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By Dale S. Preece

INTRODUCTION The Waste Isolation Pilot Plant (WIPP) is a research facility located in Southeastern New Mexico. The WIPP is being developed by the U. S. Department of Energy (DOE) to demonstrate the safe storage of defense-related radioactive wastes in bedded salt. Since the WIPP is a research facility. a number of large-scale in situ experiments have been planned and are currently under construction (Munson, 1983). These experiments were designed to study various aspects of nuclear waste storage in bedded salt such as mechanical responses and creep closures of drifts, thermal response due to heated canisters, and thermally-induced fluid migration. One purpose of the WIPP is to develop and demonstrate a general predictive structural analysis capability for a bedded salt repository. One set of experiments, called the Geomechanical Evaluation (Room G), is heavily instrumented to study the creep around several rooms with different configuration. A Layout of the Geomechanical Evaluation room is shown in Figure 1. One portion of it consists of wedge shaped pillar where two drifts intersect at an angle of 7.5 degrees. The pillar can be seen at the bottom of Figure 1. One purpose of the wedge pillar experiment is to study progressive pillar failure in the tip region. Another is to determine how nonuniform pillar thickness affects creep closure of the drifts. An important aspect of the experiments is the correlation between experimental data and corresponding pretest finite element analyses of the site. The finite element simulations serve two purposes. First, the computer simulations aid in understanding the experimental results by providing calculated stress. strain and displacement fields that cannot be measured directly. Second, comparison of calculated and measured drift closures serves as a method for validating or improving the finite element models and the constitutive models employed. The wedge pillar geometry required a 3-D finite element creep calculation. Results from these 3-D calculations will be presented in this paper. A 2-D plane strain double drift model which approximates the wedge pillar geometry at a slice perpendicular to the drift has also been performed and will be compared to the 3-D results in this paper. FINITE ELEMENT COMPUTER PROGRAMS Two finite element computer programs. JAC (Biffle, 1984) and JAC3D (Biffle. 1986), were used. JAC is a 2-D finite element program developed for quasi-static analysis of non-linear solids. It employs the conjugate gradient iterative technique to obtain a solution Spatial integration is performed using a single gauss point in each four node quadrilateral element. An hourglass viscosity technique is used to control the zero energy modes that occur with single point integration. The single point integration combined with the explicit nature of the program and exploitation of CRAY-1 computer architecture results in very efficient execution J4: results were compared to results obtained with eight other 2-D structural creep computer codes in the second WIPP benchmark exercise (Morgan, 1981) JAC3D was derived from JAC to treat 3-D finite element models and has many of the same characteristics including single point integration and hourglass viscosity. These characteristics have made 3-D creep analyses more reasonable by significantly reducing computation time The first exercise of the creep capability

Jan 1, 1986

By D. Savile Davis

D. Savile Davis* (Member): The members of the Chamber of Mines Research Laboratories, headed by Dr Cook, indeed deserve all the plaudits that have been given for the original conception of the rock-cutter, and for the research work that has gone into designing the equipment now being tested. Underground trials with the rock-cutting machines at the Doornfontein Gold Mining Company Limited have been encouraging to the extent that 29 square fathoms of rock have been cut in the past three months with two modified prototype machines. Both the hangingwalI condition of the stope and the control of stoping width have been excellent. So far, however, the rate of production has been too slow and does not compare with conventional stoping methods, due firstly to the time-consuming operation of moving the machine forward and resetting it in its next cutting section and secondly to mechanical defects. Although great strides have been made in improving the tungsten carbide cutting tool, further progress is dependent upon engineering development in the design of a more reliable and versatile machine within economic limits. Commenting briefly on the experiments that have been carried out on the Gold Fields mines, the rock-cutter, when used at the Luipaards Vlei mine, some 600 ft below surface, cut very nearly perfect slots above and below the reef on a stope face but was unable to cut slots in the reef itself. At the Doornfontein mine, however, where the tests are currently being carried out at a depth of 8,000 ft below surface and where there is a much greater pressure acting on the rock face, vertical fracture planes can be clearly seen in the hangingwall running parallel to the face. Owing to the presence of these fracture planes it has been found impossible to cut slots of any depth and the action of the machine is more one of spalling than of cutting, resulting in slabs of combined reef and waste falling from the face as the cutting tool advances. This is accentuated by the existence of a well-defined hangingwall parting plane. Most of the slabbing takes place from the slot cut below the reef to the hangingwalI of the stope. Since waste sorting in narrow conditions and poor lighting is not an efficient operation, the reef together with a proportion of waste rock, has to be removed from the stope. The ripping action of the machine as opposed to slot cutting under these particular conditions may not be a great disadvantage, as the volume of waste which can be packed depends on the ratio of reef width to stope width and with a narrow reef body some waste has to be removed to the surface. In these circumstances mining at depth with this type of machine should not be viewed as selective mining but more correctly mining without explosives. Another type of rock-cutter is being tested by the Gold Fields Group, consisting of a diamond studded rotating disc, which has, to date, cut slots 10 ft in length and averaging 7 1/2 in. in depth. It is, however, too early to comment on the practicability and economics of such a cutter. The great interest that is at present being shown in breaking rock without the use of explosives warrants comment on the Robins raise-borer, recently introduced into South Africa, and used for the first time with great success at the Doornfontein mine.

