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By T. J. Yeomans

"The Kidd Creek metallurgical site is situated twenty-six kilometres east of Timmins, Ontario. The site consists of a 4.5 niillion tonne per year concentrator, a 127,000 tonne per year electrolytic zinc plant and a 79,000 tonne per year copper smelter/refinery complex.The concentrator commenced operations in 1966 and was expanded to its present capacity in 1978. The zinc plant commenced operations in 1972, and the copper smelter/refinery in 1983.The concentrator tailings, and both liquid and solid waste products from the other plants are disposed of within a 1215 hectare tailings disposal area. Here, a method known as centraljzecl thickened discharge is used to form a stable cone-shaped hill. A low permanent perimeter dyke structure encloses the cone and directs cone run-off and thickener overflow to a treatment pond. After treatment with lime slurry, this neutralized effluent is recycled to the metsite and also discharged into the Porcupine River system.The tailings system has been in operation in its present form since 1972, upon completion of dykes and a 110 metre thickener unit. Underflow densities were continuously increased until thickener overloading became a serious problem. This necessitated the construction of a second 110 metre thickener, which was completed in 1983. Target densities have been achieved without difficulty since then.This paper discusses the operation of the tailings system, reviews past performance, and outlines future operations."

Jan 1, 1986

On the more important mines it is almost essential to have a general plan of the surface. A special feature of this plan should be the indication thereon of a suitable arrangement of contour lines, either wholly over the surface of the property or at parts where may seem most useful. The importance of having these contours marked on the plan will be best appreciated by the engineer, in selecting the best available site for a proposed new plant, in locating a tramway, in connecting the surface with the underground workings to the best advantage, or in solving many similar problems that come within his province. The closer the contours are together vertically the better, the limitation being the trouble and expense of getting them on to the plan. They may range from five to fifty feet apart but a distance of ten feet will be found, in most cases, to be the most useful.In deciding the datum level from which these contours are to be measured, it is best to depart from the ordinary sea level datum, and use instead the highest available point on the property. This datum will be called zero and all levels will be reckoned so many feet below it; thus by a contour marked 160 on the plan, it will be understood that the level of the plane of the contour is 160 feet below the highest available point on the property. This system of measuring levels downward is readily suggested from measuring the depths of shafts from the brace or of excavations from the original surface.

By Joe Roby

Rolling disc cutters are the business end of hard-rock and mixed-face tunnel boring machines (TBMs). Since their first successful employment on a TBM more than 50 years ago, disc cutter technology has constantly evolved, allowing modern TBMs to excavate very hard and abrasive rock efficiently. Disc cutters have also been employed successfully on earth pressure balance (EPB) and slurry machines, cutting rock under water and ground pressure. The range of materials excavated by machines today is broader, excavation rates are higher and cutter costs are lower than ever before, proving the value of investment in cutter development. Disc cutters are used on a range of tunneling equipment, from pipe jacked slurry micro-TBMs less than 1 m (3.3 ft) in diameter to 15 m (49 ft) in diameter, hard-rock boring TBMs. The geological conditions under which they are employed range from sands and gravels, with several bar of water pressure, to extremely hard, massive rock with UCS up to 420 MPa. Regardless of the type of machine or geology, one thing remains constant: changing worn out cutters is costly. When cutters must be changed in the middle of a tunnel, the contractor incurs the cost of downtime as well as the cost of refurbishing or replacing the cutters. In the case of catastrophic cutter failures, the project can be stopped completely for long periods, and the costs mount up rapidly while tunnel production is at a stand still. Catastrophic cutter failures include the occasion when hard-rock disc cutters are failed in groups (called a ?wipeout? phenomenon) and the operator fails to stop the machine, which results in severe damage to the cutterhead. The cause of this can be either an undetected failed cutter propagating damage to surrounding cutters or operator error in steering. Regardless of the cause, the resulting damage can take days or even weeks to repair. When cutters fail prematurely on EPB or slurry machines in certain geological conditions, it is impossible to evacuate the chamber and, therefore, impossible to get [ ] into the chamber to change the cutters. The solution is frequently an unplanned intervention shaft that must be sunk at great cost. The importance of being able to predict cutter life accurately, and in all geological conditions, cannot be overemphasized. For these reasons, cutters have been developed for specific machine types and sizes, as well as for specific geological operating conditions. Different size cutters are required for different size machines. For example, while 483- and 508-mm (19- and 20-in.) cutters are used on large diameter TBMs, it is impossible to employ cutters so large on a small micro-TBM. Also, the cutter must be designed for the specific geological conditions under which it will be operating. The disc cutter rings required to bore extremely hard rock are the most expensive of cutter rings. However, they provide little advantage in weaker rock formations where less expensive rings will do the job as well. It is important to choose the correct cutter for the machine and the geological conditions in order to get the best balance of cost and risk.

