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By T S. Giraudo

The Macraes Gold Project, operated by GRD Macraes Limited, engaged GRD Minproc Limited to assist with the evaluation of a pressure oxidation circuit, this being the key component of a recovery enhancement project. The initial study gave sufficient confidence in the process to enter into project execution. GRD Macraes Limited was the first company to license Newmont Mining Corporation's pressure oxidation technology, which aims to improve gold recovery from the carbonaceous sulfidic flotation concentrate. This paper outlines the fundamental chemistry of the process followed by a description of the plant that was ultimately installed. A summary of the commissioning is given together with an overview of the production results to date. The project ramp-up was rapid, with design temperature and throughput being achieved within days of hot commissioning. Gold recoveries have exceeded expectations, improving from around 70% prior to the installation of pressure oxidation to above 84%. The project has increased the nominal gold output at Macraes from 120,000 oz/y to 170,000 oz/y. On-going plant optimisation for process parameters is discussed, giving rise to flexibility in gold production, with key controlling parameters exceeding design.

Jan 1, 2000

By J. R. Stone, J. T. Roy

ASARC09s continuous tapper for lead blast furnace is described. Its use throughout the company's plants has resulted in higher production rates, lower labor costs, and better working conditions. It has also permitted wide latitude in blast pressures, as well as in charge and slag compositions. The practice of tapping a lead blast furnace has remained essentially unchanged for many years. In principle, it is a batch operation discharging the products from the continuous smelting of lead-bearing materials. As such, it imposes certain restrictions on the proper functioning of the smelting operation. First, it calls for a highly developed sense of timing, and a considerable amount of physical effort on the part of the operator. Dirty and hot working conditions, as well as the hazards of flying metal and slag, usually prevail. These conditions have resulted in higher labor costs, both in rate and in amount. Labor turnover has also been high. Secondly, rigid control of mechanical and metallurgical factors must be maintained. Frequent oxygen lancing of the tap hole is needed to permit tapping, thereby increasing operating costs. Rapid increases in smelting rate, plugged tap holes, or forgetfulness on the part of the operator may permit the molten material to rise above the tuyeies and freeze the furnace. Also, the operator will often blow out the tap hole after draining, filling the area with fume. Composition of charge also vitally affects the tapping. The type of lead-bearing material being smelted will naturally affect the smelting rate, and thus the frequency of tapping. The amounts and types of fluxing materials, which determine the ultimate fluidity of the slag, must be regidly controlled. Finally, the batch tapping of a continuous smelting operation of this type appears to be wrong in principle. The molten material drains into the crucible at the bottom of the furnace where it starts cooling immediately and continues to cool until tapped out. While keeping the lead molten is no particular problem, maintaining a fluid slag at all times is often difficult, and sometimes impossible. These factors all point to the desirability of a method for the continuous tapping of a lead blast furnace. The expected advantages would be more desirable working conditions, lower labor costs, increased production, and a much wider latitude in both blast pressure and charge composition. CONTINUOUS TAPPING Historical. A number of devices, e.g., mechanical tappers, siphons, and lead wells, have been tried in the past in attempting to effect continuous tapping. None proved to be entirely satisfactory for tapping lead and slag simultaneously. It was not until 1955, however, that ASARCO designed and put into operation a successful tapping unit at its East Helena, Mont. lead smelter. This unit is described in detail in U.S. Patent NO. 2,890,951. Continuous Tapping Unit—See Figs. 1 and 2. Fig. 1 is a schematic drawing of the unit in place, showing front and side views. The component parts are identified in the caption. Fig. 2 is an actual photograph, front view, of the tapping unit attached to the base of the blast furnace. Operation. The operation of the tapping unit is quite simple. Essentially it consists of a narrow tapping bay surrounding the furnace tap hole, with a deep notch in the front wall. When the blast furnace is started this notch is open down to the level of the tap hole. Starting the furnace may be accomplished in several ways but the most satisfactory method is to plug the tap hole with clay, permitting metal and slag to build up inside the furnace. Before the liquid reaches the tuyeres, the plug is removed and the furnace is drained to the settler. The blast is then stopped momentarily while the notch in the front wall of the tapping bay is dammed with chrome brick and plastic chrome ore. The blast is again started, and the height of the dam is adjusted so that the liquid in the bay will counterbalance the internal pressure of the furnace. When the desired level is reached the slag and metal will overflow to the settler. It should be noted that the molten material ,is not siphoned from the furnace and that the metal 'acts as a liquid valve controlling the flow of slag through the tap hole. During normal operations the tapping bay is filled primarily with lead, covered with a thin layer of slag. Thus, the heavier lead plays the major role in compensating for variations in furnace pressures. Due to the high specific gravity of lead, changes in the liquid level in the tapping bay will never exceed a few inches. For this reason it is seldom necessary to change the height of the dam due to minor variations in operating conditions. Some turbulence of the liquid in the bay is essential at all times. This indicates that the level of the