By Peter Keppler, Charles Jackson

WHY DO AN AUDIT? There is increased enforcement of environmental laws and regulations at State and Federal levels with attendant steep fines and even the possibility of jail sentences. We are all aware of environmental and public health and safety incidences (i.e., Love Canal, Bhopal) which have led to stringent laws and regulations (i.e., CERCLA, RCRA, SARA). In some cases, the courts have specified that a company institute an audit program as a condition of settlement of claims which related to lack of environmental control. The U.S.EPA issued an Environmental Auditing Policy Statement in the November 8, 1985 Federal Register (pp. 46504-46508) which encourages the use of environmental auditing by regulated entities. Elements of what EPA considers to be an effective environmental auditing program are outlined and briefly explained. These factors combine to indicate why there is a growing number of companies which have instituted an audit program. Briefly, the program can and should be used to monitor compliance with environmental regulations and corporate policy, to prevent violations, to improve policies and procedures, to raise the level of environmental awareness, to encourage communication among operations and all management levels about environmental problems, and to measure risks and control costs. AMAX Inc. has had an environmental audit program since the early 1970's. The AMAX Environmental Affairs Department, working with the Law Department, oversees and coordinates environmental audits and analyzes the findings which an audit reveals. Facilities are audited on a regular schedule; those having the greatest potential for environmental liabilities are audited at least annually; other facilities are audited less frequently. The question of who performs the AMAX audits was settled in favor of the audit team being independent of personnel from the facility being surveyed. Other methods of auditing could be the following: - Independent contractor (Appendix A) - Combination of independent contractor and in-house staff - Self-audit by site personnel The merits and demerits of who should conduct the audit can be analyzed to suit your particular needs. Next you need to decide on the procedure for the audit and the form for reporting and follow up. Environmental publications referenced in Appendix B can be helpful to you at this point. After deciding on the form of auditing, a schedule is prepared and followed to the closest extent possible. This schedule should be published and distributed to the sites so facilities know when they are being audited. This is not designed to be "a game of hide and seek." Various elements of the AMAX Audit Program are as follows: Team selection Pre-visit preparation Site visit Exit interview Report Using the audit Follow-up Team Selection The usual make-up of the team is the audit manager, the environmental specialist who normally provides assistance to the site or has knowledge about the operation, and a legal representative if his or her schedule allows. The legal representative will review the findings of the audit team if a personal visit is not timely or necessary. Size of the team can be expanded, but a minimum of the audit manager and the environmental specialist is best. Two sets of observations are better than one. The environmental specialist can also provide the rapport which is beneficial to creating the appropriate tone so that the site people will be as cooperative as possible. This is not to be an antagonistic exercise.

