt is somewhat higher than frozen earths, 0.017 cal/(sec) (cm)(c) for salt at 32F vs 0.005 cal/(sec) (cm) () for ice at this temperature, this disadvantage is offset by the higher storage temperature in underground cavities. For example, at atmospheric pressure, the temperature of liquid methane would be —258F. In a storage cavity at 2,000 ft, the pressure would be approximately 1,000 psia, which would allow a storage temperature of -9OF. The formation temperature would be roughly 110F at this depth, hence the total temperature differential would be 200F. Under surface storage conditions the corresponding temperature difference would be some 333 F (75 F + 258 F). The results of this difference can be compared more readily by substituting the appropriate values of thermal conductivity, temperature difference, heat capacity and density into the solution of the diffusion equation for the unsteady linear conduction of heat. For the underground salt storage, the heat flux is 15cal/(sec) (sq cm), where t is the time in seconds from the initiation of the cooling. For the surface system, and using the thermal properties of ice at the average temperature, the corresponding figure is 21cal/(sec) (sq cm). Hence, the heat losses through ice under surface conditions would be some 30 per cent higher than the underground salt storage. This ratio would vary, considerablv. Of course, for other conditions, but it indicates that an advantage over surface facilities may be obtained by storage in underground salt cavities.

Jan 1, 1965

By B. F. Decker, D. Harker

The recrystallization reaction in OFHC and spectroscopically pure copper has been followed by X ray diffraction determinations of the amount of material with the cold-worked and recrystallized textures in specimens which had been given various heat treatments. The heats of activation for recrystallization are found to be: 29.9 kcal per rnol for OFHC copper and 22.4 kcal per rnol for spectroscopic copper. A reliable and speedy method for measuring the amount of recrystallized material in a piece of rolled metal, together with a new scheme for low temperature heat treatment, have made possible a determination of the activation energy for recrystallization in rolled copper. This method for studying recrystallization rates is different from others reported in the literature.' Oxygen free, high conductivity copper—OFHC copper—of purity 99.98 pct gave an activation energy for recrystallization of 29.9 kcal per mol with recrystallization data taken in the temperature range 208-245°C. Copper of purity 99.999 pct from the American Smelting & Refining Co. gave an activation energy for recrystallization of 22.4 kcal per mol with recrystallization data taken in the temperature range 43-135°C. The great change in activation energy for recrystallization as the small amount of impurity in the metal is decreased suggests that the motion of grain boundaries is conditioned by the impurities which must be concentrated in them. Thus the OFHC copper recrystallizes with an activation energy of the order of magnitude to be expected for the diffusion of the kind of impurities likely to be present. The high purity copper, on the other hand, recrystallizes with a much lower activation energy, in accord with the notion that the grain boundaries need not wait for the impurities to diffuse with them. Copper when strongly cold-rolled has a well-defined crystallite texture which can be described by saying that the [111] direction of the face- centered cubic crystals is aligned approximately in the cross-rolling direction and that faces of the form {110} lie approximately in the rolling plane. When such rolled copper is completely annealed the crystallite texture is quite different: the [loo] direction is in the cross-rolling direction and faces of the form (100) are in the rolling plane. At intermediate stages of annealing, the copper contains both textures. If a Geiger counter X ray diffraction spectrometer is set up so as to measure the intensity of the (200) Bragg reflection from the rolled surface of a piece of copper strip, this intensity increases from a very small value in the cold-worked state to a maximum as the specimen is annealed. The same would be true for (200) reflections from planes normal to the rolling or cross-rolling directions. In view of this fact, the amount of material in the copper which has recrystallized can be measured by comparing the intensities of one of these reflections just mentioned with the intensity from a similar specimen after complete annealing. In much the same way the intensity of (111) reflections from planes normal to the cross-rolling direction could be used to measure the amount of unrecrystallized material in the specimen. In the work to be described here the amounts of recrystallized and unrecrystallized material in partly annealed rolled copper specimens were determined in this way. Experimental Procedure Samples of oxygen free high conductivity (OFHC) copper, cold-rolled 99.7 pct to 0.002 in. thickness, were subjected to heat treatments for various chosen times at four different temperatures, and high purity copper samples, cold-rolled 98.0 pct to 0.0015 in. thickness, at six different temperatures. Each sample, after treatment, was placed in an orientation sample holder for use with a Geiger counter X ray spectrometer and a reading was taken of the intensity of the (200) reflection from planes perpendicular to

Jan 1, 1951

By Jr. F. M. Perkins.