Jan 1, 1963

By Oswald J. Heinrich

THE origin of spontaneous combustion in collieries is, of course, chiefly due to bad system in laying out the pits, unclean workings, insufficient ventilation, and neglect in damming off works after they are exhausted of the main coal. This is particularly indispensable in the Richmond field, because it is difficult and dangerous here, on account of the inferior roof, to remove the last vestiges of coal ; and we are compelled also to retain all the slates and bands, as far as possible, in the pit, to help fill up, although such material is liable to fire. This, however, can be prevented to a great extent, as mentioned below. I am satisfied that the only system of working this thick coal is by cross-cutting (Querbau, ouvrage & travers) or working in benches, as practiced in France (Creusot, Rive de Gier, St. Etienne), in Silesia (Kiinigsgrube), which I have used as far as it is practicable in an old pit, almost torn to pieces by former workings. Modifications may, of course, be needed, even in the same pit, on account of irregularities in the seam. I am also satisfied for economical reasons (on account of the heavy cost for timbering) that pits here ought to be worked upwards, even if it requires a longer period of time to win the coal, and a greater outlay of capital to commence with. The various saddle-shaped " troubles " making natural minor basins in the main basin aid in the execution of this plan. A pit worked once ought to be worked out clean, given up, and dammed off forever. If sections of proper height are assumed, according to the required amount of production for a certain period of time, they ought to be worked out and filled in as much as possible, dammed off, and ultimately allowed to fill up with water, in order to protect the upper works in future from fire in the gob below. This I consider indispensable, because we may not be able profitably to fill in all the ground. To do so we have to use the waste of the mine, although it is liable to fire in time. In using this stuff, it ought to be surrounded by walls of rock not liable to take fire, laid firmly in clay, the gob being in the middle, and firmly covered and packed with clay on top. In

By Michael Radnor

The purpose of the study was to 1) examine the social and economic impacts of surface mine blasting in relation to variations in conditions of terrain and geology, population density and type, levels of ground motion and air blast; 2) consider technical, legal and government policy issues; and 3) examine the potential, nature and extent of surface mine blasting impacts and its regulations upon both communities and surface mining companies. Phase One designed the empirical research and technical analyses of Phase Two, focused upon surface coal mining and quarrying in the Eastern U.S. Several broad conclusions are evident: 1) blasting is not the major issue for communities or surface mine companies; 2) while the number of instances of severe cracking or annoyance are relatively small, the impact can be severe for individuals immediately involved; 3) blasting/drilling cost curves tend to rise significantly at intervals below 1.0 particle velocity, in/sec, dependent upon the size of a permit area and the size of a mine; and 4) the legal, complaint, cost and density analyses do not suggest the need for a general lowering of present regulations.

Jan 1, 1981

By Larry Patts, Jon Hummer, Jim Noll, Aleksandar Bugarski1, George Schnakenberg, Steve Mischler, Mike Crum, Shawn Vanderslice, Rick Anderson