Jan 1, 1991

By John R. Freeman

The die pressing of brass may be described as a method of producing irregularly shaped parts of brass and other copper alloys by hot deformation in a die under pressure. Die pressing of brass was first developed as an improvement on sand castings in order to obtain irregularly shaped articles of brass with superior surfaces, a closer tolerance in dimensions, a denser structure, freedom from porosity and elimination of the relatively large amount of scrap invariably attached to sand castings. The freedom from sandy surfaces with consequent greater tool life in machining operations was also recognized as well as elimination of losses due to the uncovering of blowholes in machining operations. Die pressing also gave alloys of superior strength. Die pressings today are made from extruded rod or shapes, according to design of the part to be made. The greater number of pressings are made from round rod, because of its relatively lower cost and ease of handling. The rod is first sheared or sawed into desired lengths, the diameter of rod used and length of "slug" or "rivet" depending upon ultimate weight and shape of the die-pressed part to be made and design of die. The dimensions of the slug are figured so as to give the minimum amount of scrap to be sheared from the pressing as it comes from the die. Extruded shapes having I, T or X sections are often used when the finished part is to have a similar shape or heavy sections are joined by relatively light sections of considerable length. A rod bent in the shape of a U is also used on occasion, for similar reasons. Proper shearing of the slugs is of considerable importance. The sheared ends of the slugs should be as flat and straight as possible. Improperly sheared ends having an angle to the axis of the bar or a rough, uneven surface are liable to give folds or laps in the finished pressing and proportionately high rejections. The property of shearing varies considerably with different alloys and should be considered in the selection of a material for a given part. Where satisfactory shearing is not possible, the slugs may be cut from the bar by power saws. Extruded

Jan 1, 1931

By N. H. MacMillan, O. C. Dias

Recent work by Westwood et al. (reviewed in Refs. 1-6) has established that many rocks, minerals, ceramics, and glasses exhibit phenomenologically similar chemomechanical (Rebinder) effects when deformed in different liquid environments, apparently as the result of chemisorption-induced variations in their near-surface flow and flow-dependent fracture behavior. In particular, it appears that there exists an empirical, but apparently general, correlation between the near-surface flow stress of an inorganic nonmetal and its surface charge- namely, that such a solid is hardest at its zero point of charge (i.e., when its zeta potential [ ] = 0) and softens as [ ] increases, regardless of the sign of zeta. Moreover, because near-surface flow and fracture phenomena are two of the important determinants of the rates of machining and comminution processes, both Westwood et al. and others7 have been able to use this so-called "[ ]-correlation" to develop cutting fluids and grinding aids that in many instances have given rise to several fold increases in performance. In addition, Westwood et al. have demonstrated that both aqueous and organic environments can give rise to similar improvements in cutting rate for a given combination of tool and workpiece, provided that both impart the same surface charge to the work- piece (i.e., transfer the same number of electrons to or from its surface). This last finding implies that in any application it should always be possible to find some cutting fluid that is at once effective, cheap, nontoxic, and nonpolluting. Thus, reports of apparently similar environmental influences on the mechanical behavior of diamond,8 certain organic crystals and polymers,9-14 and coa1-15-21 suggest that it is pertinent to ask: (a) whether the [ ]-correlation (or any analogous relationship) extends to coal; and (b), if so, whether this relationship can be used to guide the development of superior aids for grinding coal. Accordingly, correlations have been sought between (1) the zeta potentials, (2) the hardnesses, and (3 and 4) the drilling and grinding characteristics of coals of different rank in different aqueous and organic environments. This paper reports the results of that search.

Jan 1, 1980

By H. H. Otto

Fourteen years ago, J. B. Warriner presented before the Institute a paper on anthracite stripping,' describing the progress of stripping in the Anthracite Region from its beginning with an old quarry stripping at Summit Hill, which was operated more than a century ago, through various types of strippings used in 1917. This paper will briefly review the status of anthracite stripping methods from 1917 to the present day and describe in detail the stripping operation of The Hudson Coal CO. at Clinton collicry. It is not intended as a review of the work in the district—that would require a much longer paper. History and Statistics Mr. Warriner said, in part: "The sizes of shovels have gradually increased until now 70 and 80-ton shovels are the rule and larger equipment is coming in." He also referred to a large electrically operated dragline which weighed about 225 tons and had an operating radius of 125 ft. In 1917, a number of strippings used shovels of various sizes, ranging up to the 60 Marion and Bucyrus 70-C railroad type. Material was hauled away in narrow-gage side dumpers, usually of 4 or 5 cu. yd. capacities. Star-type churn drills with 4-in. bits were generally used for drilling purposes. Since 1917 wages in the anthracite field; have increased, but the higher operating costs of strippings are now being successfully combated by continuous improvements in stripping equipment, such as the development of the caterpillar or crawler shovels, the use of larger cars, better drilling equipment, the introduction of shovels of increased capacity and the adoption, in a number of instances, of standard-gage equipment for transportation. In 1918 another large dragline was introduced in the southern fields, which had been in service in the Middle West and is in use today, dipping vertically about 180 ft. to raise coal out of a stripping. About 1918, small shovels with caterpillar traction were introduced. In 1922 the 6-in. well drills began replacing the 4-in. drills, with an attendant saving to the operators. About the same time A. E. Dick placed the first pair