Jan 1, 2009

By M A. Llano-Serna, D J. Williams, L Bu

Tailings dam failures continue to occur worldwide at an unacceptable rate, averaging two events annually over the last 100 years. Furthermore, tailings dams continue to grow in size, increasing the potential release of tailings and water should they fail. Tailings are generally deposited in a slurry state, with the potential to flow if containment is lost. Understanding the generation and propagation of excess pore water pressure on failure is critical to predicting the flow and run-out of tailings. This paper presents numerical modelling using the material point method indicating that tailings run-out processes may be self-enhanced by the excess pore water pressure induced during the collapse of the containment. Moreover, it is shown that geomorphological features, such as the slope and its surface roughness downstream of the tailings dam, influence the pore water pressure dynamics upstream of the advancing toe.CITATION:Llano-Serna, M A, Bu, L and Williams, D J, 2018. Excess pore water pressure induced during tailings run-out, in Proceedings Mine Waste and Tailings Stewardship Conference 2018, pp 377–383 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 23, 2018

By Grover R. Castle

INTRODUCTION Banks are being asked to assume more and more of the project risk in connection with project financings and, as a result, they are seeking to become more sophisticated in the way that they approach these kinds of financings. For example, a large number of banks that are active with the mining companies have established mining groups which typically will include mining engineers, economists, and bankers, in order to specialize their service to the mining industry. These banks have geared up these special service staffs so that they will be better equipped in their lending activities to analyze, measure and evaluate the project risk. The hope is that a team of these specialists will be able to predict the potential success or failure of these new "green fields" mining ventures that the banks are asked to finance. This paper shall focus on this subject and try to determine just how successful these specialists can hope to become as they analyze these various projects that are brought to them for financing. The approach that shall be taken in this paper shall be to examine how successful the mining companies themselves have been in the past with this same problem. These mining companies should be the most sophisticated in their approach to the analysis of the potential for success of a new mine because they have had the most experience in evaluating these projects. After all, who knows more about developing a mine than the companies that have developed similar mines in the past? For the purposes of this paper, the writer has selected a sample of 18 mines that have been developed over the years and measured the actual performance of each of these sample ventures against the original projections. All of these cases have involved the development of large projects where the sponsors or partners in the venture are substantial and sophisticated companies and where the feasibility studies have been prepared by engineering firms that are considered very professional in the industry or alternatively where the feasibility has been done by a comparable in-house staff. The sample cases have been chosen from a period which starts in 1965 and continues through 1981 with the cases distributed over that period as indicated below: [ ] Fourteen of the projects are open pit mines, three are underground mines and one mine starts on the surface and gradually goes underground. Seven of the mines are in the the far east with the remaining three in either South or Central America. The following four specific questions were examined for each project:

