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By Samantha Beeson, Collin Poole

"This paper examines the development of two new selective flocculation techniques for recovering KCI values from low grade carnallite potash ore.Direct recovery of KCI was achieved using one of two mid-range molecular weight polymers, one non-ionic and the other strongly cationic, in conjunction with the same selective anionic dispersant. From initial response studies on pure minerals and selective flocculation tests on synthetic mixtures, investigations were extended to recover KCI values from a natural carnallite potash ore. A single-stage process achieved products grading up to 61.3% KCI at recoveries up to 85.4% for the cationic flocculant suite and up to 64.8% KCI with recoveries as high as 90.3% for the non-ionic flocculant suite, with dispersant and flocculant dosages of 150-250gr1 and 10-75gr1, respectively. These results clearly demonstrate not only the feasibility of treating such problematic ores at low reagent dosages, but also the potential elimination of desliming prior to froth flotation.The experimental results presented are considered in terms of the reagent adsorption phenomena present in the system and various aspects of floe formation and cleaning.IntroductionAs high-grade sylvinite potash ores become scarce, lower grade sylvinite and carnallite potash ores are becoming more attractive. However, these ores often contain large amounts (up to 15% by weight) of water-insoluble slimes which must be removed prior to KC! flotation, owing to their high adsorptive capacity for the amine collectors used. The slimes are generally removed by scrubbing followed by hydroclassification and any residual material not eliminated is blinded with reagents such as guar gum or starch to inhibit interference in subsequent KC1 flotation.An alternative approach used commercially in Canada by Agrium Inc. involves water-insoluble slimes flotation prior to KC! flotation. In this process, the slimes are selectively flocculated and then separated by froth flotation. Losses of KC! to the slimes product and associated brine requirements are both reported to be less than those encountered in conventional desliming, but the technique is somewhat limited in its application at present."

Jan 1, 1997

By C. W. Gibbs, E. K. Parks

In 1929 the Standard Oil Company of California commenced the development of a device for obtaining temperatures and pressures in flowing and shut-in wells and of suitable running equipment for such a recorder. Although simple in principle, the recorder has required much experimental work in its development. The various problems have now been solved to such an extent that it is believed that the apparatus affords a sturdy and reliable means for obtaining information regarding depth pressures,—apparatus that may be turned over to a well gang for routine operation. As an example of the rough usage the device will withstand a recent incident may be cited. Owing to failure of the lowering line at the conclusion of a run a recorder was dropped some 6300 ft. through 2 1/2-in. tubing filled with gassy oil. The device was fished out undamaged and with the pressure record intact, although the instrument had been subjected to a pressure greater than 2300 lb. per sq. in. for some 24 hr. after impact at bottom. Description of Recorder The instrument is of the continuous recording loaded spring type (Fig. I). Temperature is not recorded continuously but is obtained by maximum recording thermometers in an auxiliary thermometer well. The entire instrument assembly, from bottom of the recorder to top of the socket, is 6 1/2 ft. long and weighs less than 25 lb. The maximum outside diameter of the recorder shell is 1 1/4in., and all attachments need be no larger. This permits lowering the device inside 2-in. tubing. The operating principle is simple. The well pressure acts upon a piston A resisted by the extension of spring B. The spring deflection is recorded by a stylus C upon paper within the recording drum D. Springs.—To improve accuracy at lower maximum well pressures five interchangeable spring assemblies are used. These have maximum pressure ranges of from 250 to 5000 lb. per sq. in. Piston diameter is the same in all cases, and each spring is designed for a deflection of about 5 in. for maximum pressure loading. A complete spring assembly there-

Jan 1, 1934

By Chu E. Ho

Current methods of determining jet grout column diameters are highly empirical and site specific, and do not take into account the jet hydrodynamics and soil parameters explicitly. A theoretical model which describes the jet excavation mechanism is introduced in this paper, with the aim of providing an improved prediction tool that can be used in jetting design. The model is based on the assumption that the velocity distribution in the cutting jet is equivalent to that of a free jet with boundaries corresponding to the dimensions of the cavity excavated in the ground. The shape of the cavity formed depends on the erosional properties of the soil and jet expansion is limited by the resistance at the jet-soil interface. The model predicts that the limit of jet penetration is reached when the average dynamic pressure at the jet tip becomes equal to the ultimate soil bearing resistance. The reliability of the model, evaluated using results from four case histories of jet grouting field trials, is within ±15%.

