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By H. J. Ilgner

The paper reviews the results to date of an integrated research programme currently being conducted by CSIR Mining Technology, aimed at impacting significantly on concentrated mining with face advance rates of up to 20 times faster than presently achieved. Consideration is given to the interplay between innovative technologies for continuous non-explosive mining or rock comminution underground and lateral hydraulic transportation in cross cuts, haulages and eventually hydro hoisting. Novel technologies for underground comminution of reef and development waste and an innovative pumping system suitable for the hydraulic transportation of coarse comminuted rock are reviewed. Emphasis is placed on South African gold mine reef and waste samples comminuted by electric shock wave comminution technology. The innovative TORE® pump system appears to be well suited for the successful hydraulic transportation of rock, produced by electric shock wave comminution, at solids concentrations between 45 % and 67 % by mass.

Jan 1, 1997

By Kenneth O. Bayard, John D. Sullivan

"This repot continues the series of papers being prepared at the Southwest Experiment Station of the Bureau of Mines in cooperation with the department of mining and metallurgy, University of Arizona, on factors involved in the leaching of ores. A list of previous publications is given in the bibliography at the end of this paper.Four factors are essential to the leaching of any copper-bearing ore: (1) A solution that will attack the copper minerals must get into the body of the ore particles; (2) the copper minerals must be dissolved by the solution; (3) the solution containing the dissolved copper must find its way out of the voids of the rocks into the main solution; and (4) the copper in solution must be recovered by some means of precipitation.This paper considers the third step -- the extraction of the water-soluble copper. From leached ores. The usual commercial practice in leaching copper ores consists of: (1) Crushing the ore to pass through a screen with approximately one-quarter or one-third inch openings; (2) charging the crushed ore into a fat; (3) adding, usually by upward percolation, a solution that will attack the copper minerals; and (4) washing the copper-laden solutions out of the ore with solutions weaker in copper than the pregnant solution, and finally with water. A countercurrent system of leaching and washing is usually employed. In some plants the solutions are circulated during leaching and washing, and in others they are not. In some plants the wash solutions are added by upward percolation, while at Chuquicamata the so-called ""piston-displacement"" method of washing is employed.The following condensed resume of the leaching cycle of the plant of the Inspiration Consolidated Copper Co. is given to present to the reader a picture of the steps involved in a typical leaching cycle. This plant was recently described in detail.4 The ore, crushed to pass openings 11/2 inches in diameter, passes over a screen with ¼-inch openings. The undersize from the screen goes to a Dorr howl classifier that removes the primary slime from the circuit. The overflow from the classifier is treated in a slime-leaching plant. The oversize of the ¼-inch screen passes through rolls, where it is crushed to pass a 3/8-inchscreen. The undersize from the 3/8-inch screen join with the coarse material from the classifier, and the combined products become the feed to the leaching vats. There are 13 concrete lead-lined leaching vats. The complete leaching cycle is 13 days. Eight tanks are under acid-iron content at all times; three are being washed, one tank is being filled, and the one is being excavated. The electrolyzed solution from the tank house, to which amount of new acid has been added, is put on the oldest tank of ore under need This solution travels from tank to tank and finally emerges from the newest ore"

Apr 1, 1931

IT has often been stated in the past that the materials known as inert fillers were merely adulterants nsed in the manufacture of paint, and that their purpose was to cheapen paints made from more expensive pigments, such as zinc oxide or white lead. In recent years, however, practical tests in various parts of the world have proved that many of the inert fillers, when used with wisdom, possess very definite and advantageous properties.Another oft-repeated statement which had a far-reaching effect in the past, and which is not without influence to-day, is that the very best paint must always be made from a pure pigment, without admixture with either a fellow pigment or an inert filler. Both or either of these statements may be true in a few isolated instances, but for the vast majority of ordinary painting conditions there is an abundance of experimental evidence to prove that neither is true.With the desire to further investigate these points and po 'determine the best combination of pigments and fillers for use as an outside white paint on plant buildings, a number of panels were prepared in the laboratory of this company.