Jan 1, 1985

At Great Boulder Gold Mines Ltd. a number of attempts were made to remove water soluble copper and iron salts from calcined pyritic concentrates but variou~ problems arose and the methods were abandoned.A method of rapid washing has been developed and although some problems still exist a workable process was established which resulted in substantial savings of reagents used for the cyanidation of calcine.INTRODUCTIONIn the Great Boulder flowsheet crude ore is ground in cyanide solution, then thickened, agitated and subjected to two stage filtering. Secondary filter discharge is pulped with flotation circuit water, gassed with sulphur dioxide to precipitate soluble gold and activate pyrite, and after the addition of copper sulphate and xanthate, the pyrite and tellurides are recovered by flotation. Flotation tailings are thickened and pumped to the disposal area and concentrates are calcined in Edwards roasters. Calcine is washed to remove water soluble copper and iron salts and the washings pumped to the flotation section where the copper is used as an activator. The washed calcine is cyanided and returned to the crude ore circuit via the thickeners. Solution for precipitation in the Merrill Crowe is drawn from the crude ore thickeners.Prior to the introduction of calcine washing the calcine contained 7 lb/ton copper of which approximately 60 per cent was water soluble. It also contained approximately 90 lb/tol'l of water soluble iron salts. Both the copper and...

Jan 1, 1964

By Jerry D. Pfarr, Gary Loving, David G. McDowall

Introduction The Open Cut is one of the best known landmarks in South Dakota's Black Hills. It is the site of the original discovery of the Homestake ledge, or lead, in 1876. Between that time and 1945, the Open Cut produced about 36 Mt (40 million st) of gold ore, the majority of which was drawn from underground operations. In 1981, a drilling program was undertaken to determine how much ore the old timers left behind. The results of this study indicated an ore body made up of broken and caved ground, and solid blocks in pillars and crowns. Reserves are estimated at about 6.8 Mt (7.5 million st) of gold ore, grading at 4 g/t (0.116 oz per st). Total tonnage of material, including waste, is estimated at more than 60 Mt (66 million st). The mining plan is a combination open pit and underground operations. The bottom 78 m (255 ft) will be mined by the underground panel-caving method. The open pit was designed as a contractor operation up to the expected milling stage or the point of sustained ore production. The pit was designed to avoid disturbance of US Highway 85 near Lead, SD. In 1983, the South Dakota Board of Minerals and Environment approved the mining and reclamation plan. Test mining began in the north of the Open Cut using open pit methods. Upon completion of the test pit, Homestake Mining's board of directors approved an expanded mine plan. They committed an additional $27.1 million to prestrip the ore body that is expected to yield 23 t (750,000 oz) of gold. Project life is estimated at 15 years. Exploration program In early 1980, 10 diamond drill holes were drilled from the surface in the vicinity of the east side of the Open Cut. The reserves delineated by that drilling were too small to justify their development. A continuation of this study attempted to quantify the tonnage remaining in the bottom of the Open Cut in the crushed remnants of old stope-and-crown pillars. This was originally thought to be minable by improved drawhole methods using highly mechanized procedures. By early 1981, however, thinking turned toward surface mining methods. The effort to quantify minable ore tonnage ultimately involved drilling 385 reverse-circulation rotary drill holes. These were drilled from August 1981 to March 1983 in several groups of holes. Each group, or drilling program, had a slightly different immediate goal or purpose. The holes were drilled to various depths using truck-mounted Drill Tech 40 drills with doublewalled 76 mm (3 in.) pipe with 134 mm (5.25 in.) three-cone rotary bits. Initially, wet drilling was employed using a cyclone collection system. Since cutting recovery was not as high as it should have been, all subsequent drilling was dry. The cuttings were collected and deposited by way of a cyclone into the sample bag. The pickup for the cuttings using the three-cone bits was about 50 mm (2 in.) above the bottom of the hole. Overall cutting recovery for the drilling program was estimated at 75%. Samples were taken on each 1.5 m (5 ft) of hole drilled. Each sample was split into three 48-g (1-2/3 assay ton) charges. The accepted value was the straight average of the three charges. Because of the known "nugget effect," assays were nearly impossible to duplicate. Therefore, to validate the regular assays, a series of screen fire assays were run. Based on this method of assays and checks, it was concluded that 10% more ounces would be produced than indicated by mine grade. Test pit The northernmost portion of the mine as planned is known as the Terraville test pit. This portion was mined ahead of the rest and separately. It provided answers to questions on slope stability, reserves correlation, mill recoveries, mining costs, and surface mining conditions in various weather and ground conditions since many boulders and much timber were expected. The mining plan for this portion of the pit was successful. More ore was mined than was expected. Mixing of Open Cut ore with underground ore in the mill feed was successful. Impact on the surrounding communities