Capillary forces play a controlling role in water-drive displacement processes both in laboratory experiments and in actual reservoirs, but their quantitative importance may be quite different in the two cases. Because of the importance of conducting laboratory experiments which are representative of field conditions, it is necessary to understand exactly the role of capillary forces in the displacement process. Though a number of experimental investigations related to this subject are contained in the literature, there appears to be a lack of information pertaining to unsteady-state experiments in water-wet media. This experimental study was conducted to obtain additional laboratory data to clarify further the role of capillary forces in both the macroscopic and microscopic flow of oil and water in porous materials. THEORETICAL CONSlDERATlONS The capillary pressure is defined as the difference in pressure between a continuous oil phase and a continuous water phase in a porous material.' The magnitude of this pressure difference depends on the interfacial curvature and the interfacial tension. The interfacial curvature is determined by the geometry of the pore spaces, the wettability of the rock surfaces, and the quantity of cach phase present. Capillary forces are involved in a water-drive displacement process in that they exert a controlling influence on the microscopic fluid distribution which in turn is reflected in the saturation or macroscopic flow behavior. Microscopic Fluid Distribution Because of the microscopic nature of the displacement of oil by water, it is necessary to consider the flow and the fluid distribution in individual pores. On this microscopic scale the capillary Forces, which act over a distance of one or two sand grain diameters, control the distribution of oil and water under static equilibrium conditions. When an external force is applied to the fluids, such as in a water-injection experiment, the applied forces tend to distort the oil-water interfaces. However, in most fine-grained, water-wet sands, the applied pressure difference across one or two grain diameters is usually several orders of magnitude less than the capillary pressure difference. These con-siderations lead to the theory' that even during flow the capillary forces continue to control the microscopic distribution of oil and water within the pores of a porous material for all practical reservoir and laboratory flow rates. This concept of capillary forces controlling the microscopic distribution of fluids has been substantially verified by other investigators3-7 who have found a lack of dependence of relative permeability and residual oil saturation on rate of fluid injection. Macroscopic Distribution The microscopic influence of capillary forces cannot be observed easily and only the effect on the macroscopic or average saturation can be detected. The saturation, of course, is really the point of interest. During a water flood, large differences in saturation at the flood front cause large capillary pressure gradients. This, in turn, causes water to advance ahead of the flood front thereby reducing the capillary pressure gradient in this region. The result is that in homogeneous porous media capillary pressure gradients tend to cause a diffuse displacement front. At low rates in laboratory columns, the front may extend over the entire column length. When the advancing water first reaches the outflow face of a core, the water, which is the wetting phase, cannot be produced because the pressure in the water just inside the core is lower than the pressure in the oil-filled space around the outflow face. This difference in pressure is equal to the capillary pressure for the water saturation existing at the outflow face. Water, therefore, accumulates at the outflow end of the core which causes a reduction in the capillary pressure. Because the capillary pressure does not vanish except at the residual oil saturation,7,8,9 water will not be produced until the residual oil saturation exists at the outflow face. This entire effect,"' which is called the "boundary effect", results in a region of relatively high water saturation near the outflow face. At low rates of injection in a short column, this region of high water saturation may extend over a considerable portion of the column. The influence of capillary forces on the macroscopic flow of oil and water have been described by Leverett.10 For unidirectional, viscous flow in the absence of gravity segregation, the expression in dimensionless form for the fraction of water in the flowing stream, f, is

Jan 1, 1958

By Allan Neal, Jess W. Smith

FirstMiss Gold Inc., formerly FRM Minerals, Inc., is heap leaching gold from difficult, low grade ores at the Getchell Gold Mine in Humboldt County, Nevada. A two-thousand ton per day facility is leaching material from waste dumps from old operations and from small, satellite deposits around the main orebody. A three-thousand ton per day pressure oxidation mill is under construction to treat sulphide gold ores in the main deposit. The Getchell gold deposit occurs in a wide fault zone in carbonaceous, calcareous sediments and intrusive rocks. Intense clay alteration is typical within the mineralized zones. Silicification is a less dominant alteration phase within the ore. Gold in the upper oxide portion of the deposit occurs both in a free state which is leachable and complexed in fine disseminated pyrite which is not leachable. Gold in the lower sulphide portion of the deposit is almost entirely complexed in pyrite. Abundant arsenic sulphides and variable amounts of remobilized carbon are associated with the gold mineralization throughout the deposit. The waste dump material being heap leached was derived from the upper portions of the gold deposit. The dumps contain high quantitites of fines, clays and sulphides of iron, arsenic and other trace elements. No stockpiling of low grade materials was attempted during the old mining operations, therefore, the dump material is highly variable and unpredictable. A sample of representative material from one dump contained the following:

