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Aamot, Olav Crone, Research Engr., Guggenheim Bros. Labs., 3,771 10th Ave., New York, N. Y. '29 Abbott, Clarence E., V.P., Charge of Raw Materials, Tenn. Coal, Iron & R. R. Co., 1242 Brown-Marx Bldg., Birmingham, Ala. '04 Abbott, Louis Benjamin, Chief Engr., Consolidation Coal Co., Watson Bldg., Fairmont, W. Va. '15 Abeel, George H., Jr., Cons. Min. Engr 128/ S. Canon Dr., Beverly Hills, Calif. '18 Abel, Raymond L., Asst. Prof., Dept. of Pet. Refin., Univ. of Pittsburgh Pittsburgh, Pa. '28 Abel, Walter D., Cons. Engr 364 S. Cloverdale Ave., Los Angeles, Calif. '32 Abele, Louis T 1457 Euclid Ave., Zanesville, Ohio. '31 Abell, Leo M Sullivan Concentrator, Chapman Camp, B. C., Canada. '31 Abort, Robert H., Asst. Met., Research Lab., U. S. Steel Corp., Lincoln Highway, Kearny, N. J. '19 Abraham, W. E. V., East India United Service Club, 16 St. James' Sq., London, S. W. 1, England. '28 Abstain, Henry T., Engr. & Mgr., Profile-Tamarack Mines Co Yellow Pine, Idaho. '27 Acker, E. O'C., Met 128 East Washington Late, Germantown, Philadelphia, Pa. '86 Ackermat, D. E., Research Met., International Nickel Co., Inc. Bayonne, N. J. '35 Acres, T. H., Geol. Dept., Valuation & Pet. Engrg., Richfield Oil Co., 555 S. Flower St., Los Angeles, Calif. '29

Jan 1, 1934

A brief review is given of the history of the Colliery to date together with the reasons for the development of an additional seam.Investigations into the quality of the coal ami its suitability for the production of metallurgical coke and determination of the economic mining limits and coal reserves are discussed. The commencement of mining operations and the inter-seamarrangements necessary for the disposal of coal won are described with the stages in the opening out of the seam itself and the reasons for the installation of a 48-in. trunk conveyor belt. Choke and manning of face equipment are considered and also the work of secondary development and pillar extraction.It is concluded that with modern face equipment and conveyor belt lines, concentrated mining followed by immediate extraction of pillars is practicable and is giving excellent results at Burwood Colliery.INTRODUCTIONBurwood Colliery, the total holding of which now comprises some 3,140 acres of land and 1,450 acres of sea leases, has mined the Borehole Seam since 1883. The early work took place from Nos. 1 and 2 shafts at Glenrock Lagoon and later on at Nos. 3 and 4 shafts at Whitebridge to which the centre of operations shifted in 1893, shaft sinking at this latter area having commenced in 1888.As all the early operations were confined to the Borehole Seam, reserves of coal on this horizon had been considerably reduced by the mid 1930's, so a cross measure drift was driven to the Victoria Tunnel Seam which is about 300 ft. above the Borehole Seam, and fully mechanized coal winning commenced from the second seam in 1938. Subsequently this cross measure drift was extended to the Borehole Seam and complete mechanization of the south-west sections of this seam achieved by 1945.Four track mounted mechanical units were put into operation in the latter area and these worked in a southerly and...

Jan 1, 1958

It is shown that factors associated with conditions whIch are thermodynamically favourable for the formation of CuS04 in roasting lead to high copper recoveries on leaching. One of these factors is the ratio of the air to solids used for roasting. This can be varied widely without loss of temperature control in the sulphating roasting of many Australian copper concentrates.Thus, the prediction of the air to solids ratio corresponding to a maximum p(SO3) value, for a given concentrate and temperature, is of importance in fluidized bed roasting, since this could be associated with the formation of a maximum quantity of CuS04. Calculations are presented which predict this air to solids ratio, not only for a given concentrate composition, temperature, and pressure, but also for air which is enriched with sulphur dioxide or oxygen.