Jan 1, 2006

By H. C. Hardison

The Yates pool, Pccos County, Texas, has greater potential capacities than any other field in the United States. This field is notable for the large return it yields on capital invested, for low production costs and for the orderly and conservative manner in which it was developed and continues to be operated. The proven area is 20,000 acres and there are 392 producing wells, or approximately 1 well to 50 acres. The theoretical daily potential is 5,250,000 bbl. as determined by l-hr. open-flow gages. Individual well gages vary from a few barrels to 8500 bbl. per hour, and well depths range from 1000 to 1800 ft. The greater part of the production is from a very porous limestone, and by agreement no productive well has penetrated the lime more than 225 ft. The forces driving the oil from the formation to the surface are hydrostatic head, expanding gas or a combination of the two. The degree of effectiveness of either depends upon the structural position of, and porosity encountered in, the individual wells. Approximately 95 per cent. of the wells in the field flow naturally. The pipe line outlet of the field at the present time is 100,000 bbl. per day, which is less than 2 per cent. of the daily potential. Difficulties and Principles of Probation It became evident during the first year of the development of this field that operators would have to devise a scheme to allocate pipe line runs equitably among their different properties. When the idea of proration was first inaugerated, the matter of spacing wells was given consideration. This was given up through fear of legal difficulties. Storage was also a question that was given considerable thought, and a plan was finally developed which went into effect Oct. 1, 1927. It was decided that the oil runs should be prorated according to the total potential capacity of all the wells which the several operators had in the field. The potential capacity of the wells was measured by a l-hr. open-flow gage. All runs to storage by an operator were deducted from his daily allowance.

Jan 1, 1931

By T. Negeri, M. Boisclair

"A hybrid process consisting of flotation and magnetic separation has been developed to concentrate multi-phase rare earth minerals associated with a carbonatite ore that contains a significant amount of niobium. Fine grinding to a d80 of <65 microns was required to obtain acceptable separation of 92% of the contained total rare earth oxide (TREO) and niobium in half the feed mass. Whereas it was possible to obtain 85% of TREO in 28% of the feed mass through staged cleaning, it was demonstrated that at least 87% and 85% recoveries of TREO and Nb respectively could be achieved in 16% of the feed if the rougher-scavenger concentrates were calcined, quenched, reground to <25 µm and then magnetically upgraded. The paper discusses the overall concept of the flowsheet and the experimental strategies that led to this progress. INTRODUCTIONOut of the approximately 250 minerals known to contain rare earth elements (REEs), only about 60 of them contain REEs in significant amounts. Over 95% of the world’s rare earth oxides (REOs) are believed to occur in the form of three minerals (Pradip & Fuerstenau, 1991) namely bastnaesite (Ce,La,Y)CO3F, monazite (Ce,La,Nd,Th)PO4 and xenotime Y(PO)4. An estimated 51.4% of global REO resources are hosted by carbonatite deposits with the same three minerals as major constituents (Weng et. al., 2015). Any one of the REEs or commonly all of them can be found in these minerals (Ravnaas, 2010). As the demand for critical REEs increases due to the ever increasing applications in advanced technologies, many deposits containing other forms of REE minerals will have to be developed. REE minerals are extensively dispersed and intergrown with several oxides, silicates, carbonates, and phosphate minerals. Consequently it is possible to liberate them only at fine grinding and is thus difficult to separate and produce high grade rare earth mineral concentrates. Most of these minerals contain other important elements such as tantalum (Ta), niobium (Nb), uranium (U), and thorium (Th) in varied quantities. REE minerals in the same ore may exist as carbonates, fluorocarbonates, niobates, silicates, phosphates etc. in association with gangue minerals belonging to different mineralogical groups."