Jan 1, 1998

By G. K. Herzog

THE problems relating to the construction and use o{ mining equipment are basically the same as those pertaining to equipment in other industries. It is true that much mining equipment is subjected to an unusually severe combination of shock, high stresses, abrasive wear, and corrosive ? conditions, but in other industries, also, equipment must operate under similar adverse conditions, so that their experiences may readily be applied to the problems of the mining industry. Steel and iron are for obvious reasons the most suitable materials for the construction of the greater part of the equipment used in the mining industry. For a large percentage of uses, the ordinary plain carbon-steels and ordinary or unalloyed cast-iron are, and probably will continue to be for a long rime, entirely satisfactory materials. There are, however, many. conditions which cannot be satisfactorily taken care of by these materials, and alloy steels and irons must be used where the greatest possible economy is desired. In general, alloy steels and irons are used where it is necessary or economically desirable to use materials possessing properties superior to plain carbon-steel or ordinary cast-irons, such as: (a) better strength (b) better wear resistance (c) better corrosion resistance (d) better resistance to high temperatures (e) better resistance to low temperatures or combinations of two or more of these advantages. There are also conditions where materials having other special properties are required, such as, for example, certain magnetic or electrical properties for electrical machinery; but in the great majority of cases, one or more of the five advantages listed are the most desired.

Jan 1, 1938

By Murray Hitzman

Central and southern South America contain a variety of sediment-hosted base metal deposits including shale-hosted Zn-Pb-Ag, Irish-type Zn-Pb-Ag, BrokenHill-type Pb-Zn-Ag, Mississippi Valley-type Zn-Pb, and sediment-hosted stratiform Cu deposits. In this sense, the paper point out Potential for sediment-hosted base metal deposits in Central and Southern South America.

Apr 24, 2001

By Raymond Rivera, Federico Cernuschi

Este documento trata de actualizar el conocimiento de un tipo de venillas no tan conocidas que llevan el nombre early halos. El reconocimiento de este tipo de venillas puede servir mucho para la vectorización lateral y vertical de un sistema porfirítico, pero debido a la escasa información que existe sobre ellas muchas veces son subestimadas, desapercibidas o hasta incluso confundidas con otros tipos de venillas. Este tipo de venillas son más comunes en los pórfidos Cu – Mo, mientras que en los pórfidos con afinidad hacia el Au (Cu – Au / Au – Cu) suelen ser más escasas o no estar presentes, pero aun así su débil presencia puede servir como una herramienta de vectorización.

Sep 30, 2022

By C E. Loo

The cohesive zone has a significant influence on blast furnace operation and performance. The high-temperature behaviour of the reducing ferrous burden determines the properties of the cohesive zone and is studied in a laboratory using a softening and melting test. This paper summarises recent work aimed at better understanding the physico-chemical changes in a ferrous bed during softening and melting. Using a combination of standard, non-standard and quenched tests supported by optical image analysis, this work has highlighted some important differences in behaviour between lump, pellets and sinter as well as between the two lump ores and two acid pellets. As with the two lump ores, one pellet is denser than the other and the denser ore/pellet is also higher in silica.Within the cohesive zone, densification of the ferrous layer leads to reduced permeability and greater pressure drop. From the work presented in this paper, two key mechanisms of ferrous layer densification were identified. Firstly, the formation of fayalite, which melts at low temperatures, can act as a lubricant between particles and lead to densification of the ferrous layer under load. Secondly, reduction to metallic iron is topochemical and fayalite in the outer shell of the particle restricts gas flow to the centre and results in the formation of a wustite core. As temperatures rise, the wustite melts and eventually exudes out of the shell, occupying void spaces between particles and further reducing layer voidage and permeability.This work has shown that density and silica content are two key factors that result in increased densification of the ferrous layer. For porous Australian lump ore, the higher porosity and lower silica reduces the formation of fayalite and limits the formation of the wustite core, ultimately leading to a more permeable bed. The difference between the porous and dense ores was more pronounced with the introduction of prereduction to better simulate blast furnace conditions. This suggests that a porous ore gives significantly better cohesive zone permeability and blast furnace operations than a dense ore.The behaviour of fluxed sinter particles is different in that the silica reacts with lime to form dicalcium silicate. Minimal wustite core forms because relict hematite particles in sinter are only small and reduction occurs uniformly throughout the particle. Even with significant loss of material through prereduction, beds formed from sinter remained comparatively rigid. Studies using a typical blast furnace burden mix of 80 per cent fluxed sinter and 20 per cent porous ore showed that the structures of these beds at softening and melting temperatures are very similar to the sinter-only test. Measurements of bed contraction and pressure drop also confirmed that the introduction of 20 per cent porous ore into a sinter bed did not alter its properties. This is an indication that 20 per cent ore can be easily accommodated by blast furnace burdens.CITATION:Loo, C E, 2015. Important acid ferrous burden properties in the cohesive zone of a blast furnace, in Proceedings Iron Ore 2015, pp 367–376 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By D Mather

