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By Lyle F. Smith

With the demand for larger diameter raises under all ground conditions and the increasing price of drill rods and operating costs, it was evident that there was a need for a raise borer with a variable-speed drive and fast-response torque limiting control. The raise boring machine described in this paper was developed after several years of market research and discussions with mine operators and construction contractors. This paper discusses the Ingersoll-Rand raise borers with the solid-state controlled DC motor drives. These machines have gained wide acceptance in the five-year period since being placed in operation and are designed to accommodate the range of raise diameters and lengths most required today.

Jan 1, 1978

By P. G. West-Sells

At Barrick's Zaldivar Copper Mine in Chile, the addition of sulfuric acid to the ore going to heap leaching is a major expense. A process in which a portion of this acid is replaced by elemental sulfur added to the heap has been developed by Barrick and Westcoast Biotech that has the potential for significant cost savings. Initial shake flask tests with plant raffinate solution indicated that oxidation of elemental sulfur by a culture of sulfur-oxidizing bacteria, some indigenous to the Zaldivar heap, showed a significant lag period. This lag period could be reduced by dilution of the raffinate, but dilution would not be economical to implement at Zaldivar. Short (0.75 m) column tests indicated that after the initial lag period, sulfur biooxidation occurred steadily and the acid requirement for obtaining the same copper recovery as the control without sulfur was reduced by as much as 13 kg/t. These tests were followed by taller (5.1 m) column tests that confirmed the reduction in acid consumption, with no negative effect on copper extraction.

Jan 1, 2007

By Tanai Marin

Modelling of pyrometallurgical processes can become extremely challenging when the process involves complex chemical reactions, phase change, multiphase flows, heat and mass transfer and fluid flow. Traditionally modelling involving complex thermodynamic equilibrium and thermochemistry are treated separately from problems in which transport phenomena are dominant, such as CFD and heat transfer studies. The separation is also necessary due to the lack of modelling tools that can effectively couple both solution thermodynamic and transport phenomena or even multiphysics problems. However, there are, quite often, problems in high temperature process metallurgy in which it is not possible to simplify this two modelling approaches. Typical cases of this are for example, the modelling of slag freezing on furnace walls (electric or flash furnaces) to create a protective layer in front of the refractories hot wall, the formation of build up from the matte or metal phases at the hearth of the furnace, calcine smelting from charging piles into slag phase, etc. The problem is even more challenging in nonferrous pyrometallurgy as most, if not all, of the condense phases are highly non ideal solutions involving several components and exhibiting complex crystallization paths and miscibility gaps, such as the Fe-Ni-Cu-S matte, FeO-Fe2O3-SiO2, FeO-MgO-SiO2 slags, etc. In this paper, an approach to model coupled problems involving multiphysics and solution thermodynamic is presented. The approach consists in developing in-house a library containing the Gibbs energy minimization routines and using the local equilibrium concept for process modelling from a multiphysics commercial software (COMSOL). Simple test cases, such as building a non-ideal binary phase diagram on a CFD package is presented, the limitations and of other classic approaches such as equilibrium or Scheil cooling are discussed and compared to this model.

Jan 1, 2015

By Paul Billingsley

NICKEL Plate mountain coincides closely with that portion of the Hedley mining district which is of economic interest. From it has come almost the entire past production, and the active mines of the present epoch, Nickel Plate-Sunnyside and Hedley Mascot, lie in a compact group near its summit. The Hedley district, which is 210 miles due east of Vancouver, is situated in the Okanagan range of southern British Columbia, where this rather gentle uplift is deeply dissected by the canyons of the Similkameen and its tributaries. The region is essentially part of the Interior Plateau of the Province (Figure 1), but is adjacent to easterly elements of the Coast Range. The Similkameen canyon, 4,000 feet in depth and frequently less than 4 miles from rim to rim, traverses the district from northwest to southeast. The deepest and narrowest portion of the canyon lies between Hedley and Keremeos, 18 miles to the southeast. At Hedley, an important tributary, Twenty Mile creek, enters from the north in a bold canyon, making a focus of unusually rugged topography. Nickel Plate mountain occupies the sector to the east of this junction (see Figure 2) and its flanks are ringed by the cliffs of Twenty Mile canyon on the west and by those of the Similkameen canyon on the south. On the east, it slopes gently to the hanging valley of Eighteen Mile creek, and on the north, with only slight loss of elevation, it merges into the moderately hilly surface of the plateau.