Jan 1, 2016

By George Macready

TRINIDAD, British West Indies, is an island near the north coast of South America, situated between latitudes 10° and 11° N:, and. opposite the numerous outlets of the Orinoco River Delta. It is separated from Venezuela by the Gulf of Paria (salt, water) and straits over 5 mi. (8 km.) wide. The area of the island is approximately 1750 sq. mi. (453,- . 250 hectares) and the population is approximately 400,000. The climate is tropical with an annual rainfall of from 45 to 60 in. (114 to 152 cm.). The oil fields consist of several units, or fields, located in the southern half of the island. Approximately 90 per cent. of the total production has been yielded by fields situated within 7 mi: (11.3 km.) of the famous asphalt lake and on the southwest peninsula. The most important producing fields, or units, are the following, which are shown on the accompanying map: Brighton, or Pitch Lake Field, operated by the Trinidad Lake Petroleum Co., Ltd., is situated beside the famous Pitch Lake; it, even encroaches on the lake. Vessigny Field, operated by the Trinidad Lake Petroleum Co., Ltd., is situated 2 mi. (3.2 km.) south of Pitch Lake. Lot One Field, operated by the Petroleum Development Co., Ltd., the United British Oilfields of Trinidad, Ltd., and Stollmeyer, Ltd., is situated 3 mi. south of Pitch Lake upon Lot One of Morne 1&apos;Enfer Forest Reserve and adjoining properties. Parry Lands Field, operated by the United British Oilfields of Trinidad, Ltd., and the Petroleum Development Co., Ltd., is situated 3;2 mi. south of Pitch Lake on Lot Three of Morne l&apos;Enfer Forest Reserve and adjoining properties. Point Fortin Field, operated by the United British Oilfields of Trinidad, Ltd., is situated at Point Fortin, 6 mi. southwest of Pitch Lake. Fyzabad Field, operated by Trinidad Leaseholds, Ltd., is situated several miles southwest of Fyzabad Village and 6 mi. south-southeast of Pitch Lake. Barracpore Field, operated by Trinidad Leaseholds, Ltd., is situated several miles south of. San Fernando and 15 mi. (24.14 km.) east of Pitch Lake.

Jan 9, 1920

By Frederic Benard

PRIoR to 1936, the Ontario Refining Co. received all incoming blister copper from The International Nickel Company&apos;s smelter in the usual form of 460-lb. cakes, or slabs. These were received in open cars in lots of 250 pieces. At the refinery the cakes were unloaded onto narrow- • gauge cars, for convenient handling and pickup by charging cranes. In March 1936, an investigation was started on the feasibility of transferring converter copper in the molten state from the smelter to the refinery anode furnaces, a distance of 1 1/4 miles. Although common in steel practice, the transportation of blister copper in this form had not previoisly been attempted and a preliminary study of the various factors involved was therefore necessary. For this purpose, a series of experiments was made in which molten copper, removed at different stages during the furnace cycle, was held in a small portable holding furnace. This vessel was essentially a refractory-lined steel cylinder provided with ports for filling and emptying, heated by an oil burner. Dimensions of the steel shell were 3 ft. 10 in. diameter by 7 ft. 3 in. long, and thickness of refractories depended on the design and combination used for the particular experiment. The vessel was filled with approximately 2 1/2 tons of copper at a predetermined temperature, the lining having been preheated to a definite point by means of an oil burner. All openings were well luted and temperature readings were made every 15 min. by means of an immersion-type thermocouple. In this manner the rate of heat loss resulting with different refractory design could be developed. The ratio of exterior surface to weight of metal was taken into consideration in comparing the experimental results with what might be expected for a full-sized unit, and observations were also made on type and degree of slag accretion. It was not possible to get as much information on this latter point as desirable, but all evidence showed that the problem would be control of slag build-up rather than attack on refractories. From the information derived, and in consultation with the engineering staff of the supplying company, an order was placed for the first

Jan 1, 1944

By P. R. Bradley

The property of the Alaska Juneau Gold Mining Co., covering the Juneau gold belt for over a mile in length, consists of 136 claims and 24 mill sites; these are near Juneau, Alaska, in what is known as the Harris mining district. Gold was discovered in the district in 1880, by Richard T. Harris and Joseph Juneau, who had been prompted to explore the section by rumors they had heard among the Indians. Gold was first found on the beach at the mouth of Gold Creek and further search revealed the lode deposits in the Silver Bow basin, about two miles inland. The lode locations made immediately thereafter constitute the principal area of the Alaska Juneau group of claims. Gradually news of the strike spread and, during the late eighties, Silver Bow basin became a substantial producer of gold, both from placer work and from quartz milling. Numerous small mills of various types were erected to crush high-grade float and stringers gouged to shallow depths. Placer mining gradually ceased and the pioneer milling operations gave way to larger and better mills, which in time were the predecessors of the two large mills built by the Alaska Gastineau Co. and the Alaska Juneau Co. on the shores of Gastineau Channel. This evolution from arrastres to small mechanical mills, thence to the two large-capacity mills on Gastineau Channel, although a natural sequence of events, had a most important bearing on the development of the district. Knowledge of the character, extent, and value of the ore deposits gained by these early operations could not have been acquired otherwise. Spencer shows1 that the early mills were: Arrastres in 1881-82-83-84-85-86-88; a 5-stamp mill in 1882; a "newly devised grinding mill" in 1884; two Huntington mills in 1886; a 10-stamp mill in 1889; a revolving Dodge mill in 1890; a 5-stamp mill in 1891; and a 20-stamp mill in 1893. He also estimates the gold production to the end of 1893 as $1,000,000 from lode mining and $1,250,000 from placer mining. No estimate can