Fibre reinforced shotcrete has long been accepted as temporary ground support in both mining and civil tunnel linings. Greater knowledge and understanding of shotcrete as an engineering material and advances in its quality have enabled designers to eliminate the requirement of a second pass lining. The initial shotcrete can be used as the permanent tunnel lining or be incorporated into the final lining design, rather than ignored and considered as sacrificial.For the initial shotcrete lining to become a suitable permanent tunnel lining it must be able to establish and maintain stability over time when subjected to all possible loads. To ensure this, best engineering practices are required during the design phase of the project to analyse; ground conditions including rock mass properties, ground and ground water quality, rock mass loadings, excavation sequencing, and construction methodologies.The increased durability of steel fibre reinforced shotcrete over traditional tunnel support, in the right ground conditions, makes it possible to achieve a serviceability design life of greater than 100 years (as required in most infrastructure projects). By specifying a good quality shotcrete mix and ensuring adequate quality control for batching, fibre dosing, preproduction testing, excavation sequencing, and application control, it is possible to eliminate the requirement of a secondary lining. The properties of a suitable permanent lining must be well established using suffi cient preproduction testing in combination with post application monitoring.A major benefit of steel fibre reinforced shotcrete in ground support applications is the removal of any requirement for workers to work under unsupported ground. Its use can also significantly reduce, if not eliminate, the manual handling associated with placing traditional reinforcement such as lattice girders and mesh. Both of these benefi ts are significant with the ever increasing focus on health and safety.This paper highlights factors that must be considered in the design and specifications to enable an initial shotcrete tunnel lining to become the permanent shotcrete tunnel lining.

Mar 8, 2011

By Jon C. Volkwein

The U. S. Bureau of Mines investigated a method of using inert gas to prevent the formation of explosive gas mixtures in auger highwall mining of coal. A combination of gasoline and diesel engine exhaust gases was introduced into the auger drill hole using a short section of pipe located at the collar. Gas samples were taken and analyzed on site with infrared detectors for oxygen, carbon dioxide, methane, and carbon monoxide. Evacuated bottle samples were also taken and analyzed by gas chromatography at the Pittsburgh Research Center. These gas results were analyzed for explosibility. Personal exposure to carbon monoxide was also monitored. The highest methane level observed was 9.55 pct. The Inert gas levels, (carbon dioxide and nitrogen) were sufficiently high to prevent any ignition of the methane. Results showed that for all conditions during mining, gas concentrations were non-explosive. The maximum personal time weighted average sample for carbon monoxide was 20 ppm. This system provides a safe, inexpensive, simple method for preventing explosions during auger mining. INTRODUCTION The auger highwall mining method is an effective method to recover coal from a reserve when removal of the overburden by surface mining equipment becomes uneconomical. In this method of mining, a horizontal auger enters the coal seam from the surface mine bench under the highwall and the coal is drilled in a series of parallel holes. Historically, coal mined from the surface is relatively shallow, and over time, methane associated with the coal has dissipated through the surface. In most circumstances, little methane has been found associated with auger mining. However, mining technology has enabled surface mining of deeper reserves of coal. Furthermore, environmental constraints have forced the highwall extraction method to be used to remove coal under wetlands, further increasing the chances of encountering methane. Recently incidents of methane explosions at a few auger mining operations have resulted in injuries and increased testing for methane at the collars of auger holes. The fuel source of the reported explosions was not necessarily limited to methane, but may also have involved coal dust. The Mine Safety and Health Administration (MSHA) met with the Bureau to discuss what technology might be available to enable the safe resumption of mining. The discussion included the difficulty of ventilating through the solid shafts of the augers, that steel bits probably created the ignition source, and that perhaps inerting the holes with low oxygen and high carbon dioxide concentrations from the machine's diesel exhaust was a potential solution. Considering the ventilation aspects of the problem, it was not clear If ventilation could be reliably established. If some degree of ventilation to the front of the mining head is achieved, it may combine with methane to bring the hole atmosphere from a rich, nonexplosive mixture to an explosive mixture. Furthermore, it may not prevent a dust explosion in such a mining configuration. Lack of access through the shafts of auger type mining machines further limits the ability to add water or air to cool bits to prevent an ignition source from developing. Either of these approaches would also be expensive. The process of mining coal In an inert atmosphere has been considered in the past, but to our knowledge, never implemented (Department of Interior, 1970). Clearly, implementation in underground mining would be more complicated. On a mine bench open to the atmosphere, however, adding inert gas to the mining head could provide a quick, feasible method to prevent explosions at auger highwall mining operations. Also the problem of how to move the inert gas to the cutting head of the machine had to be considered. Preventing explosions on auger mining machines using inert gas requires three primary considerations: first is the source of inert gas; second, placing the inert gas at the cutter head; and third, monitoring the hole atmosphere. Any gas source having an effective inert gas concentration of 34 volume pct or greater will prevent methane from Igniting (Zabetakis, 1965). Sources of inert gas considered for this application included liquid nitrogen, modified shipboard inert gas generators (for hydrocarbon shipping and transfer), jet turbine engine (Paczkowski, et. al., 1982), the auger's diesel engine and a gasoline engine. Operation cost, purchase cost and availability limited our testing to the diesel and gasoline engines. This work tested each engine, separately and combined. To ensure effectiveness, both company and enforcement personnel need to know how to monitor the condition of the inerted hole. Measurements at depth inside the hole are possible by remote sampling through rigid tubing, but this method is Impractical for routine monitoring. Continuous monitoring of the exhaust gas stream is an alternative. The U. S. Bureau of Mines evaluated an inert gas system at an auger mining operation at a surface mine near Owensboro, KY. Coal was mined from the Number 9 Coalbed in Henderson Co. KY. Tests were conducted in January and March of 1992.