Jan 1, 1941

By Fred E. Gray

The mill of Bralorne Mines, Limited, is located on the government highway in the Bridge River mining district of British Columbia, approximately fifty miles northeast of Shalalth, the shipping point on the P.G.E. railway. Prior to 1931, the property now owned by Bralorne Mines, Limited, was operated by Lorne Gold Mines, Limited. During the early operation of the mine, the ore had been treated at various times in one of three small amalgamation mills on the property. With the succession to ownership by the Bralorne Mines, Limited, work was started on construction of a new mill of approximately 100 tons capacity per day. The new mill was put into operation in the month of February, 1932, and since that date has been in continuous operation, milling an average of slightly over 3,000 tons of ore a month. The ore treated in the Bralorne mill is a gold-bearing quartz containing a very small amount of silver. About 75 to 80 per cent of the gold is in the form of free gold, the principal remaining portion being intimately associated with pyrite and arsenopyrite. The gangue rock is quartz, together with varying but relatively small amounts of altered diorite or albitite. In some of the ore, usually the higher grade portions, there is also considerable talc, which forms much slime in grinding. The mill is situated near the portal of the main haulage tunnel and the ore is received directly from the mine ore-trains, which dump on to the grizzly over the coarse-ore bin of the crushing plant. Figure 1 shows the flow-sheet of the mill as it is now being operated. Various changes have been made in the flow-sheet and in the reagents since the commencement of operations, and doubtless minor changes will continue to be made in order to increase the efficiency of the mill. Among the more important changes were those from two-stage to single-stage grinding, and from amalgamation plates to blankets. The change in the crushing was made because too much crushing was done when both primary and secondary mills were employed.

Jan 1, 1933

By L. M. Dwarkin

Introduction Avital factor in the success of any coal mining operation is the method of roof support employed. A coal seam in? place has been described as the filling in a geological sandwich and the process of mining as the art of snatching the filling from the midst of the sandwich. The difficulty of this task can be appreciated -in the realization of the tremendous weight bearing on a coal seam due to overlying strata; a weight that can exert pressures of several thousand pounds per square inch. Over the years that mining has been carried on, methods have been evolved by arduous trial and error, making possible the successful extraction of billions of tons of coal. However, these time-proven methods are not enough. The constant demand for faster and cheaper production, the necessity of mining under increasingly difficult conditions, and the requirement for higher standards of safety, call for ever better and more efficient roof support. A recent and very promising development in answer to this need is that of Roof Bolting. Roof Support in General Before proceeding to the subject proper, a brief discussion of the older roof support methods is of value in illuminating the difficulties involved and the problems to be solved.

Jan 1, 1951

By H. A. Hamza, J. Visman

"This is a report on the combined application of ""long-chain"" flocculants and classifier cyclones in the field of slimes treatment, with emphasis on the simultaneous thickening and clarifying of industrial effluents, at low cost. The principle of the process described below involves: (1) the formation of small floes of relatively great strength, by sequential dispersion and conditioning, with average floe size of about 1501 microns; (2) the floes being of sufficient strength to withstand the shearing forces operating within a cyclone and consequently behaving as solid particles with an equivalent diameter (Stokes) that exceeds the size cutpoint of the classifier cyclone; (3) the removal of flocculated solids by means of a classifier cyclone the size cutpoint of which permits effective separation of these solids from the water. Results of test work to date indicate that a broader use of cyclones in industry is possible with the above types of flocculants because the submicron particles and colloidal material are readily collected and clear water is produced, with the attendant advantages of compact construction and low cost that are characteristic of centrifugal separation. The process generally permits classifier cyclones of medium throughput capacity (e.g. 12-in. diam.) to be used for clarifying effluents that contain minus 10-micron particles."