Jan 1, 1931

By G. Capobianco

"The coal industry is an important part of Canada's economy - as an energy source, employer, and generator of export earnings. Coal comprises 71% of our fossil energy and with rapidly advancing handling and burning technology will clearly play a major role in our national energy mix, eventually as synthetic gases and liquids.International coal markets are very competitive and Canada is disadvantaged by long rail hauls to port, and a heavy tax burden. However, high mine productivity has allowed us to increase our markets.Currently, domestic markets are dominated by coal use for electrical generation, a market with significant growth opportunities.Strong growth is expected in export thermal markets, while metallurgical coal demand is projected to show little or no growth in the medium term. IntroductionThe coal industry is a vital part of the Canadian energy scene .• Coal supplies 17% of Canada's energy.• On a per capita basis, Albertans use more energy in the form of coal than they do oil and gas.• The value of Canada's exports of coal last year was equal in value to all the cars we imported from Japan.• Coal is our largest single export to our seco nd largest trading partner, Japan.• The value of coal produced in Canada exceeds the value of any other single mineral produced in thi s country.• In Alberta, the safety record for coal mines is on a par with that of department store clerks.• The coal industry employs about 25% of the total number of people employed in all metal mines in Canada, directly and indirectly accounting for about 50 000 jobs.Coal comprises 71% of Canada' s hydrocarbon energy resources on a crude oil equivalent basis . This is more than double our oil sands resource and dwarfs our conventional crude oil and gas resources (Fig. I). In the short term, coal as a solid fuel plays a significant part in the Canadian energy scene. In the longer term, coal, in the form of electricity and eventually liquids and gases, has a tremendous role to play. In 1986, Canada produced 57.8 million tonnes, slightly down from the record high of 60.8 million tonnes in 1985. Our exports in 1986 were 26.0 million tonnes, a growth of 53% in three years."

Jan 1, 1988

By R. Kumar, N. R. Mandre

"The beneficiation of iron ore slimes by selective flocculation was studied. Polyacrylamide (PAM) and guar gum were used as flocculant, along with sodium hexametaphosphate (SHMP) as a dispersant. The samples collected were initially subjected for chemical and mineralogical studies. The mineralogical studies indicated the presence of haematite and goethite as the principal iron-bearing minerals, with quartz and kaolinite as gangue minerals. The selective flocculation tests showed that the Fe grade can be enhanced from 58.24% to 64.60% Fe at a recovery of 66.33% using polyacrylamide as the flocculant. A Fe grade of 63.20% at a recovery of 68.04% was obtained using guar gum as flocculant. The efficacy of the studies was analysed using a separation index, which indicated that the higher separation index was obtained with the tests employing polyacrylamide as the flocculant. IntroductionIndia is endowed with huge reserves of highgrade iron ore and is one of the prominent producers in the world. As the high-grade iron ores have been exhaustively mined, it has become imperative to use low-grade iron ores, fines, and slimes to meet the growing demand. Also, generation of fines and slimes during mining and processing contributes to loss of minerals to the extent of 30% of the total value (Mandre and Panigrahi, 1997). These fines/slimes are often discarded as waste into tailing ponds, containing considerable amounts of iron. Therefore beneficiation of tailings/slimes has become necessary for optimal utilization of the resources.In recent years, interest in recovery of valuable minerals from low-grade ores, slimes, and tailings has increased worldwide. These activities are initiated not only to recover minerals but also to address various environmental issues associated with fine particle treatment. The literature reveals that several technologies have been employed in this regard, including flotation, advanced gravity separation, and high-intensity magnetic separation. However, these processes are found to be ineffective for fine particle treatment due to lower collision probability in flotation, decreased gravitational effect in gravity separation, and reduced drag force in magnetic separation, among other factors (Batisteli and Peres, 2008; Das et al., 2007; Dworzanowski, 2012; Rocha, Caneado, and Peres, 2010). Therefore, it is in this regard that selective flocculation has gained importance in treating slimes in recent years."