This study is organized under the auspices of the Metal/Nonmetal Diesel Partnership formed by the National Institute of Occupational Safety and Health (NIOSH), the National Mining Association (NMA), the National Stone Sand and Gravel Association (NSSGA), the United Steel Workers of America (USWA) and the MARG Diesel Coalition. The ultimate objective of the study is to reduce the exposure of underground miners to diesel particulate matter (DPM) and gases, and to help in fulfilling the partnership’s goal of identifying technically and economically feasible controls to curtail particulate matter emissions from existing and new diesel powered vehicles in underground metal and nonmetal mines. The majority of current knowledge on the performance of the diesel particulate filter (DPF) systems and other control technologies is based on studies done under laboratory conditions, the results then being applied the results to on-highway vehicles. According to the best knowledge of the authors, only two comprehensive studies that were conducted recently at Noranda’s Bathurst Mining and Smelting Mine and International Nickel Company’s Stobie Mine under the sponsorship of the Diesel Emissions Evaluation Program [McGinn 2001, Bugarski and Schnakenberg 2001, Bugarski and Schnakenberg 2002] offered some insight into the problems associated with the deployment of modern DPFs to underground mining vehicles. The U.S. mining industry has expressed concern that this rather limited knowledge base is not sufficient to help them comply with the rule limiting DPM exposure of underground metal and nonmetal miners [30 CFR 57.5060 2001]. The partnership agreed that a series of comprehensive field evaluations of DPFs in several underground mines was needed to determine their viability of DPF systems and establish confidence in their performance. The first in this series of studies was the one conducted in Stillwater Mining Company’s Nye Mine, Nye, Montana. This study was conducted in two phases. The objective of the first phase was to establish the effectiveness of the selected technologies in reducing diesel emissions by using an isolated zone methodology. The objective of the second phase was to assess the effectiveness of diesel particulate filters in controlling the exposure of underground miners in actual production scenarios. As part of the first phase of the study a series of tests were conducted at Nye Mine from May 19, 2003 until May 30, 2003. The results of Phase I of the study are presented in this report.

Jan 1, 2005

By McNicol M. D

Petroleum exploration in onshore Gippsland was directed in the 1950s and 1960s towards plays in the basal Latrobe and Strzelecki Groups, but these became forgotten in the 1970s with the Latrobe play discoveries offshore. The principal explorationists of the time were Woodside Oil and Arco Ltd, who then departed the scene after drilling some 18 wells without major success. In the late-1980s successful hydrocarbon finds in equivalent age Otway Group sediments in the Otway Basin were made and in early 1990 Tyers Petroleum began investigations into Strzelecki Group prospects as applied to the onshore Gippsland Basin. Some 18 anticlines have been mapped, based on shallower coal exploration drilling in the Latrobe Group, which from seismic and gravity evidence are believed to unconformably overlie similar anticlinal structures in the underlying Strzelecki Group. Strzelecki targets of interest within drilling depth are the basal Tyers Conglomerate/Rintouls Creek Sandstone Formations which outcrop in the north of the Latrobe Valley and are analogous to the Pretty Hill Sandstone in the Otway Basin. In addition, intra-Strzelecki Group targets similar to Heathfield and Windermere Sandstone Members are also considered to be present. Good source rocks are abundant including coal seams and carbonaceous sediments of Neocomian and Jurassic age, which lie within the oil window. Seal is provided by the overlying Aptian-Albian aged volcanogenic sediments. The Tyers/Rintouls sandstones - a target in the Woodside/Arco wells, with two exceptions, were not reached. Encouraging intra-Strzelecki shows were reported in many of the wells. Basal and intra-Strzelecki porous quartzose sandstones remain one of the few untested plays left in the Gippsland Basin and, in the onshore, merit extra attention as they he within the oil window at economically drillable depths.

Jan 1, 1992

By Zuhair A. Nawab, Muhammad A. Al-Marhqun, Mehmet Guney

"The Atlantis-Il-Deep is a stratified metalliferous deposit located along the median valley of the Red Sea at a water depth of about 2200 m. The metal-bearing mud contains sulphides of zinc, copper and iron with significant amounts of silver, gold and cobalt. A variety of oxide, siliceous and carbonaceous gangue material is also present. Long sediment cores indicate that the thickness of the deposit may reach 25 m, but the average kriged thickness of metal-rich sediments is about 11 m.Mining geostatistics was applied to estimate the Red Sea offshore mineral resources. Experimental semi-variograms were computed along 0 ± 180 degrees and then models were fitted to analyze the mineralization phenomenon. Kriging estimates indicate that 696.330 million t of bulk sediments contain 1.891 million t of zinc, 0.425 million t of copper and 3. 75 thousand t of silver. Gold and cobalt contents were calculated from the analysis of solids in flotation concentrates as 47 t and 5368 t respectively. The significant outcome of implementing geostatistics to the Red Sea brine precipitates is not only for the evaluation of metal resources, but it also provides a useful tool in the optimization of selective dredging of appropriate areas to be mined. IntroductionThe discovery and extensive exploration of hot brines and hydrothermal sediments in the Red Sea median valley revealed the fact that the Atlantis-II-Deep is a potentially exploitable metalliferous mud in the economic sense. Sediments in the mineralized deep are composed mainly of oxides, various types of silicates and sulphides with considerable concentrations of zinc, copper, cadmium, silver and gold. The mineral-rich deposit, located at a mean water depth of about 2000 m, has an average thickness of 11 m, and is covered by a brine layer having a temperature of 60°C."