Jan 1, 1925

By M. Gensamer

ONLY recently has it been realized that preferred orientations are common in hot-rolled steels. In a recent paper, N. P. Goss1 stated that hot-rolled mild steel exhibits a texture different from that charac-teristic of cold-rolled steel. No complete study of these hot-rolling textures has been reported using the pole-figure method of Wever. The material used in the work reported in this paper was the same material used by Gensamer and Mehl2 in a recent study of preferred orientations in cold-rolled steel, to which reference may be made for the composition and thermal history of the steel. Automobile-body sheet bar, 4 in. thick was hot-rolled at two temperatures, 780° and 910° C. (1436° and 1669° F.). (These temperatures refer to the tempera-ture of the furnace in which the steel was heated.) After each pass the steel was returned to the furnace and held there long enough to return to temperature. Although no accurate method was used to measure the temperature after each pass, it was estimated by color that only a slight drop in temperature occurred. In order to determine the effect of any preferred orientation that might have persisted through the normalizing treatment, one specimen at each temperature was rolled in the same direction as the original working direction and the other at right angles to it. The total reduction in each case was 85 per cent, the specimens having been reduced from 0.250 to 0.0375 in. in five passes. Final cooling was in air. Pinhole reflection photograms were taken of all four specimens, with the rolling direction vertical and the X-ray beam at an angle of 7° with the transverse direction (the transverse direction is defined as that direction in the rolling plane perpendicular to the rolling direction). These photograms, taken with a Me target at 35 kv., are reproduced in Figs. 1, 2, 3 and 4. They show that a more pronounced texture was developed by the low-temperature rolling (780° C.), and that of the two

Jan 1, 1936

By Carlos Campos Junior, Leonardo Soares Chaves, Raul Carretero Hoz, Renato Jacome Costa, Davi Bastos Martins de Oliveira

The Vale’s Serra Sul mine, also known as S11D project, is one of the world’s largest open pit iron ore mines, located in Brazil’s Amazon region. The mine produces around 90 Mt of high-quality iron ore annually. Its unique orebody enabled the S11D project to adopt a truckless mining system with In-Pit Crushing and Conveying (IPCCs) for ore exploitation and waste stripping. The IPCCs system requires crushing the waste, mainly composed by the jaspilite lithology, and is also conveyed to waste deposits several kilometers away from the mine. Due to Amazon’s environmental restrictions, the project was designed with preliminary geological knowledge, leading to the installation of primary crushers not designed for the hardest jaspilite, with uniaxial compressive strength (UCS) up to 500 MPa. This resulted in low waste crushing performance, affecting stripping ratio and threatening future production. A project was developed to improve the fragmentation of the jaspilite, allowing higher waste IPCC production, and reducing the risk to the production due to low performance of waste crushing and conveying. The first challenge was the independence of the geological and the geomechanical models, while only the geological model was considered in blast designs, the geomechanical model describes six waste groups of jaspilite classified according to their strength. A new geological model was then created to incorporate the geomechanical classification, allowing different blasting design according to the rock strength of the jaspilite. The blasting pattern on most compact jaspilite was reduced 27%, from 22m² to 16m² in 9 7/8” diameter holes, with a proposed theoretical reduction of 23% of the P80, from 326mm to 250mm. The second challenge was the implementation of the new design, as this pattern size and the high powder factor (2.9 kg/m3) were paradigms for such a large iron ore mine. The measured results have shown 25% reduction of the P80, the crushing rate was increased by 17% (230t/h), saving around $2.4 million annually in crushing and conveying system maintenance.