Jan 1, 1993

By Gen-ichi Nakazawa, Masamichi Fujimori, Ichiro Doi

INTRODUCTION The major products of nickel industry in Japan are the electrolytic nickel (E-Ni), the ferro-nickel (Fe-Ni) and the nickel oxide sinter (NOS), totaling in production to 90 - 100 X 103 tons/year in terms of pure nickel content, and sharing approximately 20% for E-Ni, 10% or less for NOS, and the rest for Fe-Ni. The raw materials to be used for these products depend totally upon the imports from abroad; matte, the raw material used in the production of E-Ni, is from Australia and Indonesia; matte for the NOS is from Indonesia and New Caledonia; and garnierite, the raw material for Fe-Ni, is imported in bulk from New Caledonia, Indonesia and the Philippines. This report describes the acquisition of these raw materials, the production processes, trend of supply-demand, problem areas and the measures taken, and so forth, in connection with the nickel industry in Japan. ELECTROLYTIC NICKEL 1. History The nickel industry in Japan began in 1939 by the Sumitomo Metal Mining Co., Ltd. (SMM) with 150 tons E-Ni/month under the blast furnace smelting and the metal anode electrolysis. This was temporarily suspended after the end of last World War II, but resumed in 1952 with garnierite of New Caledonia as a raw material. Since then, the production capacity has been expanded incidentally to the increase of demand. However, since the smelting of garnierite contained about 2% nickel was very costly, the smelting process had been changed to smelt the briquettes of the mixture of garnierite and nickel sulfide concentrate imported from Canada and Australia. In 1975, this smelter was closed, and SMM's nickel refinery used matte from Western Mining Corp. (WMC) of Australia as a raw material. The refining process had been the metal anode one. The crude nickel oxide derived from roasting the high grade nickel matte had been melted and reduced with coke in the electric furnace and cast into anode. In 1970, a new refinery was constructed to produce 1,000 tons Ni/month under the matte anode process, and further, in 1976, the old refinery was converted into the matte anode process, thus resulting in the total production capacity of 1,690 tons Ni/ month today under the matte anode process. Shimura Kako Co., Ltd. began in 1951 to produce a little of E-Ni, and in 1955, they constructed a shaft furnace smelter with New Caledonian garnierite as a raw material to begin the E-Ni production under the matte anode process. Its monthly production reached 400 tons in 1957, but it discontinued its smeltering in 1963 to convert to the matte import from S. Le Nickel, WMC and INCO. In 1979, it changed the source of matte supply from the above to Soroako of Indonesia, discontinued the E-Ni production in 1979, and arranged for a total toll contract with Sumitomo. Meanwhile, SMM has developed a very distingished process for nickel-cobalt separation and refining by a solvent extraction method from nickel and cobalt mixed sulfide generated as a by-product in the nickel recovering process from laterite under the Nicaro process. Nippon Mining Co., Ltd. (NMC) also developed its own process. The plants under these processes were constructed in 1975, with the monthly production capacity of 230 tons of nickel and 130 tons of cobalt by SMM, and 300 tons of nickel and 110 tons of cobalt by NMC. SMM imports the raw material from the Marinduque Mining and Industrial Corp. of the Philippines, while NMC from Queensland Nickel Pty. Ltd. of Australia. 2. Production Status 2.1 Producers and materials Today, SMM holds the monthly E-Ni production capacity of 1,920 tons, of which 1,690 tons are from the matte anode process, and the remaining 230 tons are from the mixed sulfide as a raw material. NMC also uses the mixed sulfide as a raw material for a production of 300 tons/month. [ ] Most of SMM's raw material, nickel matte, is imported from WMC of Australia, partly using the Soroako matte from Indonesia. An example of material analysis is given in Table 1. Matte is a raw material for E-Ni and NOS, and the major sources of matte supply at present are Australian WMC and Indonesian Soroako, occupying ap-