Jan 1, 1988

The development of Western Mining Corporation's first commercial gold heap leaching operation is described. The laboratory testwork requirements for determining the amenability of ore to heap leaching techniques is discussed. Heap construction and operating practices are described; and production statistics and operating parameters are presented. It is shown that for a specific low grade gold ore a treatment cost optimization can be performed. Taking cognizance of the ore grade, percentage gold recovery, operating costs, and the gold sale price, it can be determined whether Carbon-In-Pulp (C-I-P) or Heap Leaching should be employed as the extraction process. The examples presented in this paper refer specifically to Kambalda and should not be used in themselves as generalities.

Jan 1, 1987

By Hon-Yim Ko

The properties of coal from the Pittsburgh seam, procured from the Bruceton and Federal No. 2 Mines were found to be similar. Strength and stiffness properties of Beehive Mine coal were determined and the material found to be isotropic. Mechanical, index and geologic properties of roof and floor coal measure rocks from Beehive mines were measured and correlated. Analytic formulations are given for coal creep under uniaxial loading, while triaxial creep is being investigated. Results of testing of coal model structures are compared with predictions from using several material models.

Jan 1, 1976

By Robert E. McIvor

"The Functional Performance Equation for ball milling circuits allows us to isolate the circuit classification system efficiency (CSE) from the mill’s grinding performance. The factors which affect circuit CSE are reviewed, with emphasis on water usage at the classifiers. It is shown that manipulation of such water usage can be carried out to improve plant grinding circuit efficiency, at times by more than 30 percent.THE FUNCTIONAL PERFORMANCE EQUATIONObservation of the feed and product size distributions going into and discharging from the mill in grinding circuits with low versus high circulating load ratios initially led to the definition of “circuit classification system efficiency” (or CSE, McIvor, 1988a), and ultimately the discovery of the Functional Performance Equation (McIvor, 1988b, and McIvor et al, 1992). The CSE is the percentage of “coarse” (typically with reference to the plus P80) size material occupying the mill (versus “fines” or product size material, finer than the circuit P80), and can be calculated by taking the average of the percentage of coarse material in the mill feed and mill discharge. As this also represents the percentage of the mill energy expended on targeted, coarse size material, it is directly proportional to overall grinding circuit efficiency and production rate. This is all clearly expressed in the Functional Performance Equation for ball milling circuits, which is derived as follows.We define “fines” as any product size material, and “coarse” as that targeted to be further ground, the two typically differentiated by the circuit target P80. For any grinding circuit, the production rate of new, product size material (fines) must equal the specific grinding rate of the coarse material (i.e. fine product generated per unit of energy applied to the coarse material) times the power applied to the coarse material."