Jan 1, 1973

By M. Kucharski

The model of the Outokumpu direct-to-copper process deals with a steady-state process the parameters of which do not depend on time. To give a prediction for the process outcome, "a modified black box" method, which ignores most of the chemical reactions in the furnace, was used. This method takes input and output parameters into consideration. The presented model describes the Outokumpu direct-to blister process in the form of the set of linear equations which give mass and enthalpy balances as well as the distribution of elements between blister copper and slag. However, only major substances where quantities in the feed charge exceed a certain level (for example 0.5 kg/ lMg of smelted concentrate) are taken into consideration. Minor elements are distributed among slag and metal phases in a different way. Some of the equations are derived from distribution coefficients. The value of distribution coefficients can be calculated from plant production data.

Jan 1, 2007

By Zolfan F. Oszter

As exploration for new mineral deposits advances farther into the Canadian North, the more the design engineer is faced with the problems of cold weather conditions. This is especially true when designing belt conveyor systems for open-pit operations where the temperature in the winter may go to -40°F or below. Problems with the mechanical components and also with the belting will arise, which in turn may reduce the availability of the system. This paper outlines the problems encountered during cold weather operation and what precautionary measures must be taken into consideration in the design of equipment, as well as in the selection of materials, in order to ensure a system which will maintain production. THE conveying of ore and waste in the pit or to the processing plant poses problems during cold weather, when temperatures may drop to -4o?F or be-low. Troubles with the mechanical components, the belting or the supporting structure may arise, and this, in turn, will reduce the availability of the conveying system.

Jan 1, 1965

By Mohamad Kharseh

There is general acceptance that climate change, which is the most important challenge facing humanity, is anthropogenic and attributed to fossil fuel consumption. Therefore, improving the performance of our existing energy systems and deploy more renewable energy resources is an urgent issue to be addressed. Geothermal refers to existing of heat energy in deep rock and sedimentary basins. This energy can be used to drive a power turbine to generate electricity. Traditionally, geothermal has been exploited in places with the plentiful hot water at relatively shallow depth. In the light of fact that ground temperature increases with the depth everywhere on the Earth, engineered geothermal systems (EGS) can be installed in any place to exploit the geothermal in generating energy. Unfortunately, the high exploration and drilling costs of boreholes is the main barrier to commerciality of EGS worldwide. In addition, there are technical problems associated with drilling big depth. In oil producing countries such problems can be addressed by utilizing whether active or abandoned oil or gas wells and, consequently, EGS can produce power at profit. The current study presents an analysis of a binary geothermal power generation system for commercial electricity generation in Qatar. For this purpose, two binary cycles are assumed the main difference between them is that the first one is air cooled while the other will be water cooled. The performance of the two cycles and the possibility of improvement has been shown. Economic analysis the power plant shows that the levelized costs of electricity is 3.6US¢/kWh and the pay-back time is less than 8 years.