Jan 1, 2017

By W. A. Haley

Wheel loaders became important bulk materials handling tools in the early fifties. As they grew in stature, productivity, reliability, and acceptance, the demand for larger, more rugged machines evolved. In the early sixties, the introduction of a 6-yard model by a major equipment manufacturer led quickly to widespread experimenting and rapid worldwide acceptance in loading blasted rock in stone quarries and smaller metal mines. As the wheel loaders grew in popularity for shot rock loading, larger models, 12- to 15-yard size were developed and saw widespread use by 1970. In the past 20 years, at most quarries and small to medium size mines, most cable operated power shovels have been replaced by large wheel loaders and more recently, hydraulic excavators as the primary loading tool. Wheel loaders larger than the 15-yard class are not used widely. Until about 1970, very few overburden haulage trucks were used in domestic coal mines. About that time, many states started enforcing stringent mine relcamation regulations and in 1977, even more stringent federal legislation was passed. These controls brought about drastic changes in coal surface mining methods in the mountainous areas of Appalachia, in smaller midwestern mines and in the large mines of the western states. Basically, simple material casting methods were replaced at most eastern mines by "haul-back" and "mountain top mining" methods using controlled placement of overburden as it is removed from the coal. These methods dictated much more extensive use of loaders/ shovels and trucks to the point that most coal mining in those areas is now by loader/shovel trucks methods. And those areas have become the home of the country's largest population of large wheel loaders and trucks in the 35- to 85-ton size class with a trend to even larger trucks and shovels. To maintain equal production or provide for growth, the new methods have caused a sharp increase in machine intensity to the extent that at some mines, up to three times as many machines were required to mine a million tons of coal in the late seventies as compared to the machine needs in the sixties. Mining costs and coal prices also took a big jump in the mid-seventies, paralleling the drastic price increases in oil and other forms of energy.

Jan 1, 1984

By David Cole

AT the time flotation appeared upon the metallurgical horizon in Arizona, the writer, under the direction of Dr. Ricketts, was engaged in remodeling and enlarging the No. 6 Concentration Plant of the Arizona Copper Co. at Morenci, and the work had been in progress nearly a year before the Inspiration experiments with flotation had disclosed the revolution in concentration that was at that timè impending. The simplified flowsheet worked out for the remodeling of the Morenci plant had been based upon the removal of the freed metal in a minimum number of stages by treatment upon tables equipped with Butchart riffles, the latter being adapted to accomplish both classification of feed and removal of the metal at one operation, substantially as described in my previous paper.1 The scheme of treatment for the slime was based upon the well-known fact that after copper sulphides, such as chalcocite and chalcopyrite, etc., are reduced to a certain extremely fine state of comminution they are definitely beyond the reach of separation upon any of the concentrating devices then known. By using drag-belt classifiers (which served as conveyors as well as separators), the overflow would be of the usual "slimes" class. Experiment had shown that when these drag-belt overflows were, properly diluted (that is, to about 5 to 7 per cent. solids in the feed under treatment) . the very fine sand, and especially the very fine but still granular sulphide particle's, would, if given a short distance to fall, quickly settle out in prime condition to yield an excellent recovery on vanners; a further valuable feature would be that the very fine non-separable final overflow tonnage might be discharged direct to tailing, thus conserving space requirements so as to permit within the old building, without embarrass-ment, the increase of capacity required.

Jan 9, 1916

By Carlos F. Perucca, Maria Virginia Ramírez A.

"Chile based SQM S.A., exploits potassium chloride rich salts at the Salar de Atacama, in northern Chile. The process includes the solar evaporation of natural brines existing in the Salar salt bed, followed by conventional flotation to recover Muriate of Potash (MOP) for fertilizer use.This paper describes a structured method developed to aid in understanding the metallurgical response of the solar evaporated salts to conventional flotation. The method uses classical metallurgical data analysis combined with laboratory and plant-scale testing, and rudimentary mineralogical studies.As a result of the study, several changes were made to the flotation circuit which produced an operation that is much more consistent and easier to control, and has an increased recovery. An optimized circuit, currently being financially evaluated, has also been conceptually designed based on the findings of this work.INTRODUCTIONSQM S.A. is the world’s largest producer and distributor of specialty fertilizers, iodine, lithium, as well as a wide array of industrial and specialty chemicals. SQM is a public corporation whose shares are traded in the Santiago and the New York Stock Exchanges.SQM S.A. exploits the mineral rights at the Salar de Atacama in northern Chile. Its production processes involve the solar evaporation of natural brines existing in the Salar salt bed, aimed at the recovery of Potassium, Lithium and Boron compounds."