Jan 1, 1968

By H. Ipek

Industrial grinding machines used in the mineral processing industries are mostly of the tumbling mill type. These mills exist in various types: ball, rod, pebble, autogenous, and semi-autogenous. Balls and rods are traditionally media shapes used in the tumbling mills. In resent years, as an alternative to balls, grinding charges with various shapes are being used. Cylpebs are slightly tapered cylindrical grinding charges with length equaling diameter, and all the edges being radiused. Cylpebs of equal diameter and length have 50% greater surface area and 45% greater weight than balls of the same material. In addition, they have 9% higher bulk density than steel balls, or 12% higher bulk density than cast balls. An important advantage of cylpebs over balls is that cylpebs produce narrower granulometric distribution due to their shapes. However, some discrepancies were found in the literature regarding the milling performance of cylpebs in comparison with balls (Chi, 2004). In recent studies, Chi (2004) has provided detailed information on cylpebs and has compared these to balls in relation to media mass, size distribution, surface area, and input specific energy. Ipek (2006) has investigated the effects of cylpebs and balls with equal mass and surface area on the breakage rate of quartz, using cylpebs and balls separately. Similarly, Lameck and Moys (2006) have investigated the effects of balls and worn balls on breakage rate of material using worn balls as grinding environment. However, no study has attempted to investigate the effects of grinding environment other than balls on the brekage distribution. As indicated by Prasher (1987), the breakage distribution function is independent of the following variables in ball mills: hold-up (Mika et al., 1967; Kelsall et al. 1969), mill speed and ball load (Herbst and Fuerstenau, 1972), media load and ball density (Kelsall et al., 1973), mill diameter (Malghan and Fuerstenau, 1976), and powder charge (Shoji et al. 1980). The aim of the present paper is to compare the effects of cylpebs and balls on the breakage distribution function of quartz with respect to same mass and same surface area. BACKGROUND Comminution as a rate process is called kinetic approach. Studies on the kinetic approach are reviewed by Oner (2001). Oner states that grinding kinetics are based on studies by Kolmogorov in 1941 and that Robert (1950) is the first to suggest this approach. Later, Melzak, Loeveday, and Reid have improved the grinding kinetics equation; and, Austin and Lynch have conducted two basic studies on the analytical solutions of these equations. Considering the results of studies comparing the size reduction mechanisms, it has been accepted that breakage constitutes of two operations. These two basic operations, first defined by Epstein (1948), are those of selection of material for breakage and the distribution of breakage material to particle sizes smaller than the initial size (Lynch, 1977). Epstein defined the two operations as: ? Pn(y), the probability of breakage of a particle size in the nth step of the breakage process. ? F(x,y), the cumulative distributions by weight of particles of size x<y arising from the breakage of a unit mass of a size y. Probability of breakage is called a selection or a breakage-rate function (Si). Characteristics size distribution after breakage is called a breakage distribution or appearance function (Bi,j). Due to the difficulties in representing particle size distributions mathematically and constantly, the batch equation based on particle size fraction is suggested. Thus, as is the case in breakage distribution and specific rate of breakage, the breakage parameters were defined by batch grinding experiments conducted in laboratories (Oner, 2001). When effective breakage occurs in the mill, the breakage of material with single feed size fraction usually fits first order grinding kinetics and the specific rate of breakage values can be calculated through the slope of semi-log plot. The values of the breakage distributions are deduced from the size distributions at short grinding times where there is mainly size 1 material breaking and only small amount of smaller sizes to rebreak (Austin et. al., 1984). B values can be calculated by using BII. [ ] where j is the top size of the charge, Pi(t) is the cumulative percent undersize of i th size interval at the short grinding time which gives no more than about 30% broken out of the top size interval. Cumulative primary breakage distribution, Bi,j, is also defined in an empirical form by [ ] where xi is the top size and Bi,j is the weight fraction of primary breakage products. The parameters F, ? and ß define the size distribution of the material being ground. On plotting size versus Bi,j on log scale, the slope of the lower portion of the curve gives the value of ?, the slope of the upper portion of the curve gives the value of ß, and F is the intercept as in Fig.1. [ ]

Jan 1, 2009

By Steven R. Lowe

Construction of a pump station, consisting of an inlet chamber, wet well and pump chamber, required excavation to a depth of 9.7 m (32 ft) below ground surface and more than 7.6 m (25 ft) below the ground water table in a rural, but rapidly developing part of southern New Jersey. Private houses are located [ ] within a few hundred feet of the pump station that use shallow wells for water supply. The pump station design anticipated support of the entire pump station on shallow foundations bearing directly on the underlying natural soils. However, during construction, uncontrolled ground water flow caused ground loss and damage to the bearing qualities of foundation soils. As a result, remedial repairs were required including support of the pump station on drilled pile foundations and releveling of the pump chamber that had settled. This case study is an example of where a poor understanding of the significance of upward flowing ground water had a major impact on the cost of construction. Anticipated subsurface conditions The pump station layout is shown in Fig. 1 with the locations of two borings made prior to construction superimposed. One of the borings was located within the footprint of the deep wet well excavation. Generally, this is not advisable where a deep excavation is required below the ground water table. This is because there is a history of projects where ground water seepage, or pressure, has broken out into an excavation through an ungrouted borehole and compromised the integrity of foundation soils. Logs of the borings did not indicate that the borings [ ] were grouted upon completion.