Aug 1, 2013

By M. Maeda, H. Sasaki

"Precious metals such as platinum, palladium, and rhodium are commonly used for automotive catalyst converters. To recover these metals with lower environmental load and cost, exposure to zinc vapor was proposed as a pretreatment prior to acid leaching. Alloying with zinc is expected to enhance dissolution of precious metals and reduce the amount of toxic chemical agents required for the leaching. This study examined the effect of the alloying pretreatment using model catalysts. Catalysts were immersed in leaching solutions for a predefined time with or without zinc vapor pretreatment, and the amount of dissolved metals was then evaluated. INTRODUCTIONAutomotive catalysts account for large portions of the world consumption of platinum (Pt), palladium (Pd), and rhodium (Rh) (Butler, J., 2011). Because of their limited production and high prices, recovery of the precious metals from spent catalysts is obviously an important issue. A representative structure of an automotive catalytic converter can be described as a monolithic ceramic support with a honeycomb structure which is coated with porous oxides such as those of aluminum (Al2O3), cerium (CeO2), and zirconium (ZrO2) (Heck & Farrauto, 2001; Matsumoto & Shinjoh, 2007). Small particles of precious metals, originally several nanometers or less, adhere to the oxides.Typical recovery processes for precious metals from scraps are shown in Figure 1 (a)-(c) (Habashi, 1997; Yoo, 1998; Shibata & Okuda, 2002). Precious metals can be recovered by using non-ferrous smelters as Figure 1 (a) shows. Smelted along with metals such as copper (Cu), precious metals and ceramics in the scraps are absorbed into the metal phase and slag phase, respectively. Because the metals separate from Cu as anode slime in successive electro-refining, such a process can recover precious metals as co-products of Cu. However, in order to prevent negative effects on the principal process, a specialized procedure is needed. Figure 1 (b) is a specialized approach in which scraps are melted at elevated temperature with Cu called the Rose process. Here Cu works as a collector metal, and the absorbed precious metals are then concentrated by oxidation of Cu. Such a process can separate precious metals from ceramics constituting catalyst converters. Smelting at high temperature, however, requires extensive equipment as well as a high amount of energy. On the other hand, hydrometallurgical processes may be effective to separate precious metals from ceramics into an aqueous solution, because chemical agents can act on these metals selectively (Figure 1 (c)). Leaching of precious metals, however, requires strong acid and oxidizing agents because of their corrosion resistance; this makes liquid waste disposal and the large environmental load controversial. To solve the problem, a metal vapor pretreatment was proposed (Okabe, Yamamoto, Kayanuma, & Maeda, 2003a, 2003b; Kayanuma, Okabe, Mitsuda, & Maeda., 2004; Kayanuma, Okabe, & Maeda, 2004; Itoh, Kayanuma, Miyake, Kimura, & Maeda, 2005). A flow diagram of the process using zinc (Zn) vapor is shown by Fig. 1(d). Exposed to Zn vapor, precious metals are expected to form alloys which dissolve more easily than pure precious metals, thus reducing the use of chemical agents and environmental load involved with successive leaching. In previous research (Sasaki, Miyaki, & Maeda, 2010; Sasaki & Maeda, 2010, 2012), dissolutions of precious metal alloys formed by Zn vapor treatment were examined in hydrochloric acid (HCl) using an electrochemical method, and it was confirmed that these metals dissolved at higher rates from the alloys than from their pure states (Figure 2). In this study, the availability of the alloying pretreatment was demonstrated by testing on model catalysts in leaching solutions."

Jan 1, 2012

By Vilkul Yuri

Operation of the iron ore deposits of Krivbass which has been going for 130 years, has led to serious problems related to reaching great depths in open pit and underground mining. Extraction of the huge volume of rock mass (about 9 billion m3) resulted in mass violations of the natural state of the subsurface. At present most mining operations are carried out either by combined open pit and underground mining or by open pits in areas underworked by underground mining. Now the process of movement of the strata underworked by underground mines is not completed ? work fronts of open pits approach the areas of possible funnel, terrace and fissure formation, movement continue to develop. With deepening mining operations the remaining reserves are worked off under layers of incumbent rocks of > 900 m. Significant underground voids arise from mining "blind" deposits at large depths. Heaving of voids over the mined area lasts for decades; it may result in not only unpredictable time of funnel emerging and area development of movement synclines, but also the threat of air blows in mines. Forced caving of incumbent rocks as a way of controllable backfilling is difficult to achieve for economic and technical reasons. This is much more expensive and dangerous than stoping. Furthermore, the mining practice does not know examples of enormous volume backfilling under a km thick rock layer. Currently, RMI "KNU" is conducting research and pre-design work directed at creating a specialized pilot Geoinformation System (GIS) of continuous remote monitoring of rock mass in areas of underground voids influence in Krivbass area. Implementation of this system will enable to determine promptly in time and space the areas of the beyond cut off stress condition, to inform about possible occurrence of events that may result in accidents and to improve safety of mining along with other systems of forecasting and controlling the stress condition in rock mass. RMI "KNU", with a large number of professionals from many Ukraine?s leading mining research centers, has developed a research and design work program that includes the following principal stages: 1. Analyzing geomechanical conditions of combined mining of magnetite quartzite and rich ore deposits; determining mechanisms of moving of rock strata underworked by underground mining and places of hazardous area monitoring. 2. Investigating anthropogenic factor influence on the safe state level of rock strata underworked by underground mining (laboratory studies). 3. Developing processing techniques for continuous remote monitoring of rock mass in areas of underground cavity influence. 4. Modeling the state of rock mass based on results of pilot GIS hardware measurements. 5. Developing an expert diagnostics system for rock mass state in areas underground void influence. 6. Experimental investigation and testing the GIS, its improving for efficient application. After testing the pilot GIS it is supposed to be implemented at all mining enterprises as an integral part of iron ore mining and processing technology in the Krivbass area.