Jan 1, 2004

By A. W. Stradling, M. J. Fairweather

"The selective flotation of lead, zinc and iron sulphide minerals is key to the economic performance of concentrators treating zinc-lead ores. Diagnostic metallurgy is a tool to quantify mineral behaviour as a function of particle size and liberation. The result of such analysis may be used to select preferred strategies to improve metallurgical performance. The zinc flotation circuits at three Teck Cominco operations: Sullivan, Polaris and Red Dog are compared here. This paper analyses operating data from each of these circuits. The effects of mineralogy and performance by particle size are emphasized. Metallurgical performance at these three concentrators varies substantially. Liberation size is the major factor in determining metallurgical performance. Complex mineralogy requires a fine grind to achieve effective liberation. Fine grinding results in offsetting decreases in concentrate grade through entrainment. The economic treatment of complex fine-grained deposits requires optimisation of regrind size for liberation without unduly compromising concentrate grade through gangue entrainment.INTRODUCTIONUntil recently Teck Cominco Metals Ltd. operated three mines treating zinc-lead ore. In order of increasing mineralogical complexity they are: the Polaris Mine on Little Cornwallis Island in Nunavut in the eastern Arctic, the Sullivan Mine in southeastern British Columbia and the Red Dog Mine in northwestern Alaska. Both the Polaris and Sullivan Mines have recently ceased to operate due to depletion of ore reserves.The ore type and associated gangue, mineralogy and iron solid solution vary significantly between these deposits. Table 1 summarizes some of the important differences"

Jan 1, 2003

By C. Guney Olgun, Charles Elks, Sherif L. Abdelaziz, James R. Martin, Christopher Fox, Pier Luigi Iovino, Fatih Catalbas, Pierre Gouvin

"Energy piles are a new innovative renewable energy technology to access and exploit the relative constant temperature of the ground for efficient heating and cooling of buildings. Examples of energy piles include drilled shafts, continuous flight auger (CFA) piles, driven piles, and micropiles. A field test setup was installed at the Virginia Tech Geotechnical Research Facility to study energy piles. The field test consists of a total of five micropiles 25 cm (10 inches) in diameter, four of which were equipped with circulation loops. The test piles extend to a depth of 30.5 meters (100 feet) and are heavily instrumented with strain gauges and temperature sensors. Several observation boreholes were formed around the test piles to monitor ground temperatures. Thermal conductivity tests and field load testing were performed to investigate the behavior of Energy Piles. This paper presents results of the thermal conductivity tests and thermo-mechanical field load test performed as part of this field investigation.IntroductionThere is a developing trend around the world to explore alternative energy sources. The main driving forces are growing global energy demand, depleting natural resources and the adverse effects of greenhouse gas emissions from fossil fuel consumption. Geothermal energy is one of the promising renewable sources that can be utilized to offset such trends. However the use of geothermal energy has been limited mainly to certain hot spot areas where it is used either for district heating and/or electricity generation. Ground temperatures below a depth of about 6 meters (20 feet) remain stable compared to outside air temperatures, typically lying between 8-23°C (46-73°F) at most U.S. regions. The constant temperature and heat storage capacity of near surface soils in any region represent a tremendous potential of stored energy that can be used for heating and cooling of structures."

Jan 1, 2017

By Peter N. Blakey, Jean-Paul Paquette, Barton A. Thomson

"FROM THE BEGINNING OF THE PIT, it was known that the surface mining method had a projected life of about ten years, and that the ore extended to depths recoverable only by underground mining. When serious consideration was given to underground mining in 1968, it was realized that tremendous advantages could be gained by starting the underground mine as early as• poss1ble. These included: 1. The availability of underground ore for grade control, which would become more difficult as the pit deepened. 2. The need to develop new systems, equipment and skills to take full advantage of the technological revolution in underground mining. 3. Early knowledge of rock behaviour through trial mining on a small scale prior to finalizing plans for major blocks. In addition, the pit would be available to guarantee the mill supply if problems were encountered. 4. With sufficient time and early mining experience a much better job could be done on all mine planning. Because ample reserves of economic underground ore were virtually assured, the desired underground mining capability became a function of two restraints: first, the time required to excavate and construct the permanent headframe, shaft and ore-handling facilities necessary to handle large volumes of stope ore; second, the time for lateral and stope development on at least one level which would normally follow shaft sinking. The decision made was to move immediately on the permanent hoisting installations and at the same time embark on the planned ramp from surface, which would allow much stope preparation to be completed prior to .bringing the shaft on stream. Both projects were placed on an urgent basis early in 196S' and the following decisions made shortly thereafter."