Jan 1, 1982

By H. Lyn Bourne

Daily everyone depends on the great variety of glass products, so much so that glass is often taken for granted. In fact most people do not realize how versatile glass has become. Consider the various uses and then try to imagine a day in which we are not influenced by glass. Common uses include container ware, table ware, window glass, lead crystal, automobile glass, and fiber glass. Several less common, but important, uses include laboratory ware, pharmaceutical, TV bulbs, light bulbs, glass ceramics, optical glass, fiber optics, and laser glass. Corning, Inc., a leader in specialty products, uses nearly 1 000 different compositions to manufacture about 60 000 different products (Edwards and Copley, 1977). Glass is such a complex product that-definitions vary and exceptions can be found for most definitions. Glass is an inorganic amorphous (non-crystalline) solid. Most glasses are produced by melting of a mixture of oxide raw materials, and then cooled to room temperature. Soda-lime-silica composition.s account for about 90% of all glasses melted (Anon, 1973). The properties of the glass product come mainly from its chemical composition. All of the different glasses require melting a combination of raw materials and forming the molten material into the desired shape. Both the melting and the forming processes use sophisticated technology and these technologies require experts to manage these production systems. The manufacturing process is continuous and takes place in tonnage quantities, so adjustments in the batch to achieve the desired finished product requires a great deal of expertise. Raw materials are fed to the batch mixing area in very large quantities (tons in most cases). As a result, impurities in the range of 0.1% result in addition of that impurity within the molten glass in kilogram amounts. More than twenty different industrial minerals are consumed in the manufacture of various kinds of glass (O'Driscoll, 1990). This chapter describes the major and minor ingredients of the various glass batches. It discusses the roles of the various oxides in the glass batch and most importantly considers the mineral raw materials which supply the glass industry. Each of the raw materials is described in detail in other chapters so the geology and mineralogy sections are kept brief here. Container glass, by far, accounts for the most production; followed by flat glass, fiber glass, and specialty glass of which table ware accounts for the greatest tonnage. [Table 1] shows the general production data for 1987 through 1990. Statistics for many of the uses do not appear because production volumes are small compared to the major uses. The glass industry is organized in four categories: containers, flat glass, fiber glass and specialty glass. The US Department of Commerce, Bureau of Census, publishes production data about the glass industry in three different categories: 1) glass containers, 2) consumer, scientific, technical and industrial glassware, and 3) flat glass. The Bureau has very complete statistics about the glass industry in these three categories but they report production data in different units according to industry standards. Therefore, [Table 1] gives the production data in dissimilar units. The production of most glass articles follows similar steps. The raw materials are mixed and the resulting batch is fed into the furnace. In soda-lime-silica glasses melting begins between 600 and 900°C. At these temperatures CO, and other gasses are released which create bubbles in the molten glass. To remove the bubbles and insure complete melting the temperature is raised to between 1 500 and 1 600°C. This is the melting-refining stage during which the refining agents in the glass batch serve to aid in the release of gas bubbles, homogenize the melt, and prevent the formation of scum on the surface of the molten liquid. At the conclusion of the melting-refining stage the glass is too fluid for working and the melt is cooled to about 1 100°C to attain the proper viscosity for working and forming to begin. After the glass article has been made, it must undergo annealing (slowly and uniformly reheated and cooled) to remove thermal stresses that were created during the forming process.

Jan 1, 1994

By J. S. Rinehart, W. C. Maurer

The mechanics of impact crater formation in rock, particularly sandstone, has been sutdied, the velocity range being approximately that normally associated with oilwell gun perforators. The bullets were small steel spheres having diameters of 3/16, 9/32 and 7/16 in; impact velocities ranged from 300 to 7,000 ft/sec. The craters have two distinct parts — a cylindrical hole (or burrow) with a diameter the same as that of the impacting sphere, and a wide-angle cup comprising most of the volume of the crater. The burrow is fornred as material in front of the projectile is crushed and pushed aside, forming a cylindrical hole surrounded by a high-density zone. The clip forms as fractures are initiated in front of the projectile and propagate along logarithmic spirals, approximaling maximum shear trajectories, to the free surface of the rock. A most significant observation (made for the first time) was that, below the base of the cup in one type of sandstone, there are a group of similar fractures, not extending to the surface, which are spaced uniformly a few millimeters apart. Each fracture follows roughly the contour of the base of the cup and appears to require a certain threshold impulse to initiate it. These fractures comprise a relatively high fraction of the total, newly exposed surface area. The volume of the material removed by crushing varies as the first power of the impact velocity and the volume removed by fracturing, as the second power of the impact velocity. Penetration varies linearly with the impact velocity and is inversely proportional to the specific acoustic resistance of the target material, the proportionality constant being dependent upon the shape of the projectile. INTRODUCTION Yield of oil from a producing well is frequently enhanced by firing bullets and shaped charges through the well casing into the oil-bearing rock, forming craters and fractures from which oil can flow more readily. The purpose of this investigation has been to develop a better understanding of the mechanics of impact crater formation in rock, particularly sandstone, the velocity range being approximately that normally associated with oilwell gun perforators. FORCES OPERATIVE DURING IMPACT When a projectile moving at considerable velocity strikes a- massive target such as oil-bearing sandstone, intense and complex transient stress situations develop within both the projectile and the rock or sandstone against which it is striking. Usually the struck rock fails, the missile or projectile penetrating into the rock to some depth where it comes to rest or is forcibly ejected from its burrow by expansion of a plug of target material compressed in front of it. When the impact velocity is very high, the projectile itself may fail, breaking apart or becoming distorted; this situation is not considered here, the discussion being limited to nondeforming projectiles. Many experimental studies'.' have been carried out to determine the nature of the mechanics of crater formation and the salient features of the forces coming into play, some of the earliest studies being the French Army experiments performed at Metz between 1835 and 1845.' The stratagem in most instances has been to make a post-mortem examination of the crater, measuring volume and depth of penetration and deducing force relationships from these observations rather than performing the more difficult (usually almost impossible) feat of measuring stresses during penetration. In many materials, the force acting during penetration of the projectile is found to be the sum of two components—(1) a constant force, independent of the velocity, representing some inherent strength of the target material; and (2) a component, proportional to the square of the velocity, representing inertial forces. For such materials, the average force per unit area acting on the projectile at any instant while it is in motion and being decelerated may be written F/A = a + bv2 . . . (1) where v is the velocity of the projectile at that instant, A is the cross-sectional area of the penetrating projectile taken normal to its trajectory, and a and b are constants which are dependent upon the target material and the shape of the projectile. It follows that the total penetration s is given by .........(2) where v, is the velocity of the projectile when it just strikes the target. Values of a and b for spherical projectiles impacting in a loose sand-gravel mixture and compacted earth were obtained in the Metz experiments. For sand-gravel, a and b are 620 psi and 0.0115 (psi) (ft/sec)', respectively; and for compacted earthworks, a and b are 432 psi and 0.0008 (psi) (ft/sec)'. Figs 1 and 2