Jan 1, 2011

By George A. Morgan

Output of mineral products was estimated to be unchanged in 1983 from that of 1982 with the exception of petroleum. The country began production of crude oil from the Seme Oilfield, and all output was exported. Wd and charcoal constituted 72% of Benin&apos;s total energy consumption, and imported petroleum products and electric power accounted for the remainder. Cement was produced from imported clinker. The gross domestic product (GDP) was $1,061 million in 1982, the latest year for which information was available.2 COMMODITY REVIEW Gold.-Exploration was underway at Perma by the U.S.S.R. as part of the United Nations development program. Gold had been produced previously at Perma, where a grade of 0.289 troy ounce of gold per ton was reported, An exploitation program for the deposit was to employ about 100 people, and output was targeted to reach 1,600 troy ounces per year. Recoverable reserves were put at 96,000 troy ounces.

Jan 1, 1985

By Robin J. Batterham

We still need to innovate, but it is getting harder. The real price of metals is declining, return on shareholder value is lower than other industries, and there is a shorter-term focus on staying viable in a capital-intensive industry where timeframes for implementation of new technologies can be long (+I5 years). In this environment, the levels of R&D spent by major minerals companies are remaining steady at around 0.2%-0.5% of revenue. This is not large and to remain viable in the future, we need to be smarter about the way in which we do R&D. The mining industry is a knowledge-based industry. Driving down the cost curve demands innovation, which demands knowledge. To be smarter and more effective in our approach to innovation, a new paradigm is required Generating, capturing and using knowledge has evolved from Mode 1 (discipline based and clear distinction between basic and applied research) to Mode 2 (multidisciplinary, team-based approached with flow between basic and applied activities). Mode 2 is now seen to be much more efficient and to deliver better appropriability (Tegart 2003). To survive, we need to innovate, and the best way to do this is collaboratively. Some examples and pointers for the future will be given in this paper. THE CURRENT SITUATION The Rand report and other recent studies have painted a grim outlook for the mining industry. The real price of metals continues to decline (Figure 1) and profit margins remain thin. This is reflected in shareholder returnwhich are well below other industries (Figure 2)-and current management strategies-which reflect something of a "survival mode." The current focus for most mining companies is on short-term, incremental gains

Jan 1, 2004

By Jill Murray

This study focuses on the amosite mining region in South Africa and associated health effects, compared to other mined asbestos fiber types. Historically, dust and fiber levels were high in the amosite mills and mines, and many miners and members of the surrounding communities were exposed to the fibers. Research has shown that amosite produces both benign and malignant disease. Nevertheless, the mesotheliomagenic potential of amosite is several fold lower than crocidolite. The risk of disease associated with amosite exposure is difficult to quantify. Reasons for this include the scarcity of available information, including fiber measurements, and case ascertainment, as well as the juxtaposition of the amosite and crocidolite asbestos seams in South Africa. © 2007 Elsevier Inc. All rights reserved. Keywords: Amosite; Mesothelioma; Crocidolite; South Africa; Asbestos

Jan 1, 2007

Urban development to the south of Sydney is approaching areas of current and proposed coal mining. Already some structures and residential areas are affected by mining subsidence.Subsidence resulting from either conventional pillar extraction methods or longwall mining has been monitored for 15 years to enable the characteristics of the surface deformation to be determined.The vertical and horizontal displacements of survey stations on the surface are monitored and the results arc used to calculate subsidence, changes in slope of the surface, curvatures along the subsidence profile, and the ensile and compressive strains. These features are in turn related to the underground extraction to give a description of surface subsidence and the development of subsidence with time.Subsidence investigations have been carried out over all or part of the workings of six collieries in the Southern Coalfield.Two case histories illustrate the procedure in the investigation of subsidence phenomena. The first example is the study of the subsidence over a series of longwall panels in one seam and the second is a study of subsidence over pillar extraction in two seams one above the other. From the results of the investigations the effects of panel widths on subsidence are determined.This and other information assists mine planning in areas where surface movement and deformation are important considerations.