Aug 1, 2013

By D. A. Livingstone

Abstract Golden Manitou Mines, Limited, nine miles east of the town of Val d'Or in northwestern Quebec, are presently milling, at the rate of 1,000 tons daily, a relatively complex 3.11-phide ore and producing, in decreasing order of value, zinc, silver, gold, lead, and copper. The products are a zinc concentrate, a silver-lead-copper concentrate, and a silver-gold precipitate from cyanidation. Although several independent pre-production metallurgical investigations into the treatment of the ore were carried out, current practice is different from any of the recommended methods and, while still leaving room for improvement, is considerably more effective than earlier performance. Flow-sheet developments from the commencement of operations in 1942 are dealt with briefly, showing the reasons for making numerous changes, the shortcomings of the various schemes tried, and, finally, covering data pertinent to present-day practice. From the recovery standpoint, silver has presented the greatest difficulty. The writer's association with the Company coincided with the start of production. Characteristics of the Ore The predominant metallic minerals present in the ore, in decreasing order of their abundance, are: pyrite, sphalerite, galena and chalcopyrite, with lesser amounts of argentiferous tennantite-tetrahedrite, arsenopyrite, covellite, bornite, and chalcocite. The sphalerite is mostly light in colour, sometimes colourless, and practically free from combined iron. Some of the sphalerite is very closely associated with minor amounts of pyrite, chalcopyrite, and galena. Pyrie is abundantly disseminated as medium to fine grains; more rarely it is massive and contains occasional grains of sphalerite, galena, and chalcopyrite.

Jan 1, 1952

By D. R. Nagaraj, Napoleon Tercero, Raymond Farinato

"Dialkyl dithiophosphinates are a unique class of collectors used in the flotation processing of Cu- Au, complex polymetallic, and precious metal ores. They are known to exhibit higher selectivity and flotation rates for Cu sulfides, galena, and precious metal values in the presence of other gangue sulfides, than the traditionally used collectors such as dialkyl dithiophosphates and xanthates. They also have an attractive HSE profile. Historically, for most reagents used in flotation, practical methods of chemical usage have preceded detailed scientific understanding of how they work; this applies also to dithiophosphinates. Recently, however, the research community has shown significant interest in this chemistry, and numerous studies have been conducted to understand the fundamentals of interactions of dithiophosphinates with several base metal sulfides and precious metals. A critical overview of fundamental studies, laboratory ore flotation, and plant practice of dithiophosphinates is given in this paper. The findings from these studies are rationalized by drawing on concepts and insights from coordination chemistry. A comparison between the interactions of dithiophosphinates and dithiophosphate analogues with various minerals and metals is also made throughout. We believe that such a critical analysis/overview will be invaluable in identifying gaps in knowledge regarding this important set of compounds, which in turn may provide a basis for the design of studies that can advance understanding of their flotation performance. INTRODUCTIONDialkyl dithiophosphinates are an important member of the class of thiophosphorus compounds that are widely used in flotation beneficiation of base metal sulfide and precious metals ores. Ore flotation studies and plant practice experience have demonstrated their unique properties compared to the traditionally used sulfide collectors such as dialkyl dithiophosphates (their closest analogues) and xanthates. The predominant dialkyl dithiophosphinate used in the industry today is sodium diisobutyl dithiophosphinate (a.k.a. DIBDTPi; AEROPHINE® 3418A1 introduced by American Cyanamid in the late 1970s; Figure 1). The widest application of AEROPHINE 3418A is found in complex polymetallic ores of Cu-Pb-Zn-Ag, Pb-Zn-Ag, and Cu-Zn where it has demonstrated improved recoveries (and flotation rates) of sulfide minerals of Cu, Pb and Ag and selectivity against sphalerite, iron sulfides, and penalty elements (Mingione, 1991). More recently, it has also been used as a collector in the processing of Cu, Cu-Ni, and Ni ores (Guitard et al., 2015) and Cu-Au ores (Dunne, 2005). In general, it is considered to possess the collecting power of xanthates together with the selectivity (against pyrite) of dithiophosphates (Hope et al., 2003a, 2003b), while requiring lower dosages than those reagents (Guitard et al., 2015). AEROPHINE® 3418A has also found an important place in several targeted formulations to improve flotation performance. Other benefits of this reagent include higher flotation rates, having a better HSE (health, safety and environmental) profile2 compared to xanthate, and exhibiting resistance to hydrolysis and oxidation (Guitard et al., 2015). Compared to diisobutyl dithiophosphate (DIBDTP), which is known to contribute significantly to frothing (often leading to over-frothing), AEROPHINE® 3418A is observed to be froth neutral (i.e. it does not alter the existing froth characteristics). As is the case for other flotation reagents, the design and application of AEROPHINE® 3418A has preceded understanding of its behavior at a fundamental level. Recently, there have been numerous fundamental and application studies designed to develop a better understanding of its interaction with sulfide minerals and precious metals. This should help in optimizing its application in flotation and provide a basis to better rationalize its unique selectivity and activity characteristics. Past studies on the interacti"