Jan 10, 1987

By Shamsher Prakash

Equivalent spring constants for analysis of footings is a function of Young's modulus of pile material (Ep), shear modulus of soil (Gs), and geometry of the footing and/or piles. Shear modulus is strongly strain dependent in soils. Therefore, the spring constants are strain dependent, which is acceptable to the Geotechnical engineers. However, structural engineers prefer to use spring constants as a function of displacement, rather than strain. In order to develop these relationships, appropriate relationships between strain and displacement are needed. Also, modulus degradation with strain needs be built into these relationships. Description of spring constants kx , ky, and kz in x, y, and z direction and k?, k?, k?, about x, y, and z directions are relatively simple. However, description of kxø and ky?, the cross-coupling spring constants depend on 2 - components of displacement and rotation. Thus, there are 8 spring constants and likewise there are 8 damping constants. Approximate strain-displacement relationships on which displacement dependant k's and c's are based are presented with appropriate equations, and a numerical example is included illustrating k's and c's as functions of displacement.

Jan 1, 2002

By Ray L. Tobie, Douglas E. Julin

GENERAL DESCRIPTION Block caving is a distinct caving method applied mostly to large, massive, ore bodies because of its in¬herent low cost and high production capabilities. Areas of sufficient size are removed by undercutting so that the mass above will cave naturally. Drawing of the caved ore at the bottom of the ore column causes the caving action to continue upward until all of the ore above the undercut level is broken into sizes suitable for handling. When properly applied, block caving results in a lower mining cost per ton than any other under¬ground method. There are three distinct forms: (1) dividing the horizontal area into rectangular or preferably square or nearly square blocks, drawing evenly over the entire area to maintain an approximately horizontal plane of con¬tact between broken ore and caved capping; (2) divid¬ing the horizontal area into panels across the ore body, retreating by undercutting manageable areas from one end of the panel to the other and maintaining an in¬clined plane of contact between the broken ore and caved capping (thus the name panel caving); and (3) no division of horizontal area of the ore body into defi¬nite blocks or panels (this is termed mass caving). Undercutting may be from wall to wall, retreating from one end of the ore body to the other, maintaining an inclined plane of contact between the broken ore and caved capping. The total active caved area is deter¬mined by the size of block that will not produce undue stress on workings below the undercutting level and by total production requirements. This type of operation is also referred to as panel caving at some properties. SUITABLE ORE BODIES Block caving in its various forms is applicable to deposits of various shapes and ores of various strengths. Its success is governed by rigid requirements and limi¬tations. In unsuitable deposits or where improperly em¬ployed, the loss of ore may exceed that of any other mining method. Good planning, systematic work pro¬cedures, careful supervision, and good judgment con¬tribute to its success. Ore Characteristics Included as a necessary characteristic in an ore body suited to a successful block-caving operation is a proper fracture pattern. Ore hardness should be another gov¬erning factor, and the toughness or softness of the ore should be considered. There must be sufficient horizon¬tal area available for expansion of the undercut, if neces¬sary, to start the caving process. Large, massive, ore bodies usually meet these conditions. Veinlike deposits must be wide and dip rather steeply to conform to the block-caving limitations. The capping must be of such a nature as to cave when the underlying ore is being removed since the weight of the overburden aids in crushing the ore. The most favorable capping character¬istic is that it breaks into coarse pieces and resists attrition as the block is drawn. Soft capping, by breaking into fine pieces and sifting into the ore or channeling through the ore to the drawpoints, can seriously dilute the ore and hinder the extraction of clean ore from the blocks. Types of Ore In block caving, a fairly uniform distribution of val¬ues in the ore is necessary. Grade values may range from low grade to high grade, but most often the sys¬tem is applied to low-grade ores. Wide veins, thick beds, or massive deposits of homogeneous ore overlain by ground which will cave constitute suitable ore bodies. The ore must be such that it can be supported while blocks are being developed and undercut but breaks up readily when caved. Some applications include porphyry copper, disseminated molybdenite, hematite in the Lake Superior iron district, limonite deposits in northern Spain, the diamond mines of South Africa, asbestos and nickel in Canada, and massive magnetite in Pennsylvania. Size and Shape Outline of the ore body should be fairly regular, and the sides of the ore body should dip steeply. It may not be economical to mine small portions of ore extending into the walls of the deposit, and low-grade inclusions in the ore cannot be left unmined. Some dilution of the ore, as mining progresses, by low-grade or barren cap¬ping is inevitable and, unavoidably, some clean ore is left between drawpoints. Determining Cavability The use of rock mechanics is sometimes advantageous in helping to determine the cavability of an ore deposit. The US Bureau of Mines has information and testing ability which can assist in determining the cavability of ore types. Several competent consulting orga¬nizations knowledgeable in rock mechanics can also as¬sist in evaluating the cavability of an ore body. Some recent work has been done to better define the cavability index of various ore bodies; an example of how this was determined is discussed by McMahon and Kendrick (1969). The intensity of the fracture pattern is a critical parameter to be analyzed. The question to be answered is, "How intense must the fracture pattern be?" Table 1 is a tabulation of the size of rock fragments that appear at the drawpoint of various block-cave mines around the world. These are indicative of the fracture-pattern spac¬ing and, as can be seen, they vary substantially from the almost sandy material at the Mather mine to fairly sizable blocks at the Grace mine. Several sets of fractures are essential to promote good caving. Ideally, two vertical sets at nearly right angles to each other and a third set nearly horizontal are required to insure a good caving ore body. Since initial caving of an ore mass is almost always by the action of gravity acting on these planes of weakness, it is doubtful