Jan 1, 1980

By G Dunstan

The Ridgeway gold mine has been an enigma of the international underground hard rock community for nearly 20 years. It nearly didn’t eventuate as it was more than 500 m below a farmer’s paddock in Central New South Wales when the exploration team, looking for open pit deposits, got a sniff on its last attempt and then drilled the high-grade discovery hole. It took separate studies to find an economic mining approach. Newcrest, then a mid-tier mining company, took the venture that set it on a path of growth and built what was to become the backbone of the company for the next 15 years.The initial mining method was sublevel caving (SLC) for the upper 300 m of the orebody. The nameplate production rate for the SLC was 4 Mt/a, which was reached in June 2002 shortly after the commissioning of the materials handling system. The ultimate production rate was subsequently lifted to around 6 Mt/a through a series of innovative and efficiency initiatives. In June 2007, approval was granted for the development of a block cave (BC) operation to extract the ore reserve situated immediately below the existing SLC operations, Ridgeway Deeps. The BC was designed at approximately 6 Mt/a to match the concentrator and a systematic improvement process through the entire underground system again saw delivery rates consistently around 10 Mt/a.In line with the production rate increases, Newcrest was also committed to advancing caving knowledge through a number of targeted research and development experiments and new processes. These were primarily related to improving cave ore recovery through systematic draw marker trials, cave propagation through preconditioning techniques with associated monitoring systems. New technologies were also developed to improve the safety and efficiency of the mine through data collection, cave establishment techniques, production dispatching and autonomous loading.CITATION:Dunstan, G, 2016. The truth is stranger than fiction – the story of Ridgeway gold mine, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 19–30 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By Peter J. Nicholson

Since the reintroduction of soil mixing to the United States over 10 years ago, the work installed by soil mixing technology has grown dramatically. From an annual turnover of less than ten million dollars per year the volume of soil mixing put in place has grown to an estimated $40 million per year in the year 2000. The majority of work done to date has involved walls, either for excavation support, permanent retaining structures or hydraulic cut-off walls. The other major use for soil mixing has been for mass stabilization of soils whether for isolation of contaminants or for retaining structures. This paper will discuss the uses of and costs of typical soil mix structures. BACKGROUND Soil mixing is defined as the in-situ combining of reagents, typically cement, with native soils, the purpose being to modify or improve the ground. This history of soil mixing covers only the use of cement slurry, aggregated with soil in-situ by mechanical mixing methods only. Soil mixing by use of high-pressure hydraulic means, also known as Jet Grouting, will not be discussed. In addition this will not cover ex-situ (above ground) mixing, or mixing dry reagents with soil. Only mixing applications for building, civil and heavy construction applications will be discussed, with only passing reference made to uses of soil mixing for environmental remediation work.

Jan 1, 2001

By D. Wolfe, K. Reid, T. Nunn

"The Aurora North mine has been operating for more than ten years, and is the source of over half of Syncrude’s bitumen production. In order to minimize front-end rejects, and to allow for high throughput, the ore top size at Aurora has been set at a nominal 125 mm. These relatively large lumps have a significant effect on both pump performance and pipeline pressure gradient. Current design methods underestimate the solids de-rating effect of lumpy slurries in pumps, and do not accurately describe the distribution or magnitude of the lump effect on pipeline pressure gradient. Data from the Aurora hydrotransport systems has been compiled, to quantify the effect of lumps on the pump head ratio and pipeline pressure gradient, and how this effect changes as lump ablation occurs along the length of the pipeline. A model has been built, using a logarithmic curve, which predicts the effect of 125 mm lumps on pipeline pressure gradient as a function of velocity and the distance from the start of a hydrotransport pipeline. This model is applicable only to Aurora hydrotransport operation, but similar models could be developed for other operations using the same methods. Using accurate pipeline pressure gradients and pump head ratios when designing future hydrotransport pipeline systems will help improve their reliability and operability."