Jan 1, 1924

By R. Boom, J. Halder, B. Snoeijer, B. Deo, A. Overbosch

Operational data of BOF and the slag samples for different starting conditions of phosphorus(0.06–0.26 per cent P) and silicon content (0.3–1.2 per cent Si) of hot metal have been analysed. The contribution of parameters that are well known to affect phosphorus distribution at tap, like basicity, temperature, FeO content of slag, slag mass etc., is investigated through the models of ionic theory of slag, optical basicity, regular solution approach and molecular theory of slag. The best overall results are obtained by the model based on molecular theory of slag in which several operational parameters are also incorporated. The investigations of different slag samples, based on optical, SEM, EPMA and X-ray studies, reveal the effect of MgO and Al2O3on slag morphology and phosphorus distribution in different phases. It is important to consider phosphorus distribution ratio in the solid and liquid part of slag. The solid part of slag, which is mostly dicalciumsilicate, can contain up to 5 per cent phosphorus. The phosphorus content of the liquid part of slag may depend upon the phosphorus content of hot metal or phosphorus load of slag. It is found that incorporation of effect of dicalciumsilicate in the model improves the accuracy of prediction. For better process control, the addition of iron ore towards the end of blow must be avoided while treating high phosphorus hot metal or during the production of ultra low phosphorus steels. Keywords: steelmaking, BOF, phosphorus distribution, slag morphology, MgO, Al2O3,dicalciumsilicate, process control, models

Jan 1, 2004

By R. Allington, G. Walton

The approach has been applied with success to the restoration of chalk workings in southern England, and research and trials relating to restoration blasting and the replication of limestone dale landforms have been undertaken. An essential prerequisite for application of the technique is rigorous geomorphological analysis of the locality to ensure that final quarry landforms are consistent with the local geomorphological and landscape setting. It is always important to ensure compatibility of the final landform with the intended after-use, and to achieve this it is also essential that such restoration schemes have a sound geotechnical basis. A disadvantage of the approach is that final slopes tend to be flatter, especially in weaker strata where natural slopes are not steep; the implications for reserves may be unacceptable, particularly in quarry extensions

Aug 1, 1994

By P. A. Witt

"Highland Valley Copper is a conventional truck-shovel mining and flotation concentrator operation located 75 km southwest of Kamloops, British Columbia (Fig. 1).The ore is situated within a porphyry copper deposit and is mined at a rate of 135,000 mt/day from two separate open pits. The nearby mill produces concentrates containing 375 million pounds of copper and 5 million pounds of molybdenum metal yearly. In the Valley pit, head grades are roughly 0.40% Cu at a 0.23% Cu equivalent cut-off. The Valley pit strip ratio currently sits at 0.55:1 with waste quantities made up of separate rock and unconsolidated overburden units. From the mill, concentrates are trucked to nearby Ashcroft and then hauled by rail to ocean port facilities in Vancouver. The mine employs 950 people and is planned to operate until 2009. The operation is a partnership between Teck Cominco (63.9%), BHP Billiton (33.6%), and others (2.5%). Management is shared by the two principal owners and directed by a management committee."

Jan 1, 2005

By Hyung-woo Kim, Sup Hong, Sung-Soo Kim, Chang-Ho Lee

A deep-seabed integrated mining system, which consists of four subsystems such as riser pipe, buffer, flexible riser, and tracked vehicles, is presented in Figure 1. In order to collect more manganese nodules, multiple tracked vehicle systems should be used in a deep-seabed integrated mining system.