Jan 1, 2016

By J. A. Russell

AS THE LOCAL electric generating plants increased in size and capacity and costs of producing electricity came down, it became apparent that it was no longer good economy to operate individual boiler plants at the Company's various collieries to power steam hoists and compressors. Great economies could be made by burning the Coal at central power plants and using electric power at the collieries. During the period after the end of the war, a steady and continuing programme of replacing first the smaller steam-driven units, and later the steam-driven compressors, was undertaken, and finally the larger hoists, those at No. 12 Colliery in New Waterford and No. 20 Colliery in Glace Bay, were the last units operated by steam.

Jan 1, 1958

By Donald Carlisle

"IntroductionTHE OPINION is held by many persons outside the mining industry, and by some within it, that mine •operators customarily find it advantageous to mine •only the richer parts of orebodies and to leave behind, perhaps permanently, leaner material which could be mined only at a lesser profit. This opinion and the inference which might be drawn from it that 'gutting' of mineral resources is a common feruture• of the mining industry are prejudicial to the intertsts of the industry and deserve careful scrutiny. Oddly enough, a theory of mine economics which gives much support for the opinion was advanced first by members of the mining profession itself, has been repeated •several times in mining literature, and to my knowledge has never been adequately refoted.According to the theory, selective mining of the rich ore first on a large scale will yield substantial gains and these gains do not merely reflect the desire of 'money minded investors' for a quick return but derive from the logical application of compound interest rates and the concept of ;present value. W. O. Hotchkiss and R. D. Parks (1) illustrate the reasoning with a hypothetical example in which an orebody of 4,000,000 tons may be mined either in its entirety in forty years at $1.00 per ton profit or selectively with the recovery of only the richest half in twenty years at $1.50 per ton profit. They show that, with Hoskold rates of 4 and 8 per cent, the present value of profits from mining only the richer 2,000,000 tons exceeds the present value of profits :from mining the whole orebody by about 20 per cent."