Jan 1, 2014

By Shimi Tzobery, Jay Davenport, Raymond Henn, Sam Bandimere

Over the past several decades, the use of cement-based grouts using ultra fine cement as a replacement for portland cement has begun to gain acceptance as an alternative to chemical solution grouts used in permeation grouting. The amount of information currently available on grouts made with ultrafine cement is limited. The authors used a previously developed field testing program to perform additional tests to evaluate the permeation performance of portland and ultrafine cement-based grouts. For the first time, the 2003 testing program included an attempt to determine the 28 day compressive strength of the sand permeated by the grout.

Jan 1, 2005

By A. L. Wilder, S. N. Dixon

Introduction Coeur d&apos;Alene Mines Corp.&apos;s largest precious metals property is located in the historic Rochester Mining District 40 km (25 miles) northeast of Lovelock, NV. The property encountered cold weather operational problems soon after its fall start-up in 1986 due to its elevation of over 1830 m (6000 ft). The problem of ice buildup on the heaps because of sprayed solution application was faced immediately. It was felt that allowing ice to build up all winter long until a spring thaw was impractical due to the large area under leach. Further, the operating cost and delivery schedule for a solution heating system was unacceptable. The development and installation of a leach solution distribution system using drip emitters made efficient, cost-effective winter operation possible. Other benefits of this system have also been observed and are discussed here. General process description 15,422 kt/day (17,000 stpd) of - 1.27-cm (-1 /2-in) crushed ore from the three-stage crushing plant are delivered to the leach pad using 77.1 t (85 st) rear dump haul trucks. The ore is drifted into place with a D-9 bulldozer. Leach panels are contiguous and are approximately 8861 m&apos;(90,000 square ft) in area built in 6-m (20-ft) lifts. New panels are built on top of older areas to a final height of 61 m (200 ft). Each panel is ripped and cross-ripped prior to leaching. Barren solution is distributed to the heap using drip emitters at rates of 0.02 to 0.41 L/min/m2 (0.0005 to 0.01 gpm per sq ft), depending on the age of the panels. The pH of the leach solution is 10.7 with a cyanide concentration of 0.75 kg/t (1.5 lb per st). Approximately 50% of the silver and 80% of the gold are finally recovered. Pregnant solution percolates though the heap and flows by gravity into one of two 9.46 ML (2.5 million gal) pregnant solution ponds. The solution is then pumped to a conventional Merrill-Crowe process plant. Clarification takes place in three 9464 L/min (2,500 gpm) capacity filters. The solution is then pumped to a packed vacuum deareation tower for the removal of dissolved oxygen. Typical deareated solution contains 0.7 parts per million dissolved oxygen. Precipitation of gold and silver is accomplished by adding a zinc dust slurry to the deareated solution at the suction of the filter press feed pump. Precipitated gold and silver are recovered in three recessed plate and frame filter presses. Barren solution is discharged into a 11.7 ML (3.1 million gal) pond where cyanide makeup occurs. This solution is pumped back to the heap for further leaching. The precipitate filter cake, containing approximately 75% dore (Ag + Au), is then fluxed with anhydrous borax, soda ash, sodium nitrate and fluorspar to yield a neutral, bisilicate slag. The fluxed precipitate is then charged into a propane-fired melting furnace and heated to 1150° C (2100° F) for 3 1/2 hours. Slag and dore bullion are poured into conical cast iron pots yielding buttons of 800 to 1000 troy oz. The dore typically contains 98.5% silver and 1 % gold. Slag is crushed and tabled to recover the trapped dore blebs and beads. Concentrate from the table is returned to the furnace. Table tails are sent to the crushing circuit and out to the leach pad. Solution application The area kept under leach at Rochester is approximately 130 000 m2 (1.4 million sq ft). Barren solution is delivered to the pad at 21.2 kL/min (5600 gpm) for a resultant application rate of 0.16 L/min/m2 (0.004 gpm per sq ft). A traditional solution sprinkling system using No. 12 Senninger Wobblers with individual pressure regulators was installed at the onset of leaching activities. The Wobblers were placed at 9.1-m (30¬ft) staggered centers and were fed off of a gridwork of Yellowmine plastic piping. Solution flow rates were moni¬tored to each panel. The onset of cold weather with an average nighttime temperature of -12° C (10° F) made it apparent that continual operation would not be possible with the sprinklers. A significant amount of ice was built up on top of the heap, making maintenance and pipe removal dangerous, if not impossible. Leach solution application was restricted to daylight hours to inhibit ice formation. Process plant flow rates were reduced to maintain steady-state operating conditions. However, as daylight temperatures dropped below freezing, ice continued to accumulate due to the sprays. Besides the obvious operating hazards brought on by the growing icefield, there was also the potential environmental hazard associated with an early thaw melting the ice too rapidly for the solution containment facilities. One other option for preventing ice formation was heating of the barren solution prior to spraying. Initial plant design allowed for expansion of the propane storage and distribution system as well as modification of the barren piping for a solution heater. This option was not exercised because the operating costs for an adequate system would have been prohibitive, and timely delivery of a system was not available. An investigation was conducted on the various drip irriga¬tion products available, since subsurface solution applicators would eliminate ice formation altogether. Systems utilizing external flow emitters were ruled out because of their ten¬dency to clog when buried. Emitter systems using perforated tubing were also eliminated from consideration due to their inability to adequately control flow over required lengths of tubing. An in-line emitter system was finally selected which demonstrated clog resistance and adequate flow control, enabling direct burial.