Jan 1, 2009

The prime objective of geometallurgy is to improve the profitability of mines through the use of spatial models of rock properties that have a significant impact on value. Although a key property is the grade of the component of interest, because it directly influences the revenue from saleable material, increasingly attention is being focused on other attributes. These 'non-grade' variables may have a less direct, but equally profound impact on value, either by reducing potential revenue (eg lowering recovery, deleterious elements, etc) or increasing mining or treatment costs (eg comminution, grindability, throughput, reagent use, etc). Collectively, these variables have been described as 'geometallurgical'. There are three main areas of activity required to realise the value of geometallurgy: 1. acquisition of appropriate data through sampling and measurement, 2. building spatial models of geometallurgical variables, and 3. using these models to optimise value. Each of these is built on the foundation of the preceding step. In building reliable spatial models, it is critical that the amount of error introduced to data by the sampling and measurement processes is understood and quantified, in order that the inherent variability of the component of interest can also be properly understood. Management of the quality of grade variables is now applied (with different levels of diligence and success) across nearly all sectors of the mining industry, not least because of the necessity to adhere to Codes of practice (eg JORC, 2004). The necessity for quality management and at least the broad framework of what to do in practice is reasonably understood by many relevant professionals. There are also a number of tools, including software products, statistical analyses and presentation formats, available for those managing grade data. It is also abundantly clear, that in many cases people have little understanding of why they are doing particular quality control activities, or how to sensibly interpret the results. The same is not true for many geometallurgical variables. In almost all instances the authors have seen, the magnitude of sampling and analytical/measurement error variance embedded in geometallurgical test data is untested and unknown. This is partly because until relatively recently, metallurgical data has mainly been collected for the purpose of process and mill design, with few, expensive tests carried out on 'average' samples of the orebody. Increasingly however, there is a move towards collection of much larger data sets from spatially distributed sampling, and thus a growing necessity to understand the errors inherent in both sampling and analysis of these types of variables, and how to manage them. In order to understand how the quality management practices applied to grade variables can be extended to geometallurgical variables, it is essential to understand the theoretical foundation that supports these practices. This requires an analysis of the concepts that underpin the 'principles of additivity' (Matheron, 1963), the 'theory of sampling' (Gy, 1982, Pitard, 1993), and 'statistical process control' (Shewhart, 1930). Relevant concepts are then integrated to form a foundation to develop methods to improve the quality management of geometallurgical variables. Application of this framework will significantly increase the integrity of geometallurgical data, the reliability of geometallurgical models and ultimately the value derived from implementing geometallurgy. This is an ABSTRACT ONLY no paper was prepared for this presentation.

Mar 1, 2010

By Daudi R. Waryoba

Keywords: Wire Drawing, OFHC Copper, Recrystallization, Orientation imaging Microscopy (OIM), Texture inhomogeneity Textural inhomogeneities have been investigated in oxygen free high conducting (OFHC) copper wire drawn at room temperature to a true strain of 2.31, and isothermally annealed at 250°C and 750°C for annealing times ranging from 10 s to 1 hr. Local orientations were mapped by means of orientation imaging microscopy (OIM). The microtexture of the drawn wires showed a strong <111> + weak <100> duplex fiber texture at the inner core, whereas the mid and surface regions had a comparatively weak texture. Annealing at 250°C resulted into a recrystallization which originated from the mid section, proceed towards the surface regions and ends in the inner core. Recrystallization resulted into a strong<100>+weak<111> duplex fiber texture. A similar textural inhomogeneity was observed during short annealing at 750°C. However, prolonged annealing gave rise to abnormal grain-growth that proceed from the inner core to the outer surfaces with a dominant <111> fiber component at the inner region and mixed components of <111>, <100>, and <112> at the outer surfaces.

Jan 1, 2005

By L. W. Vollmer

Introduction During the past three years at least fifty strings of 9% nickel steel tubing have exhibited excellent resistance .to the variably severe corrosive conditions typical of many of the sweet condensate wells in Mississippi, Louisiana, and Texas. This experience, and the characteristically outstanding performance of other ?nickel alloy steels under hydrogen sulphide corrosive conditions, dictated the selection of 9% nickel steel tubing for the second well drilled in the sour condensate Pincher Creek field of southwestern Alberta. The damaging effects of the Pincher Creek reservoir fluids on well equipment had been manifested in the discovery well by the rapid and not yet fully understood failure of API Grade N-80 tubing, stainless steel wire lines, and alloy steel fishing tools. Two very unexpected failures in the 9% nickel steel tubing, after only six days of service in production tests, instigated an extensive and continuing investigation to ascertain the probable cause of failure. This discussion presents some of the results obtained thus far which cast a little light on the perplexing behaviour of steels in hydrogen sulphide environments in general, and in sour condensate wells in particular. History of Tubing Failures Walter Marr well No. 1, in which the 9% nickel steel tubing failures occurred, produces from the Madison lime at the depth interval approximately 12,200 to 12,750 feet. The gas-oil ratio is about 26,000 and the gas contains 10.2 per cent hydrogen sulphide and 6.3 per cent carbon dioxide by volume.