Jan 1, 1953

By M. C. Harris

The past twenty to thirty years have seen unprecedented research activity aimed at understanding the performance of flotation froths. In 1990, Finch and Dobby coined the term froth recovery factor, Rr, to represent the efficiency of the froth in delivering particles from the pulp-froth interface to the concentrate. Originally defined for flotation columns, Rr has also been found useful in modeling mechanical flotation cells, and a number of techniques, based on quite different approaches, have been developed for measuring the froth recovery in industrial flotation cells. None of these techniques is universally successful. This paper reviews the concept of the froth recovery factor, and the basis of each of the techniques that have been proposed to measure Rr, and examines the reasons for this parameter being so difficult to pin down. Advances in Mineral Processing Science and Technology Proceedings of the 48th Annual Conference of Metallurgists of CIM Sudbury, Ontario, Canada Edited by

Jan 1, 2009

By Arthur Notman

In view of the alarm expressed in many quarters over the increased production of copper to come from Canada and Africa in the future, it is of interest to study the history of the principal producing districts of the world and to attempt to forecast the future part to be played by them. Tables 1 and 2 record the world production and that of the United States, together with their respective rates of increase, both in percentage and actual amounts. The important point in this record is the decline in rates of increase for both the world and the United States, but more particularly in that of the latter, since 1881. The increase for the United States in the last decade is less than that for any decade since 1881, and is only 20 per cent of that for the decade 1911-20. The increase for the world is less than that for the two preceding decades, and is only about 65 per cent of the war decade 1911-20. Referring to Table 1, the decade 1871to1880 showed an increase of 32.l per cent in the world's production, which was the low until the last decade, ending 1930, which shows an increase of only 24.5 per cent. These decades both represent periods of economic stress following wars. In the first case the recovery was very sharp, the percentage increase in 1881-90 reaching a high of 86.4 per cent. It would seem dangerous, however, to assume that a similar high will be reached in the next ten years, because of the severity of the depression brought about by the Great War and the probability of its influence extending well into the next decade.

Jan 1, 1931

By R. J. Enright

Eight mines in canada and three in Australia were visited between September 1987 and Februacy 1988. All mines had encountered problems with sulphide dust explosions and had introduced defence mechanisms to prevent or limit the extent of these explosions. Techniques included dispersion of limestone by various methods, use of water foggers, placing of water and inert stenuning in blast holes, and removal of settled dust to prevent secondary explosions. The techniques were evaluated and it was concluded that certain practices should be modified or discontinued as they were of little value in preventing dust explosions.

May 11, 1988

By K. Sugawaral, W. Tongampl, K. Hiruta, Takasaki Y., A. Hosoi, T. Katol, A. Shibayamal

Printed circuit boards of discarded personal computers, televisions, mobile phones and so on contain a significant amount of valuable metals. For the recycle of valuable metals contained in the printed circuit board wastes (P CB W), the effect of temperature on chlorination-volatilization of metals under chlorine gas flow was studied. The experimental sample used was powder of two types: crushed PCBW (Sample 1) made from PC and industrial machines, and crushed-combusted PCBW (Sample 2) made from Sample 1. The principal element in Sample 1 and Sample 2 was copper, about 13 wt% and 20 wt% respectively. Sample 1 was composed of pure metals such as copper, tin and lead. In comparison, Sample 2 consisted of oxide metals such as copper oxide, tin dioxide and silicon dioxide. The two samples were heated at predefined temperatures (ranging from 300 °C to 900 °C) in chlorine gas. The chlorine gas flow into the furnace was set at a rate of 100 mL/min. As a result, copper in Sample 1 began to volatilize at 600 °C and completely vaporized at 700 °C. While in Sample 2, copper started to volatilize at 600 °C and reached 100% vaporization at 800 °C. For Sample 1, zinc finished volatilizing at 500 °C, while for Sample 2, zinc volatilization reached only about 80% (maximum). Also, lead, nickel and titanium showed similar results. Consequently the volatilization rate of Sample 1 is higher than Sample 2 and the temperature dependency was verified for copper, lead, zinc, nickel, antimony, titanium and chromium. In fact, the volatilization rate of these metals increased with the temperature. This result reflects the difference between pure metal and metal oxide, as Sample 2 was combusted before the chlorination-volatilization test. Moreover, the chlorination reaction may be accelerated by a reductive atmosphere because Sample 1 contains carbon. In this study, the volatilization behavior of some valuable metals in PCBW by chlorine gas was observed. Further improvement in the rate of volatilization can be achieved by optimizing the experimental conditions.

Jan 1, 2011