Jan 21, 2025

By James Humes

THE active mines in the Park City district at the present time are the Silver King Coalition, Daly-Judge, Daly West, and Silver King Consoli¬dated. Several other companies, such as the Daly, American Flag, Ontario, etc., are doing considerable development work. In the Silver King Coalition mines the ore is a replacement of a dark siliceous limestone bed, varying in thickness from 18 in. to several feet. In places the ore has reached a thickness of 25 ft. This limestone bed lies within 100 ft. of the underlying Weber quartzite. There are other lime beds, similar in appearance to this one, in the Park City formation, which is approximately 700 ft. thick; but not one of them has been found to contain any traces of ore. Directly under and in contact with the ore-bearing bed is another, composed of gray lime well studded with dark concretionary nodules; and this one is an unfailing indication of the proximity of the ore-bearing bed. These ore deposits make out from or have connection with some one of the ore-bearing fissures, which were undoubtedly the channels for the circulation of mineral-bearing solutions. It is therefore proper to run prospecting drifts in the fissures, although they seldom contain much ore. As a general rule, drifting in the fissures is not costly. Mining and mucking cost from $2.50 to $4 per foot, and very little timber is required. There are instances, as in the Daly fissure in the Daly-Judge mine, where the expense of keeping drifts open in the fissure is so great that it has proved economical to make the drifts in the foot-wall rock and run crosscuts into the fissure at short intervals. In the Park City mines it is not easy to determine beforehand the best method for extracting the ore from a newly discovered ore shoot in the bedded deposits. While we may get one dimension from the drift run in the fissure, that will not give us a correct idea as to the other dimensions of the orebody. The fissures here referred to have a N. 35° E.

Jan 2, 1915

By David E. Albus, Bidjan Ghahreman

A single-family home constructed on the side of a slope in Los Angeles, CA, was subjected to lateral forces due to soil slope creep on the pile foundation supporting the home. Impacts to the property included cracking and separations in the street, retaining walls, exterior concrete flatwork, and decorative concrete features in the rear yard area. The home was supported by concrete piles that penetrated the surficial soil mantle and were embedded in underlying bedrock. However, the home exhibited minor distress that suggested even the home was being affected by soil creep. Selected points on the home were monitored with an optical survey method and manometer over a period of two years to assess if the home was undergoing vertical or lateral movement. Monitoring data indicated the home was in fact moving down slope in a manner consistent with slope creep. As a result, the authors were concerned that continued movement due to slope creep could result in over-stressing of the piles that support the home and ultimately lead to failure of the foundation system. An evaluation was then undertaken to evaluate the behavior of the foundation system due to forces from slope creep. Research was performed during the course of evaluating the conditions to consider various approaches to modeling pile behavior of the piles supporting the home. Some research has been performed to field measure creep displacement of soils and rock in slopes. These previous works have found a diverse behavior of the soil displacement with depth (velocity profile) indicating the process is very complex. An example of various velocity profiles measured in field experiments is depicted in Figure 1. Other research focused on theoretical models of velocity profiles and exhibited similarities to some of the field measured cases.

Jan 1, 2018

Frothing in Seawater Flotation Systems

Sep 13, 2010

By Dennis R. Dolinar

Screen material in the form of welded wire mesh and geogrids are used in underground coal mines to prevent the fall of small pieces of rock from the roof between roof bolts. Further, if the screen is installed during the production cycle, roof fall injuries can be reduced significantly. Therefore, the National Institute for Occupational Safety and Health (NIOSH), because of the safety implications, conducted an evaluation of screen materials commonly used in U.S. coal mines to determine their support characteristics and identify the parameters that could affect their performance with respect to controlling the fall of rock from the roof surface. To evaluate the load and stiffness characteristics of the screen, a test frame was designed and installed in the Mine Roof Simulator (MRS) at the NIOSH Pittsburgh Research Laboratory (PRL). With this test set up, the screen is bolted at four corners of the frame and a center load is applied. The test set up allows for up to 20 in of screen deflection while the load and screen defection are continuously recorded. A series of tests were conducted to evaluate the effects of various parameters such as bolt tension, the type of load bearing surface and the size of bearing plates on welded screen performance. The most common screen material used in U. S. coal mines is an 8-gauge wire welded in a 4-by-4-in spacing or aperture. The peak screen load was normally limited by wire breakage. The conditions at the bearing plate influence the nature of the wire breakage. However, the screen stiffness was controlled by slippage at the bearing plates, and by weld and wire failure. The size of the bearing plate had a significant effect on the performance of the welded screen. Therefore, the design capacity of the 8-gauge screen is evaluated based on plate size. For a 6-by-6-in bearing plate the average peak load is 2,900 lb with a stiffness of 250 lb/in. For an 8-by-8-in bearing plate, the average peak load is 4,500 lb with a stiffness of 430 lb/in. These test results are based on a 4-by-4-ft bolt spacing. A geogrid mesh was also tested.