Sep 1, 2014

By T. Kerr, C. Mattich, J. Wilson

"In refractory gold concentrates, gold is occluded in sulphide minerals preventing efficient cyanide leaching. Hence, roasting is used to oxidize sulphur and organic carbon, the latter being necessary to minimize recovery loss through the preg-robbing effect. In 2009, Outotec commissioned the circulating fluidized bed roaster at Syama gold mine in Mali. Since 2013 Resolute and Outotec have been implementing improvements to the roaster performance with the capacity increasing from 8000 t/month in 2013 to above 13,000 t/month in 2016, while achieving a residual sulphide sulphur and organic carbon contents of below 0.1 wt.% and 0.4 wt.%, respectively. Thus a gold recovery in the CIL of above 90% is attained. In 2016, Outotec audited the Syama roaster. The analysis resulted in the Low Carbon Roaster (LCR) upgrade concept which (i) allows re-processing of calcine dust from the spray cooler and ESP after pelletization and (ii) allows the roasting of a further fine-organic carbon rich feed, i.e. de-slime circuit tailings. The LCR upgrade aims to reduce organic carbon to <0.2 wt.% and increase roaster-leaching gold recovery by >3%. Project development included extensive testwork. The LCR process is expected to become operational during FY 2019.INTRODUCTION Double refractory sulphide ores exhibit the characteristics of gold locked within pyrite grains, in the presence of preg-robbing natural (organic) carbon. Hence, to achieve high recovery during subsequent processing steps, i.e. cyanide leaching, oxidation of sulphide sulphur is required, producing a porous calcine by way of a roasting step. This allows for the gold contained within the pyrite grains to be accessible during cyanide leaching. Furthermore, it is necessary to oxidize the organic carbon. Oxidizing the organic carbon is essential as gold recovery loss occurs as a result of gold-cyanide complexes being adsorbed on organic carbon, this phenomenon is termed as preg-robbing (Miller et al., 2016). The gold absorbed on the organic carbon is lost, reporting to the tailings stream. Hence, the aforementioned oxidation of sulphide sulphur and organic carbon is the process goal. Nonetheless, the latter must occur at a controlled temperature which is concentrate/ ore specific, since the pores of the produced calcine should not sinter, as this makes the gold contained in the calcine less accessible during leaching, leading to gold recovery loss. The tight temperature control requirement is a decisive factor in the choice of a Circulating Fluidized Bed (CFB) as the roasting reactor, due to the excellent mixing characteristics of the CFB which results in temperature homogeneity of (+/-10°C). Roasting of double refractory ores, as described above, can be included in the flowsheet after comminution and grinding, a process termed as whole ore roasting, or by including a flotation circuit termed as concentrate roasting. Roasting the flotation concentrate results in less solid flow, increased sulphide sulphur grade and organic carbon content resulting in a higher heating value and higher gold content. A typical composition of the Syama ore and concentrate is given in Table 1. Ore average values presented correspond to the period 2015, while concentrate values are typical for the flotation performance throughout the period 2012–2015."

Jan 1, 2017

Lower Miocene Maryville Coal Measures at Waitewhena Coalfield, in the Taranaki coal region, include seams of sub-bituminous coal up to 4 m thick. During coalfield exploration Fletcher Challenge Ltd drilled ten cored holes, all of which encountered at least one seam. Coal quality is variable. Ash ranges from&nbsp;six to 12 per cent&nbsp;and sulphur from 0.5 to&nbsp;three per cent. A basal seam, overlying the Lower Mokau Sandstone, appears to extend throughout the area under investigation. A rider seam, of variable thickness, closely overlies the main seam in most drillholes. The rider converges with the main seam to the north and vertical trends in coal properties suggest that the seams are united in some locations. A less extensive seam sometimes occurs at the top of the coal measures beneath the Upper Mokau Sandstone. The lithostratigraphy of the intervening coal measures is complex, and coexistence of fluviodeltaic, lagoonal/lacustrine, and marine depositional environments is provisionally inferred. Lateral variations in coal properties (clastic derived ash, volatile matter, inertinite, hydrogen), thickness of the sedimentary parting between the basal seam and rider, texture of the coal measure sediments, and distribution of the upper seam all indicate persistence of a relatively low lying northeast-trending zone which may have been structurally controlled. Available data indicate that coal measure accumulation was terminated by marine transgression from north to south.

Jan 1, 1987

"The characterisation of resources and description of variability across different ore domains is key to accurate prediction of probable metallurgical performance. Such characterisation testwork often has limited drill-core from which to work and this immediately confines the characterisation method to smallscale. For flotation feeds, the JKMSI (Mineral Separability Indicator) is being developed as a potential characterisation tool, building on earlier work by Chudacek, (1990), which used shaken test tubes, and was called the EMDEE device. In addition, the consistency of flotation recovery for specified size fractions has shown to be preserved, irrespective of the amount of material in that size fraction obtained by different milling conditions, so that the overall recovery can be estimated from knowing the particle size distribution (McIvor and Finch, 1991; Bazin et al, 1994, Runge, 2007 and Wightman et al, 2010) The JKMSI unit uses less than 10 g of material per test in narrowly sized classes and results to date have demonstrated the value of the test. This paper discusses the potential of the test by demonstrating the link between JKMSI performance and mineralogy of five different ores from Laguna Seca, Escondida Minera Limitada as well as the correlation to plant scale performance obtained in an earlier survey (Coleman et al, 2006). Further work to tie down the proof of representatively of the small sample has been identified and is discussed and included in the paper. INTRODUCTIONSince the uptake of the flotation process at the turn of the twentieth century, small scale, laboratory methods have been sought to characterise flotation response. Although laboratory batch flotation tests are the most widely used method to evaluate reagents, characterise ore variability and obtain suitable model parameters using various scale up factors Runge (2101) other methods using less ore would be useful where ore samples were limited.A method using test tubes for testing ores’s flotation response to reagents was first developed by Ian Walk (Wark & Wark, 1935; Wark; 1938; Sutherland and Wark 1955). In these tests a 50 cc stoppered measuring cylinder served as a means for a qualitative experiment, the aeration being effected through violent shaking by hand. In these experiments about 10 g of mineral was ground by pestle and mortar under water after which the ground material was deslimed and adequately conditioned with the appropriate reagents."