Jan 1, 1974

By Ehsan Alavi, Brian Hagan, Colin LaVassar, Glen Frank, Derek Dugan, Rick Capka, Lisa Mori

"Tunnel face instability is one of the most common issues in EPB tunneling. The evaluation of an ideal face support pressure is a critical element in the design and construction phases of EPB tunneling. In this study, measured tunnel face pressures in three EPB tunneling projects (University Link U230 and U220 projects and the Northgate Link Extension N125 Project) for Sound Transit in Seattle, WA are compared with the calculated analytical methods. This comparison shows that the analytical methods generally overestimate the required tunnel face support. The best practice for managing the face pressure in EPB tunneling operation is discussed based on the experience gained through these projects. Furthermore, the algorithm that is developed for identifying ground loss in these projects is addressed. INTRODUCTION Earth pressure balanced tunneling relies on using muck within the chamber to support the face. Muck is pressurized by controlling the advance rate of the tunnel boring machine and the speed of the screw conveyor. Target face pressure is typically calculated by using the geotechnical assumptions in the geotechnical baseline report. The concept behind EPB tunneling is that the application of sufficient face support pressure will prevent tunnel face instability and the resultant over excavation. Tunnel face instabilities caused by a lack of sufficient face support can lead to ground loss and surface settlement (Anagnostou & Kovári 1996, Kirsch 2010). A certain support pressure must be achieved to counteract water and earth pressure with an adequate factor of safety. While the water pressure can be predicted accurately, the determination of the resultant earth pressure on the face is less precise. This becomes manifested in the relatively high partial safety factors on the earth pressure."

Jan 1, 2016

MIM Holdings Limited (MIM) committed to a long-term future in copper in the Carpentaria Mineral Province in 1996 with the development of the Enterprise and Ernest Henry mines, and an upgrade of MIMsÆ copper smelter in Mount Isa, and copper refinery in Townsville. The upgrade of the copper smelter will enable the smelter to produce 250 000 tpa of copper with a single primary smelting unit, the copper Isasmelt furnace installed in 1992. The copper Isasmelt furnace currently smelts concentrate containing 145 000 tpa copper. To reduce the risk associated with the copper smelter upgrade, and shorten the ramp up time of the upgraded smelter, two high-rate trials were undertaken in 1997 to demonstrate that the copper Isasmelt furnace could, with an additional 525 tpd of tonnage oxygen, treat concentrates containing 250 000 tpa copper. During the trials the concentrate feed rate to the Isasmelt furnace was increased from its average rate of 100 tph of concentrate to 150 tph of concentrate. The increase in the capacity of the furnace was achieved by: supplementing the installed tonnage oxygen capacity of 595 tpd with an additional 300 tpd supply; reducing variability in the furnace feed rate from ¦25 per cent to less than ¦5 per cent; and reducing tramp air. The Isasmelt furnace smelted concentrate at a rate of 140 to 150 tph for four hours 25 minutes during Trial 1, and at a rate of 150 tph of concentrate for five hours and 30 minutes during Trial 2. The oxygen utilisation was 100 per cent during the trials with the process remaining under control. As a result of the development work for the trials the average treatment rate of the furnace under normal conditions increased by ten per cent.

Jan 1, 1998

By L. Xia, B. Hart, G. Gingras, M. Furlotte

Plant and laboratory test work was performed in order to optimize a copper/zinc flotation separation circuit at Matagami concentrator, Canada. The testing identified a significant difference between the laboratory testing and the plant flotation, particularly for the sphalerite recovery in the rougher copper concentrate. Tests indicate that sphalerite recovery of the flotation concentrate in the laboratory testing was significantly higher than in the plant. The ToF-SIMS surface analysis identifies surface species responsible for the discrepancy in Zn recovery. It shows that sphalerite flotation is in response to the collector attachment and a significant enrichment of Cu on the surface of sphalerite grains. A comparative surface chemical analysis was performed on both sphalerite and chalcopyrite grains from the plant flotation and laboratory testing in order to identify surface species variability between the two testing. Two significant observations were revealed. First, there is considerably more hydrophilic iron oxyhydroxide on sphalerite surface from the plant flotation sample relative to those from the lab testing. The secondary coating likely completely or partially passivates sphalerite surfaces, reducing Cu enrichment and hence sphalerite floatability. Secondly, comparing the surface chemistry on sphalerite and chalcopyrite, the hydrophilic iron oxyhydroxide species have preference for the sphalerite grains relative to chalcopyrite. Despite numerous studies on the effect of galvanic coupling during grinding, there is limited documented evidence on the effect of the milling environment on the flotation and plant operations. This observation suggests that interaction between sulphide minerals and grinding media have a significant influence on subsequent flotation. This paper discusses the results of a detailed surface chemical evaluation of sphalerite grains from both the plant and laboratory testing in the context of grinding environments and pulp chemistry.