Jan 1, 1982

By F. T. Cawood

This research report was produced in response to the release of the Mineral and Petroleum Royalty Bill. It involved are view of the mineral investment environments of selected countries in order to form an opinion on the appropriateness and competitiveness of the proposed royalty structure for South Africa. Many emerging countries have recently attracted a significant share of foreign investment, as a direct result of changes in policy. These are, by implication, the least risky and most favourable destinations for the ?investment dollar?. The five most competitive countries, namely Chile, Argentina, Peru, Mexico and Brazil have been identified as succeeding with their policy visions of creating favourable domestic investment environments. The investment rules of these countries served as a basis for the compilation of a Competitive Investment Framework (CIF). This is a template of best practice against which other investment environments, such as that of South Africa, can be measured. The framework indicates that mineral royalties should ideally be no more than three per cent, demonstrating that the regime proposed in the Royalty Bill is out of step with international best practice. Mineral royalties can never be viewed in isolation and in the assessment of the rates proposed in the Royalty Bill, one must consider the total taxation package. Such assessment appears in Section 5 of this report, demonstrating that there is reason for concern if National Treasury were to impose the rates stipulated in the Royalty Bill. The alternatives, conclusion and recommendations of the review appear in Sections 9 and 10. These are the result of independent analysis and assessment of the fundamental principles. It is done through the application of a Competitive Investment Framework that avoids the highly subjective and emotional views that often accompany a process like this.

Jan 1, 2003

A commonplace, sometimes overlooked or taken for granted, article of everyday existence is the tin can. Produced from thin gauge low carbon steel, and commonly coated with a mere 60 millionths of an inch of pure tin, the product is however strong and corrosion resistant. Tin-plate serves a myriad of purposes-the most important being to meet the requirements of the canning industry. Australia's newest steel product, tin-plate, was first produced on a commercial scale during the early part of August, 1957. The commissioning of the Hot Dip Tinning Plant (Fig. 1) of The Broken Hill Proprietary Co. Ltd. at Port Kembla, marked the culmination of a proposal dating back to 1939, the advent of World War II having caused its deferment to the present time.Cornish tin is known to have been exported as early as 300 B.C. to Syria and tinned vessels existed even in the year A.D. 23. The tin coating of plate could be described as an ancient art, believed to have had its origin in Bohemia in the thirteenth century. During the next 450 years, the process spread through Saxony to Pontypool in Monmouthshire, but it was a further half century before any measure of success was obtained.Perhaps the two greatest milestones in tin-plate manufacture were the rolling of iron plates first carried out in 1728, thereby outmoding the hammer formed sheet, and more recently, the development of the electro-deposition process. By the nineteenth century, several important changes had arisen; namely, the substitution of steam for water power, coke for charcoal in the production of the base metal, and...