Jan 1, 2013

By G. E. Perry, R. D. Seba

The steam soak process is the most widely applied and most successful thermal supplemental recovery process in use today. This process, which consists of injection of steam in various quantities into a well that is subsequently used as a producer, can be repeatedly applied to the same well. The treatment, if successful, will increase the production rate of the well for a given period. For this reason, the steam soak process should be considered primarily as an acceleration method rather than as a method for producing oil that could not be recovered by primary means. However, application of this process to reservoirs containing viscous crude may produce oil that otherwise would have zero present value or would be produced at such low rates by primary methods that it would not be profitable. A major necessity for the successful application of the steam soak process is a source of natural energy. This source of energy may exist in one or more of the following forms: expansion of fluids, compaction, gravity drainage. This study was undertaken to determine the effect of gravity drainage on steam soak production and recovery and to develop a method for predicting production rates, ultimate recovery, and heat requirements when applying this process to a reservoir in which the primary driving mechanism is gravity drainage. (For additional discussion of some of the factors that affect recovery of viscous oils by the steam soak process, see Ref. 5.) Model The model developed for analyzing the performance of the steam soak process in gravity drainage reservoirs is a modified form of that developed by Matthews and Lefkovits.1,2 The model used in this study consists of two concentric cylindrical volumes of productive formation with cap and base rock of impermeable material. The inner cylinder, extending from the radius of the wellbore to some predetermined radius, is referred to as the hot zone; and the outer cylinder extending from the exterior boundary of the hot zone to the outer drainage boundary will be referred to as the partially heated zone. Fig. 1 is a schematic cross-section of the model, indicating its thermal properties. Horizontally the reservoir is divided into two regions, the upper region having a high gas saturation and residual liquid saturation, and the lower region containing a high liquid saturation. The boundary between these two regions is referred to as the free surface, which drops and changes shape as fluid is produced from the reservoir. The model was used to analyze the steam soak process by applying the following assumptions: 1. The horizontal formation is composed of one or more homogeneous noninterconnected layers containing incompressible fluids. Each layer is treated separately. 2. The hot zone is heated by the initial soak and is maintained at the steam temperature corresponding to atmospheric pressure by resoaking. 3. The height of the free surface at the producing well is at all times zero (producing well is pumped off). 4. The position of the free surface at the start of

Jan 1, 1970

By Maria I. De Rosa, Charles D. Litton

The National Institute for Occupational Safety and Health (NIOSH/PRL) conducted a series of large-scale experiments to evaluate the effectiveness of optical flame detectors, photoelectric smoke detectors, and combined ionization and photo­electric smoke detectors for rapidly detecting mining equipment cab fires. The detec­tor alarm times were then used to trigger the discharge of a fire inerting system inside the cab to suppress cab material fires. This paper discusses the types of fire detectors tested, the experiments that Were conducted, and the results that were obtained. Con-" elusions are that rapid detection of equipment cab fires can be achieved to trigger the discharge of a fire inerting system inside the cab to protect the operator in the cab.

By Stephen T. Young

The construction of Phase 2 of Hong Kong Station, its retail podium and International Finance Centre Two, the tallest building in Hong Kong, will be completed in the middle of 2003. This paper describes the design and construction of the slurry wall for the Station (both Phase 1 and Phase 2) and the two commercial buildings at the Station. The slurry wall is used as the permanent earth support wall for the station box and the excavation depth ranges from 70 feet to 120 feet. Top-down construction technique was adopted to support the excavation of the station, while circular cofferdams relying on hoop stiffness were constructed to support the basement excavation of the two towers. The designed behavior of the slurry wall is also compared to the as built behavior.