By W. K. Hartley

INTRODUCTION AND HISTORY The Kimberley Division of De Beers Consolidated Mines Limited includes four underground mines in Kimberley viz. Dutoitspan, Bultfontein, De Beers and Wesselton Mine and two outside mines, viz. Finsch, an opencast operation and Koffie- fontein, an underground operation. Figure 1 is a map of the Kimberley district showing the location of these mines. Finsch is 150km west of Kimberley and Koffiefontein is 80km south of Kimberley. The head office of the De Beers Company is in Kimberley. The diamond sorting and trading companies for diamonds produced in South Africa are also in Kimberley. Diamonds were first discovered in the Kimberley area at Hopetown in 1866 on the banks of the Orange River. This, and later finds, triggered the start of the great "Diamond Rush". In 1870 the Jagersfontein (now closed) and Dutoitspan pipes were discovered, followed by those at Bultfontein, De Beers and Kimberley Mine (now closed) in 1871, Koffiefontein in 1880 and Wesselton in 1891. Finsch Mine was a more recent discovery in 1961. Finsch, due to go underground in 1986, and Koffiefontein, are fully self-contained with their own treatment plants, housing, services etc. The four Kimberley mines deliver ore by conveyor to a central treatment plant and use joint ser- vices. ,+lining is presently carried out mainly by Block Caving which has replaced the earlier Chambering method. De Beers Mine, on a limited scale, and Wesselton Mine have recently started to use Sub-Level Caving, and both Bultfontein and Du- toitspan are scheduled to use the same method at a later date. The combined ore tonnage produced by the Kimberley Mines is 17 400 tons per day, Koffiefontein 12 000 tons per day and Finsch Mine 17 000 tons per day. THE EVOLUTION OF MINING METHODS Initially, the open-cast mining was carried out to as great a depth as safety would permit. In the early stages of open-cast mining, the Kimberlite was removed in horizontal slices from levels 13m to 20m apart. The rock broken in the workings was raised to the surface by means of endless rope haulage inclines. As lower levels were attained these rope haulages were replaced by aerial hoists, inclined shaft haulages or inclined skip-ways. These methods were dependent upon the strength of the unsupported pipe walls, which became increasingly unstable with depth and a system was adopted whereby the ground mined in open excavations was tipped down passes to tunnels connecting with vertical hoisting shafts sunk in the country rock. As work in the open pit became more and more impracticable due to extensive falls of ground, this system was replaced by a combination of open-cast and underground mining, and eventually by wholly underground methods. Chambering With the advent of underground mining, a system known as "Chambering" (or "Inclined Chambering") was evolved. It was essentially a combination of shrinkage stoping and caving, whereby advantage was taken of the high pressure exerted by the loose rock that had accumulated in the open excavations due to the collapse of the pipe walls. This mining method was introduced around 1890 and was successfully employed, with little change, up to the early nineteen seventies. In this system each mine was served by a main hoisting shaft situated about 300m in the country rock. Main haulage cross cuts were driven from the shaft to the pipe at 200m (600ft) vertical intervals. A sub-vertical shaft, situated close to the pipe provided access to the working levels. Chamber levels were established from the sub-vertical shaft at 13m (40ft) intervals. See figure 2. On each of these levels successive 6,8m wide “cuts" were mined out across the pipe, from one rock contact to the other, and extended vertically to the footwall of the level above. Every 6,8m cut was made up of a chamber and a "back-pillart1, each approximately 3,401 wide. The chamber was mined by overhand shrinkage stoping. The levels were carried "en echelon", with each chamber on a level approximately three cuts ahead of the chamber on the level