By J. Ellzey Hayden, Lucien Eaton

The Cliffs Shaft mine of the Cleveland Cliffs Iron Co., located in the city of Ishpeming, Mich., is the largest producer of hard hematite ore on the Marquette iron range. The two shafts, A and B, lying 820 ft. (250 m.) apart on the east-west trend of the ore formation, were sunk in the early eighties, and, except for four years, have been steady producers of iron ore. Their location, on a hill overlooking the general offices of the company on the one side and the city of Ishpeming on the other, has made the shaft houses landmarks. In the spring of 1919, as the old wooden shaft houses had become shaky and out of line, and would soon be unsafe, it was necessary to replace them with the least possible delay. Three plans of construction were, therefore, considered: To replace the old shaft houses by somewhat similar structures made of wood; to use steel and radically change the design; or to build reinforced-conerete shaft houses outside of the wooden structures, completely enclosing them. The first plan was disapproved, because it is anticipated that the mine will have a long life and the proximity of other buildings makes the danger from fire rather great. Delay in receiving material, and the necessity for closing the shafts during erection, were other factors in the situation. The second plan was disapproved on account of the high cost of steel, slow delivery, and interference with hoisting during the period of erection. The presence near one of the shafts of a large bed of hard gravel suitable for concrete made the cost of the third plan considerably less than the second, and little, if any, delay in hoisting was expected. Our expectations in this regard were fully realized. Furthermore, the work could be done by unskilled labor, which was plentiful, and the erection could be quickly started. When the proposition was submitted to W. G. Mather, president of the company, he recommended that, on account of the prominence of the buildings due to their location, an effort should be made to combine in the design as far as practicable something of architectural beauty. Following his recommendation, the original plans were submitted to the

Jan 1, 1922

By Bruce Hebblewhite

This paper focuses on the recent achievements and practices developed through the Minerals Tertiary Education Council (MTEC) initiative of the Minerals Council of Australia (MCA), in conjunction with the three partner universities involved in mining engineering education – University of New South Wales (UNSW), West Australian School of Mines, Curtin University (WASM), and University of Queensland (UQ). The collaboration between these partner universities and industry has led to the development of shared course material, cross-institutional teaching arrangements, and development of a number of flexible, distance-based delivery programs. The paper will then move on to future initiatives. Importantly, the exciting challenge is the future initiative planned to cement this collaboration by the formation of a “virtual” national school of mining engineering – Mining Education Australia (MEA). This will provide a common 3rd and 4th year program to all three universities from 2007, involving innovative teaching and learning arrangements, greater student exchange potential and a more significant input by industry into the education process. Such an initiative is not without its challenges, and has a long way to go, before becoming a reality. However, it offers greater attractiveness and marketing potential to prospective students, as well as economic efficiencies, whilst delivering a better quality, and hopefully greater quantity of graduates to industry. It can also form a launching platform for a much greater level of international collaboration, through a proposed International Mining Education Network (IMEN) which is currently being floated amongst a number of international educators.

Jan 1, 2005

By Radianto Elprama

For the realisation of the underground railway line RandstadRail in the city area of Rotterdam a bored tunnel is under construction. Diaphragm walls up to a depth of about 40 m were used to construct the building pits for the start shaft, the Blijdorp Station and the end shaft. The water tightness of the diaphragm walls was determined by leakage tests that were executed after the construction of the walls, but prior to the excavation of the building pits. The tests showed satisfactory results with a hydraulic resistance of the diaphragm walls of at least 500 days. To determine the water tightness of diaphragm walls it is recommended to follow the specific procedure as described in this paper.