Jan 1, 1952

By Oswald J. Heinrich

(with figures on plate V.) THE origin of spontaneous combustion in collieries is, of course, chiefly due to bad system in laying out the pits, unclean workings, insufficient ventilation, and neglect in damming off works after they are exhausted of the main coal. This is particularly indispensable in the Richmond field, because it is difficult and dangerous here, on account of the inferior roof, to remove the last vestiges of coal ; and we are compelled also to retain all the slates and bands, as far as possible, in the pit, to help fill up, although such material is liable to fire. This, however, can be prevented to a great extent, as men¬tioned below. I am satisfied that the only system of working this thick coal is by cross-cutting ( Querbau, ouvrage à travers) or working in benches, as practiced in France (Creusot, Rive de Gier, St. Etienne), in Sile¬sia (Königsgrube), which I have used as far as it is practicable in an old pit, almost torn to pieces by former workings. Modifications may, of course, be needed, even in the same pit, on account of irregu¬larities in the seam. I am also satisfied for economical reasons (on account of the heavy cost for timbering) that pits here ought to be worked upwards, even if it requires a longer period of time to win the coal, and a greater outlay of capital to commence with. The various saddle-shaped "troubles" making natural minor basins in the main basin aid in the execution of this plan. A pit worked once ought to be worked out clean, given up, and dammed off forever. If sections of proper height are assumed, according to the required amount of production for a certain period of time, they ought to be worked out and filled in as much as possible, dammed off, and ulti¬mately allowed to fill up with water, in order to protect the tipper works in future from fire in the gob below. This I consider indispen¬sable, because we may not be able profitably to fill in all the ground. To do so we have to use the waste of the mine, although it is liable to fire in time. In using this stuff, it ought to be surrounded by walls of rock not liable to take fire, laid firmly in clay, the gob being in the middle, and firmly covered and packed with clay on top. In

Jan 1, 1873

The final criterion of the payability of gold production must necessarily be the cost per ounce. While the cost per ton or per cubic yard is a convenient comparative measure, and a necessary one for operating control, it is obviously insufficient as a means of judging whether or not, in a given set of circumstances, a gold production operation can be carried out economically.. An alternative method, which brings in the ton or cubic yard as a unit of measurement, is to determine what the cost per ton or per cubic yard is equivalent to, in terms of pennyweights or grains at the ruling price of gold. But this is really only another way of stating that the cost per unit of product is the ultimate measure of profit or loss, and not the cost per unit of input material.It is accordingly of interest to determine the cost per ounce of Australian gold production and to compare it with the cost of gold production in some other parts of the world which are important gold producers. The attached Appendix, which it, is believed comprises practically all Australian mines of any consequence (except Big Bell and the de Bernales group, for which. insufficient information is available), shows this information for financial years ending between 31st December, 1938, and 31st December, 1939, the production of gold dealt with being some 1,084,000 oz., or 66% of the total Australian production for 1939. For comparative purposes the cost of producing gold at representative mines in New Guinea, New Zealand, Fiji, Canada, India, West Africa and Rhodesia, and the average cost of production at the Rand mines, has also been included.It is obviously necessary in a study of this sort that the basis of the cost should be clearly set out and should be as nearly as possible uniform. The &quot;cost&quot; as calculated herein comprises the following items:

Jan 1, 1940

Mineral liberation data are routinely obtained for a wide variety of applications, including ore characterization and characterization of various metallurgical test products obtained in bench-scale laboratory testwork. Newmont Mining Corporation owns several properties in the Carlin Trend containing refractory gold ores. These ores are generally difficult to treat and have to undergo several processing stages for gold recovery. At Carlin?s Mill 5 operation, flotation is used to pre-concentrate some of these difficult ores to produce a sulfide concentrate to feed the roaster. It is of interest from a flotation perspective to be able to study the floatability of pyrite, arsenopyrite and arsenian pyrite, since these minerals all have different quantities of gold associated with them, and have to be recovered to ensure a successful separation has been achieved. In this paper, a detailed MLA study on several Carlin ores was undertaken on samples from flotation tests, using air or nitrogen as flotation gas, specifically to study the differences in floatability characteristics, which ultimately impacts overall gold recovery. I