Jan 1, 1958

By Frank E. Federspiel

In 1992, the Bureau of Mines conducted a mineral resource investigation of the 78,000-acre Garns Mountain study area in southeastern Idaho. The area had been proposed for Wilderness. It is underlain by Paleozoic to Mesozoic-age sedimentary rocks about 11,000 feet thick. Rocks have been faulted, offset, and folded as part of the northerly-trending Idaho-Wyoming Overthrust Belt. Study area mineral occurrences are: phosphate in the Permian Phosphoria Formation; coal in the Cretaceous Frontier end Bear River Formations; and uranium in carbonaceous and phosphate-bearing shales and coal of these formations. Four zones of phosphate are estimated to contain about 528 mitt ion tons of rock with over 10% P205; samples of phosphate-rich rock also contained as much as 3950 ppm chromium, 3280 ppm vanadium, 23 ppm gallium, and 5540 ppm zinc. Fragments of poor-quality coal and carbonaceous shale were found in the area; coal was mined near the northeast border. Uranium content up to 0.016% U3O8 is in carbon-rich shales. Oil and gas were drilled for, with minimal success, in folded sedimentary rocks (frontier and Nugget Sandstone) in the Horseshoe anticline, near the east border. Examined workings included 12 dozer cuts, 4 trenches, 2 prospect pits, and 2 caved adits. No production is known, and no leases or claims are held for the area.

Jan 1, 1992

By Margaret Armstrong

The increasing cost of developing new mines and the need to mine more marginal deposits have led mining companies to use more sophisticated estimation methods such as kriging. But until recently, coal mining companies have shown little or no interest in these techniques. However, the stringent quality controls which are now placed on coal have led to a surge in interest in improved estimation methods. This can be seen from the -, number of research projects by private companies and governmental agencies on the application of geostatistics to coal reserve estimation. At present, most of these programs are centered on the problem of estimating the tonnage and the average grade (i.e. ash percent, sulphur content, calorific value, etc.) of the coal in-situ. See, for example, Sabourin (1975), Woods (1976, 1979), Pauncz and Nixon (1979) and Armstrong (1980). These studies show that the established techniques of linear geostatistics can be applied successfully to coal deposits. The study by Woods is particularly interesting because he had access to the actual production figures as well as the initial sample data from the no 2 seam in the Greenside Colliery owned by the Anglo-American Corporation. He showed that the estimates obtained by kriging were consistently closer to the actual production figures than those given by three other estimation methods. This confirms the superiority of kriging as a method of estimating block values for coal. Another advantage of geostatistics is that it can be used to determine the optimal spacing between drill holes. An example of this is given by Armstrong for the coking coal deposit in the Bowen Basin, Australia.

Jan 1, 1981

By Stephen L. Chryssoulis, Stamen Dimov

"Preg-robbing has been known for many years, where increased losses are suffered because dissolved gold is removed from solution by adsorption onto carbonaceous matter. In direct cyanidation circuits it is reasonable to conclude that the preg-robbed gold is either in a gold cyanide [Au(CN) - 2, AuCN] or metallic [Au°] form. However, in circuits with a pre-oxidation step possibilities increase and multiply even further because of water recycle streams containing apparent and not-so-apparent gold ligands. It is in these circuits where speciation of sorbed gold, i.e. the identification of the gold compound, is useful. The importance arises from the fact that proper identification of the sorbed gold species can be used to determine how, when, where and why gold was removed from solution, thereby identifying the exact cause of the problem and providing clues as to how it may be rectified.INTRODUCTIONTo the best of our knowledge, this is the first attempt ever to speciate gold preg-robbed by carbonaceous matter using a surface sensitive microbeam technique. This approach enables the direct determination of gold species sorbed on carbonaceous particulates, opening an entirely new field in the understanding of the chemistry of gold sorption on carbon. The reasoning behind this effort was to add to the current knowledge of the detrimental effect chloride ions have on gold recovery by the pressure oxidation and in particular why in the ensuing CIL step, gold is not being desorbed from the surface of the carbonaceous particles (Table 1)."