Jan 1, 2006

By W. P. Johnstone

The computer assisted generation of geological models to represent the geometry of, and grade distribution within, orebodies is a rapidly growing area of the geological profession. Such models are increasingly used as a necessary starting point in determining the mineability of mineral deposits in terms of practical mine design and economic considerations. The transition from traditional techniques such as field or underground mapping and the manual calculation of grade and tonnage estimates, as well as related activities such as the interpolation of cross-sections, the use of stereographic projections, core logging and hand contouring, to computerized applications, is by no means complete. In fact, many geological professionals deliberately avoid the use of computer simulations based on the perception that the resulting models are simplistic and that their construction is not founded on sound geological principles. The purpose of this paper is to assess current thinking on this topic by considering; (a) an overview of advances in the drafting of geological maps, (b) the theoretical basis for multi-dimensional geological modelling, (c) the case for combining traditional and modern techniques, and (d) the benefits to the mining value chain.

Jan 1, 2003

By Tapiwanashe Chenje, David Hewitt, Peter Radziszewski

"Steel media wear is a major cost component in mineral processing. In the case of existing operating mill, steel media wear prediction can be used to help compare the performance of steel media from different suppliers before completing marked ball tests. It can also be used to investigate how ores from different site of a given mine affect steel media wear. In the case of Greenfield cases, steel media wear prediction can be used to estimate the cost of steel media for a given feasibility study.This paper presents a wear testing and prediction methodology that ties together abrasion wear and corrosion wear testing results through the use of a simplified discrete element charge motion model. The experimentation dimension to this methodology takes samples for steel media, mine site ore and water to determine abrasive and corrosive wear parameters of a given media, ore, water set.The test methodology development is illustrated along with a comparison of wear predictions with observed wear rates in some operating mills. A case study is included that illustrates how this methodology was used to estimate steel media wear performance. The paper closes with a discussion on the use of parts of the experimental methodology to investigate other abrasive wear conditions such as liner wear, paste backfill piping and ore breakage as well as avenues for future test development.INTRODUCTIONThe cost of comminution processes represents a significant part of the overall costs of a mining operation. Grinding media/liner wear alone accounts for 15-25% of the total operating cost. The annual steel grinding media consumption is approximately 2.5 mtpa or 0.3% of total world steel production [Smorgon Steel Group, 2005]. Using an estimated grinding media cost of about US$0.75/kg [Halbe and Smolik, 2002], gives a media cost of over US$1.875 billion per annum. Mining continues to shift towards exploitation of lower grade ores as the higher grades get depleted. These low-grade ores require finer grinding for the liberation of the value minerals, and hence it can be assumed that grinding media consumption will also continue to rise.The Bond abrasion test, dating back to 1963, [Weiss, 1985] is still used to estimate the total mill media wear. This is a purely empirical test based on the ore. From Bond abrasion index tests, it is possible to determine a wear rate using the following function:"

Jan 1, 2009

By J. Nell

An overview is given of the matte melting process used for the beneficiation of South African Platinum Group Metal (PGM) ores. Most producers use six-electrode furnaces to melt the concentrate and Pierce Smith converters to oxidize excess iron sulphide from the furnace matte. Concentrates from two reef-types (the so-called Merensky and UG-2 reefs) dominate the feedstock supply. Concentrate from the Merensky reef is low in chromium and do not pose significant melting problems. Concentrates from the UG-2 reef contain much higher levels of chromium, which results in the formation of chromium-rich spinels during melting. This is attributed to the low solubility of chromium in furnace slag under typical operating conditions (an oxygen partial pressure of 10-8-10-9bar at temperatures of 1550°C to 1600°C). During the converting process care is needed to avoid excessively oxidizing conditions, which result in cobalt, nickel and copper losses to the slag. High nickel losses result in the formation ofFe2NiO4(trevorite), which often forms solid solutions with FeCr2O4and Fe3O4. In future, the main challenge in the processing of South African PGM ores will be to cope with ever increasing levels of chromium in the concentrate. A short-term solution to the problem would be to melt the concentrate under slightly reducing conditions to promote the formation of CrO, which is more soluble in the slag than Cr2O3. A longer-term solution to the problem would be to abandon the current matte-smelting process in favour of an alloy-based process, which would offer total flexibility with respect to the amount of chromium that can be tolerated in the feed. Key words: Merensky, UG-2, furnace, slag, matte, converting, spinel, chromium.

Jan 1, 2004