Jan 1, 2009

Geopolymerization is a process by which Al-Si minerals undergo dissolution and polymerization in an alkaline solution to form a monolithic material with a high mechanical strength (Davidovits, 1991; 1994). In this study, Na-feldspar (albite) and K-feldspar (microcline) have been selected to investigate their abilities to undergo geopolymerization. The interaction between alkali metals present in the alkali-feldspars and alkali metal cations contained in alkaline solution during geopolymerization has also been studied. It has been found that both albite and microcline can be applied in geopolymerization to produce geopolymers with a high compressive strength. Although the dissolution of M-feldspar (where M = Na and/or K) in concentrated MOH solution is inhibited when the alkali metal cations of the alkaline solution differ from the alkali metals structured in M-feldspar, the resulting geopolymers are found to demonstrate a higher compressive strength. Moreover, a proper molar ratio of Si/Al in the geopolymeric gel phase and a higher average hardness of the undissolved particles remaining in the geopolymers are found to greatly increase the ultimate compressive strength of the geopolymers.

Jan 1, 2002

A major problem in open pit mine planning is to determine the optimum pit limit of the mine. The optimum pit limit is the one with maximum discounted net profit and satisfies technical conditions. A number of optimisation techniques have been proposed for determination of pit limits. Among these techniques, those based on graph theory are suitable to solve the three-dimensional (3D) problem. In this paper, two maximum flow methods named, Dinic&apos;s algorithm and Push-Relabel algorithm have been used to find the optimum pit limit of an iron ore mine. This method has been implemented in Gol-e-Gohar iron mine in south east of Iran in order to find the final pit limit. Issues, like properties of the mine and orebody, are studied. These properties include wall slope, sulfur and phosphor grade restrictions. The result shows that these properties have significant effect on the determined pit limit.

Dec 6, 2010

By Dean Bray

During the past three years, a new rear-dumping dragline bucket technology has been patented, developed and tested at North Dakota coal mines. Field testing has shown that this new concept in dragline bucket technology can improve strip-mining efficiencies by reducing cycletimes and increasing dumping radius. This article details the process, participants and the results for the development of a rear-dumping dragline bucket. Participants in the process In November 1995, the board of directors of Dakota Coal authorized the execution of a field test and option agreement and a license agreement with the partnership of D.F. Schmidt, electrical/maintenance supervisor, and Bill Lengenfelder, president of Modern Industrial. The board also authorized entering into a dragline bucket field-testing agreement with the Falkirk Mining Co.

Jan 1, 1999

By Thomas M. Barczak

The era of the shield-supported longwall face began in the United States a little over 25 years ago. The most fundamental development of the shield in the past 25 years has been an increase in size and capacity and a progression toward two-legged designs in favor over four-legged shields. Is the “bigger-the-better” design philosophy justified? The advantage of the two-legged design is largely attributed to its active horizontal force caused by the forward orientation of the leg cylinders. How do we know this? You might be surprised to learn that shield loading data may suggest otherwise. In addition, what about setting pressures? Should they be set as high as possible? Some people have advocated that setting pressures should equal the yield pressure. Does this make any sense? Is yielding a good or bad thing? These are some questions still worthy of investigation that are addressed in this paper. The first major change in gateroad support was the use of cable bolts to replace conventional wood cribbing particularly in Western mines. Standing support changed very little until the mid-1990s, when Strata Products1 introduced the Hercules and Propsetter supports in Eastern mines and Burrell Mining Products introduced the Can support in Western mines. Since then, there has been a dramatic increase in new roof support technologies for longwall gateroad support. Historically, standing support, including shield supports, have been designed based on a simplistic model requiring the support to have sufficient capacity to support the gravity loading of a perceived detached rock mass. This approach ignores the stiffness of the support and the resulting ground deformation that is known to be critical to roof stability. Recently, a new approach based on the ground reaction concept has been promoted by the National Institute for Occupational Safety and Health (NIOSH) to account for the interaction of the support with the rock mass behavior in achieving ground stability. In addition to gateroad support design, this concept has ramifications for shield capacities, setting pressures, and recovery room support design that challenge current practices in these areas. Although longwall mining has matured into the safest and most productive underground mining method used today, traditional concepts of roof support design and practice should still be evaluated and challenged. Ultimately, advancements in the science of ground control are made by recognizing deficiencies in current theories and thinking outside the box.