Jan 1, 1990

By R. L. Virta

Sixteen companies mined fuller’s earth in 11 states during 2004. Production during 2004, based on a preliminary survey of fuller’s earth producers, was 3.96 Mt (4.3million st) valued at $340 million. This compared with3.6 Mt (3.9 million st) during 2003, valued at $346 million. The 3.96 Mt (4.3 million st) probably includes some crude clay tonnage. And production of finished product probably was closer to 3.2 Mt (3.5 million st) in 2004. Georgia was the leading producer state of attapulgite-type fuller’s earth, followed by Florida. Georgia also was the leading producer state for montmorillonite-type fuller’s earth. It was followed, in descending order, by Mississippi, Missouri, Florida, Virginia, Illinois, California, Tennessee, Kansas and Texas.

Jan 1, 2005

By J. E. Friant

The Mini-DiscTM cutter drill bit is a radical design departure from the traditional tri-cone rockdrill bit. Miniature disc type cutting tools as small as 3.25 inches (82.5 mm) diameter are fitted to a bit body to cut individual concentric kerfs. The action is the same as that commonly used on large tunnel boring machines but miniaturized. The high performance rate and efficiency of a disc cutter are the result of producing a large proportion rock chips rather than sand or powder. The Mini-DiscTM drill bit, referred to as the “Disc Bit” has other unique features as well. It can be re-greased between runs. It can be rebuilt in the field in about 15 minutes, cutters can be replaced and the Disc-Bit appears to require less down hole hydraulic power than conventional bits. Thoroughly tested in the laboratory, the bit has recently been undergoing field trials in well drilling, blast hole drilling and shaft drilling. As any engineer knows, the laboratory can not shake out all the problems but it eliminates the more glaring weaknesses. Performance results have been exhilarating, and a few mechanical problems were discovered and largely overcome. This paper is a progress report.

Jan 1, 1999

Interest in the tantalum-niobium minerals has been greatly stimulated in Western Australia by the fact that these rare minerals occur at numerous localities in that State, and in the adjoining parts of the Northern Territory, sometimes in mineable concentrations, as at Wodgina and Greenbushes. A large proportion of the world&apos;s tantalum comes from these deposits. The late Dr. E. S. Simpson, in particular, made an extensive study of the chemistry and mineralogy of these minerals, and the following observations are based upon the examination of specimens which he and Mr. H. Bawley, Government Mineralogist of Western Australia, have submitted to me for examination from time to time. The fact that so many of the specimens have been analysed provides a unique opportunity for the interpretation of the phenomena observed in the polished sections.TANTALITE-COLUMBITE SERIESIt has long been considered, on the basis of crystallographic form, chemical analyses, and specific gravity measurements, that ferrotantalite (FeO. Taz05), manganotantalite (MnO. Taz05)&apos; ferrocolumbite (FeO. Nbz0 5), and manganocolumbite (MnO. Nbz0 5 ) form a complete isomorphous solid solution series (Simpson, 1937) and this is confirmed by study of polished sections of these minerals, Pehrman (1929) examined a series of five analysed specimens with Nbz0 5 contents as follows:-