Jan 1, 2004

Resolution Copper Mine is planned to be a 2000 m deep panel cave mine with virgin rock temperatures above 80°C in rock with a high crystalline silica content. The planned run-of-mine production rate is 120 kt/day. The project is in prefeasibility evaluation. This paper discusses features of the ventilation system design, which include multiple ventilation shafts with total flow of about 3000 m³/s and both surface and underground refrigeration systems with more than 140 MW total capacity. This will be a very challenging mine to ventilate, but this work has demonstrated that it will be technically achievable with the application of existing technology.

Jan 1, 2014

By R. Venugopal, Saxena. V. K., U. Sankrityayan

"There is a dire need of a technology which can break the mineral interfaces to enhance liberation with minimum energy consumption in mineral beneficiation processes. Understanding of mineral crystal system and interfaces is essential to develop such methodology. So, this study is carried out to find out the role of mineral crystallography in milling and liberation. Data related of mineral crystallography, Vickers Hardness Number, Bond Work Index and beneficiation processes were collected from literature and detailed statistical analysis has been carried out to establish a correlation among these parameters. It was found that atomic weight and cell parameters shown statistically significant relationship with Vickers Hardness Number of minerals, which controls the milling characteristics of ores. Liberation characteristics of five different kinds of ores were analyzed considering the types of interfaces between valuable and gangue minerals. It was found that phase interfaces play vita role in liberation and coarser liberation can be achieved in the minerals of incoherent crystal systems like manganese and zinc ores. Study of this subject can help us to understand the effect of stress wave propagation in solids and its effect on crystal interfaces. It can also help us to design a technology of future to break the mineral interfaces.INTRODUCTIONCommunition is energy intensive mineral processing operation and consumes up to 2% of the total electric power generated in the world. Energy consumption due to comminution is increasing every year as there is growing needs to exploit lower grade ore bodies and process rejects (Blazy et al., 1994; DOE, 2005). Low grade ores require crushing and grinding to liberate the valuable minerals which are further separated from gangues by various beneficiation methods like gravity separation, floatation, leaching etc. Currently, comminution methods are selected based on the reduction ratio and energy consumption but effect of grinding methods on subsequent processing methods specially on liberation also need to be studied to optimize the entire beneficiation process. Most of the attempts made in past to minimize the energy consumption in milling to enhance the mineral liberation were focused on compression mechanism, chemical additives, thermal treatment, microwave treatment and electric disintegration (Kilpelm and Austin, 1982; Norgate and Weller et al, 1994; Bearman et al., 1997; Kumar et al., 2010; Wang el al., 2011; Singh et al., 2012). But the role of mineral crystallography on milling and liberation behavior of minerals not fully explored yet."

Jan 1, 2013

By M. F. Trice

PYROPHYLLITE is a hydrous aluminum silicate (A12Si4010(OH)2)1 that occurs in both the foliated and the massive forms. The foliated variety resembles talc in that it has a greasy feel, a pearly luster, perfect basal cleavage, and usually is white, although in some deposits apple green, gray, brown, russet, and nearly black specimens may be found. The massive deposits of pyrophyllite yield pearl gray to light tan aggregates, some of which are so friable that they may be crushed with the hands into a fine gritty powder. Frequently specimens of the massive variety are studded with crystals of radial structure. The mineral has a specific gravity of about 2.7 and a hardness of 1 on Mobs' scale.1 The chemical composition of commercial pyrophyllite reflects the mineralogy of the deposits, which, according to Stuckey,2 contain pyro-phyllite and quartz, together with chloritoid, sericite, and a few other minerals of negligible importance. The latter, he says, "are noticed in but small quantities to the extent they might occur as accessory constitu-ents of an igneous rock or as products of regional metamorphism or weathering. . . . Quartz is abundant everywhere except in the very best grades of pyrophyllite." Analyses of specimens from several deposits are listed in Table 1, showing that the silica content varies from 57 to 73 percent and the alumina from 22 to 33 percent. Oxides of iron and the alkaline metals are collectively present, with one exception, to an extent of about 1 to 1.5 per cent.