Jan 1, 2016

By H. F. Zurbrigg

Introdduction The main line of ascent in the development of the nickel industry may be traced back through the International Nickel Company of Canada, Limited, to its earliest antecedents, the Canadian Copper Company and the Orford Copper Company. Mining geology is but one aspect of the overall development of this organization, and the stress it has placed on geology grew according to the need. In the early days, intensive geology was not required because ore supply was not one of the main problems of this or any of the pioneer companies. The difficulty then was to find economic methods for reducing the ores to nickel and copper, and markets for absorbing nickel in the face of a limited demand from consumers heavily prejudiced in favour of the New Caledonia product. "Very early in the history of the region it was patent to everyone that to share in the full benefits of the industry the same individuals or company must control the whole of the operation necessary to manufacture the finished product' (1, p. 45). These conditions were fulfilled for the International Nickel Company in 1902 when the Canadian Copper Company (incorporated 1886), with its ample ore supply, took the initiative to merge its interests with those of the Orford Copper (Refining) Company (incorporated 1887) to form the new company which, with reorganizational changes, emerged as the producer of the overwhelming bulk of the world's nickel. The Mond Nickel Company, Limited, entered the field in 1899 with a refining process, and over a period of years acquired ore through purchase of mining properties. This Company operated successfully and independently until 1929 when its assets were joined with those of the International Nickel Company in order to permit the most economic exploitation of the Frood-Stobie orebody in which both Companies were interested.

Jan 1, 1948

By P Rubie, B Morgan, J Russell, C Stewart

The relationship between efficient liner exchange, concentrator availability and concentrator profitability is well documented. Russell Mineral Equipment’s (RME) relationship with OZ Minerals’ Prominent Hill site personnel enabled sound collaboration between the two parties. Both parties worked together to identify the sources of productivity and efficiency loss and developed solutions to reverse them.This paper chronicles the identification of production inefficiencies and the resulting removal of mill relining from the critical path through the materialisation of mill relining optimisation strategies. Furthermore, this paper outlines other recent mill relining optimisation strategies developed by RME and tested at Prominent Hill, to demonstrate and achieve further productivity, efficiency and safety improvements.CITATION:Rubie, P, Russell, J, Morgan, B and Stewart, C, 2017. Mill relining optimisation – a story of collaboration, in Proceedings MetPlant 2017, pp 335–345 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 11, 2017

By A. M. Tweedy, Felix A. Vogel

Since the publication of Prof. J. W. Richards's paper on The Schumacher Briquetting Process,' this process has been in operation on a practical scale in two plants in the United States, and a few notes on the results obtained may be of interest. An experimental plant, consisting of a toggle press and a hydraulic press, was installed at Johnstown, Pa., together with the necessary bins and other equipment, and an attempt was made to briquette a mixture of Mahoning fines and blast-furnace flue dust, in a ratio of 7 parts of ore to 3 parts of dust. The results were unsatisfactory, in that the briquettes made were not hard enough to stand the severe handling to which they were subjected in the various transferences necessary between the press and the furnace. The trouble was due to several causes: First, the mechanical arrangement of the plant, which did not lend itself to a proper hardening and loading of the briquettes into the cars. Second, the irrrgular composition of the flue dust, as to both its lime and its fine coke contents. Third, the variability of the moisture content of the ore, and its viscosity when wet. Fourth, the ratio of fine ore to flue dust; that is, it was attempted to bind in too large a quantity of fine ore for the small amount of flue dust used. The first trouble could be overcome, and has been improved upon in the plant as it stands to-day. The second trouble was important, because of the small amount of dust used as a binder, and might have been easily removed by mixing the flue dust of the different furnaces, which, however, was never done. The third difficulty was more serious, but could have been partly solved, had it been possible to meet this along with the fourth difficulty, by decreasing the percentage of ore in the briquettes. It was proposed to do this, but it seems that the problem at the works in question is, primarily, to prepare Mahoning fines for reduction in the blast furnace, and only secondarily to use up the flue dust, and a solution along this line did not seem to be interesting.

Jan 1, 1914