Jan 1, 1981

By R A. Pax, B Cornish

"Although it is well known and easily experienced that the feed size distribution influences the throughput and stability of operation of a semi-autogenous grinding (SAG) mill, the details are poorly understood since it is often only the input and output measurements that are used to identify unfavourable in-mill conditions. Load cells and power monitoring are sometimes used but are generally poor indicators of the mechanisms that lead to such unwanted conditions.At the MMG Century operation a 12-microphone array around the SAG mill using non-contact acoustic measurement methodology has been used to diagnose the in-mill conditions to facilitate improved throughput and stability of operation. Use of acoustic measurement and control has provided a demonstrated increased throughput of at least 20 per cent as well as improved stabilisation of the mill. Since the information derived from the acoustic noise from a mill allows discrimination of the behaviour of the steel balls and the rock, a number of in-mill conditions can be identified with signals that include toe and shoulder angles, water holdup and the ball to rock ratio.Feed size variations, measured online with an imaging system, have been used to determine their effect on the internal charge behaviour. It has been found that circumferential segregation of steel balls and rock occurs as well as poor mixing of large rocks. Both of these effects are not conducive to optimal breakage rates of different size fractions. Longitudinal segregation effects also occur. This paper presents and discusses these results and illustrates how the new acoustic knowledge can be used to manage and improve mill operation and stability.Citation:Pax, R and Cornish, B, 2016. Understanding Size Effects of Semi-autogenous Grinding (SAG) Mill Operation as a Pathway to Solving Feed Disturbances - Case Study using the MMG Century SAG Mill, in Proceedings 13th AusIMM Mill Operators’ Conference 2016, pp 321–330 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 10, 2016

By S. Makni, S. Koppalkar, A. Bouajila

This paper reviews briefly the gravity recovery involving conventional gravity separators and centrifugal gravity separators. The advantages of semi-continuous gravity separators will be discussed. Techniques for upgrading the gravity concentrates will be addressed briefly. The research efforts in optimizing semi-continuous gravity separators will be recalled. Innovative gravity circuit for eliminating or minimizing the use of complex processing schemes will be reviewed. On-going gold gravity research at COREM has revealed that gold particle shape transformation is the main factor explaining the lower recoveries of the plant units. Efforts to devise possible remedies for improving gravity recovery at plant will be discussed.

Jan 1, 2011

By F Minz, P Lamberg, A Mwanga, C Lund, C Wanhainen, M Amiri Parian, J Rosenkranz

A geometallurgical model is currently built in two different ways. The first and the most common way relies on geometallurgical testing, where a large number of samples are analysed for metallurgical response using small-scale laboratory tests, eg Davis tube testing. The second, mineralogical approach focuses on collecting mineralogical information over the orebody and building the metallurgical model based on mineralogy. At Lules University of Technology, Sweden, the latter method has been adopted and taken further in four ongoing PhD studies. The geological model gives modal composition by the help of element-to-mineral conversion and Rietveld X-ray diffraction. Texturally, the orebody is divided into different archetypes, and liberation measurements for each of them are carried out in processing fineness using IncaMineral, a SEM-based technique. The grindability and liberation spectrum of any given geological unit (sample, ore block, domain) are extrapolated from the archetypes. The process model is taken into a liberation level by mass balancing selected metallurgical tests using the particle tracking technique. The approach is general and can be applied to any type of ores. Examples of ongoing studies on iron and massive sulfide ores are given.CITATION:Lamberg, P, Rosenkranz, J, Wanhainen, C, Lund, C, Minz, F, Mwanga, A and Amiri Parian, M, 2013. Building a geometallurgical model in iron ores using a mineralogical approach with liberation data, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013 , pp 317-324 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 29, 2013

By Lawrence Devon Smith

The selection of the production rate is one of the most crucial decisions to be made in the development of a mineral property. However, it is not uncommon for the production rate to be decided arbitrarily, before any technical evaluations begin, and it is often not reconsidered when better data becomes available. A number of methods are used in the mineral industry to select the production rate for a project. There is general agreement that the size of the ore reserve has a strong influence on this decision, but what other factors are involved? Is there a way to identify an optimum rate? This paper examines the factors that influence the selection of a production rate from the perspective of project economics. The NPV maximum value should be viewed as a “failure” point. The “two times capital” maximum is proposed as a lower limit. A practical check is suggested to confirm that a selected rate is in an optimum range. Production rate assumptions are much more likely to be reviewed for marginal projects, but they should also be reviewed for viable projects.

Jan 1, 1997