Jan 1, 1940

By S Smit, P J. J Strobo

ZINIFEX Century Mine is a zinc, lead and silver concentrate producer. Flotation is the primary method of minerals separation in this circuit. The flotation circuit is quite large with 79 flotation cells. The levels in each of these interconnected cells are controlled at more than 45 level control stations. Due to the multivariable interaction throughout the circuit, good level control was found to be difficult to achieve by just using normal Proportional-Integral-Derivative (PID) control. As a first step in advanced process control, the FloatStar Level Stabiliser, a multivariable controller developed specifically for flotation circuits, has been commissioned at Century mine to solve the level control problems. The first section of this paper therefore discusses and shows the results and improvements obtained by using the FloatStar level stabiliser. The results and benefits of the FloatStar Flow Optimiser are discussed in the second section of this paper.

Jan 1, 2004

Heap leaching with thiosulfate is only used in one location for the treatment of a refractory preg-robbing gold ore initially pre-treated with bio-oxidation. The chemistry of thiosulfate leaching is complex. Further work is required to develop a better control of the system. A mild refractory gold ore containing pyrite and chalcopyrite was used in this study to investigate the use of thiosulfate as an alternative heap leaching technology. Preliminary bottle rolls tests indicated that similar gold extraction was obtained with cyanide and thiosulfate. In the column leach test, effects of thiosulfate, copper and ammonia concentrations, and their ratio on both gold extraction and reagent consumption, were assessed. The range of reagent concentration were: 0.1 - 1.0M (NH4)2S2O3, 0.03 - 0.10M CuSO4*5H2O, 1.0 - 6.0M NH4OH. The solid-liquid ratio was in the range of 0.83:1 to 5: 1. Best results showed that 72% of gold was extracted in 50 days with 0.3M (NH4)2S2O3, 0.05M CuSO4*5H2O and 6M NH4OH or a combination of 1.0M (NH4)2S2O3, 0.03M CuSO4*5H2O and 3M NH4OH. The reagent consumption at solid/liquid ratio of 3:1 was 37 - 48.6 kg/t (NH4)2S2O3 and 0 - 0.62 kg/t CuSO4*5H2O. The results of thiosulfate leaching were also compared with cyanide leaching.

Jan 1, 1999

By Forrest T. Moyer

Injury experience in the coal-mining industry was worse in 1965 as both the frequency and severity rates of injuries increased 2 and 6 percent, respectively, over comparable data for 1966 (tables 1-8). The retrogression in the safety record during; 1965 resulted frog the less favorable record at bituminous-coal mines, which more than offset the improved injury experience at anthracite mines. The 259 fatal and 11,138 nonfatal injuries occurred at a rate of 45.77 per million man-hours of exposure and resulted in a severity rate of 8,960 days lost per million man-hours. Although the number of fatalities was the second lowest annual total in the recorded history, it was 17 higher than in 1964. At anthracite operations the total of 6 fatalities in 1965 was the lowest annual figure in statistical history. The annual numbers of nonfatal injuries at ill coal mines and the resulting rates of occurrence have varied little in the past 5 years (table 1). The total days lost from all injuries rose from 2.1.3 million in 19b4 to 2.23 million in 1965 (table 5). Fatal and permanent total injuries, with a standard time charge of 6,000 days each, accounted for 1,650,000 days; Permanent partial injuries, 191,932 clays; and temporary total disabilities, 383,911 days. The severity rate for fatal injuries was 6,241, 8 percent higher than in 1964, and the nonfatal severity rate of 2,718 days lost per million man-hours was 2 percent higher than in the preceding year. The average severity of permanent partial injuries ire; 730, or 48 days fewer per injury than in 1964. The average days lost for each temporary total injury was 36 days in each year.

Jan 1, 1968