Jan 1, 2003

By D. R. Brox, J. Levell

The Montana Tunnels Mine was mined out to a depth of 900 feet at the end of 1999. The first signs of instability were recognized in 1996 along the upper slope area of the northwest sector when the pit was mined out to a depth of approximately 500 feet. To extend the life of mining operations, it has been proposed to extend and steepen the northwest highwall to a depth of 1000 feet. Slope stability analyses were carried out using the UDEC software program to assess the stability and deformation of the proposed highwall design. Back analysis of an instability in 1997 along the upper slope section and highwall deformation to date has indicated reasonable agreement with the observed mode of failure and displacement monitoring results. The results of detailed analyses of the proposed highwall design have indicated very limited additional deformation, further development of tension cracks is unlikely and large-scale instability is not expected to occur. A total of 40% of additional mining has been completed and ongoing displacement monitoring to date has confirmed very limited additional movement and no further tension crack development.

Jan 1, 2001

By Michael Gauna

An evaluation and analysis of a coal pillar bump in a West Virginia mine indicated that the pillar bump occurred because of the unique combination of controlling factors. The pillar?s development stability factor of 2.6, based on Mark-Bieniawski strength and tributary loading, would typically indicate an adequate pillar size for long-term main entries. A combination of multiple-seam mining geometry and geological conditions resulted in stress concentration leading to the coal pillar bump. No high extraction mining had been performed in the area and the overlying works appeared to have adequate stability factors. However, it was interpreted that long-term flooding had softened the overlying pillar system and rendered all but the largest pillars ineffective. Contributing factors include large overlying pillars transmitting stress through the approximately 60 ft of interburden and the failed pillar?s location at a transition in the immediate roof from shale to hard, massive sandstone. Other contributing factors include sandstone channeling forcing the development section to mine the pillar smaller than surrounding pillars in an attempt to avoid the channel, the relatively high overburden of 1,100 ft, and the pillar?s position in the bottom of a trough within the coal horizon. Pillars subjected to similar conditions on the other side of the section did not fail where the roof was composed of shale, even though the roof exhibited evidence of high stress concentration. An analysis was undertaken to determine appropriate pillar size for future mining. The conditions were evaluated by investigating mine historical conditions, assessing pillar bearing pressures, and using Lamodel 2D coupled with traditional Mark-Bieniawski pillar strength and tributary area loading stability factors to determine the stress levels for the pillar. Recommendations for future pillar sizes were based on the critical load bearing capacity identified by the analysis. To date, no further pillar bumps have occurred.

Jan 1, 2008

By R. Scott

This article describes and evaluates the establishment of Kuyasa Mining including the search for a suitable deposit, the ongoing lack of finance, and market constraints. Despite the dramatic initial growth and success of their Ikhwezi Colliery near Delmas, Mpumalanga, Kuyasa?s experiences offer some sobering arguments for significant changes to mineral rights ownership, the attitude of financial institutions and government policy if South Africa is to have a dynamic and successful junior and small-scale mining sector.

Jan 1, 1998

By D. Peatfield

The design and operation of coal preparation plants are governed by the inherent quality of the raw coal to be processed, market specifications and the saleable tonnage requirements. Topography influences the layout of the plant but does not affect the process design. The coal market comprises many sectors, each having specific requirements with respect to both size and quality. In turn, these have led to different levels of preparation and the development of specific types of coal preparation plant design. Developments in processing techniques and processing equipment have also influenced the design and now give the preparation engineer numerous permutations in process and equipment selection. This paper gives an overview of the coal industry in general and summarizes some of the developments in the design of plants and technical developments in the areas of dense medium processing, fine coal treatment and equipment design.

Jan 1, 2003

Jan 1, 2000

By B. M. Nyombolo

The interest shown in our paper by Drs Schaekers and Dew, and their favourable comments, are appreciated. However, some of their remarks require a response. Stability testing procedures Regarding the issue of stability-testing procedures, we concur that no long-term procedure has been accepted by either environmentalists or the mining/processing industry as a valid method to test the long-term stability of solid wastes. We agree that the method we used is no exception, and we certainly did not claim that it should be recognized as such. What we did attempt to portray was that we had designed the test procedure in an attempt to simulate, at a very small scale, a tailings dam environment.

Jan 1, 2001

By James J. Tansey

A Deep Foundations Institute meeting with the University of Michigan at Ann Arbor in March of 1984 featured legal rights under Differing Site Conditions. A workshop conducted by the authors and Gerard Ittig, Esq. added to the value of the symposium. This paper will be of interest to your attorneys in any subsurface disputes.

Jan 1, 1985