Jan 1, 1940

By Bjarne Ekberg

"The most economical way of dewatering mineral concentrates is to reduce the moisture content in the material mechanically using thickeners and filtration avoiding thermal drying whenever possible. In addition to this, the dewatering process should produce low moisture cake, high quality liquid, consume little energy. As a part of the dewatering process The CERAMEC® capillary filter has been proven to meet these requirements in over 150 installations. The filtration is done by using ceramic filter plates with small pore size. The capillary action in the filter medium secures gas free filtration at good vacuum, and due to the small pore size, filtrate is crystal clean and particle free. The energy consumption is minimal. This paper will discuss results from filters installed at Pyhäsalmi Mine Oy in Finland where the technology was has been used since the late eighties. Results from more recent installations like Los Pelambres in Chile and LKAB in Sweden will also be discussed.INTRODUCTIONThe idea of using capillary action to make dewatering more efficient was invented two decades ago. Ten years later whole filter plants utilizing the new method were delivered to mine concentrators for dewatering of mineral concentrates. Further development of ceramic membrane plates made it possible to meet the requirements in iron ore dewatering. In addition to low cake moisture, high capacity and low energy consumption, there are a lot of environmental advantages in using capillary filtration. One of them is clear filtrate.CAPILLARY FILTRATIONA micro-porous filter material behaves like a large number of narrow capillaries forming a network. When a pressure difference ?P is applied over the wetted, hydrophilic pore with the radius r, the force striving to push the water out of the pore will be ?Ppr2. The force striving to hold the water in the pore will be the vector force 2pr? cosa of the surface tension ? and the cosine contact angle a. Equalling those forces and solving the equation for ?P we get"

Jan 1, 2003

By Goldney L. H

The authors review the literature and discuss the application of methods to the design of thickeners for ilocculated metallurgical pulps. The use and evaluation of flocculents is described. Experimental data show that design based upon laboratory settlingdata is acceptable. The economic aspect of design is also discussed.INTRODUCTIONFor many years there has been little change in the methods for the design of thickener installations. While modifications have been made in the method of applying the data all methods have been based on the observee settling rate of the solids. The development of a theoretical approach based on physical data would involve so many variables which affect the settling rate that the more practical approach of direct observation seems unlikely to be replaced as the basis of design.The advent of synthetic fiocculents has introduced a modification in operation which affects design and should be considered. The authors review the subject and attempt to bring it up to date with present day procedure. Until recently the original work of Coe and Clevenger (1916) remained the basis of thickener design. They described their observation of the different zones encountered in settling, namely clarification, feed, transition, and compression or thickening- zones. The solids pass through the transition zone in settling from feed density to some higher density where settling ceases and &#183;some compression or compaction effect controls. the density obtained. The point at which the settling zone ceases to exist is known as the critical point.

Jan 1, 1959

By R. G. Vardiman, M. R. Achter

Observations of dislocation structure in niobium are presented using two techniques. Triangular pits developed by electroetching are confined to orientations in the stereographic triangle between the (111) pole and planes of the (1ll) zone and require the presence of carbon. Some unusual electroetching phenomena at other orientations are noted. A chemical etch of dislocations decorated by the addition of carbon to 50 ppm reveals the complete dislocation line structure in a surface layer up to 30 p deep, with a resolution of less than 1 p. This effect is independent of surface orientation. The structure of subboundary nets may be resolved for misorientations up to about 1&apos;. Dislocation structures found in strain anneal grown crystals, such as helices, prismatic loops, and high-density areas, are observed. The measurement of dislocation density is discussed and values obtained in different ways are compared. IN the course of some work on the response of refractory metals to thermal-mechanical treatments, it was desired to examine the defect structure of relatively low dislocation density crystals of niobium in bulk form. Surface and bulk methods for the observation of dislocations have been reviewed by ~melinckx,&apos; and include etch pitting, decoration, and various X-ray techniques. The dislocation line structure has been developed by decoration in a few metals, although observations were limited to certain orientations, or by the coarseness and nonselectivity of the precipitate. The most commonly used technique is etch pitting. In niobium, etch pits have been developed chemically&apos;0," in material of high impurity content and electrolyti-cally&apos;&apos; in relatively pure material. In this paper, etch pits developed electrolytically, and the dislocation line structure of a surface layer developed by decoration and chemical etching, are examined, and observations are presented from both techniques. EXPERIMENTAL METHODS Our starting material was electron-beam-melted niobium, swaged to 4 or -in.-diam rod. Chemical analysis is given in Table I. Single crystals were grown by a strain anneal technique using 1 to 2.5 pct of strain and growth temperatures of 1700" to 2300"~.&apos;~ Some crystals grown by zone melting were also examined. The desired surface orientation was exposed either by spark machining or by sectioning with a cut-off wheel and prepared for etching by electro-polishing.

Jan 1, 1969