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By William B. Senseman

FoR reasons well known to mining engineers, wet grinding is quite universal in plants having to do with the extraction of metallic values from crude ores. In the processing of the nonmetallic and industrial minerals, dry grinding is the more common method. The crude materials usually contain less moisture than is required for wet grinding and the products usually are marketed in dry form; most of them as powders. To grind wet, therefore, would entail the addition of water and an expense for drying the product, which is not justified. Furthermore, the material would be a dried cake that would still have to be pulverized or disintegrated before use. For producing fine products, air separators have largely replaced the earlier practice of sizing with screens and bolting reels. Mills in which the air-separating feature is an integral part of the design are now commonplace. Mill designs are many and air systems have developed along various lines, but for the purpose of this discussion, only the fundamental idea need be emphasized— the idea of using air for sizing and removing pulverized material from a grinding mill. About 20 years ago, when the use of mills equipped with air separation had become common practice, heated air or gas was introduced to such a system for the removal of moisture. The purpose of this discussion is to outline the developments and trends that have followed this experiment. It is an important development, for mills equipped with air separation are now universally used for pulverizing not only nonmetallic and industrial minerals, but a great variety of other materials. All such mills are essentially the same. Dried material is pulverized, removed by an air stream, trapped in a cyclone collector, and relatively clean air from the cyclone is returned through a fan to the mill. There is no theoretical need for a vent from the system. The practical need is to vent the air that leaks into the system; and, since the cyclone is not 100 per cent efficient as a dust collector, this vented air carries some dust with it. The initial attempt at drying was made on a system like that described, by the addition of a furnace. The minus pressure within the mill draws hot gas from the furnace. Infiltration air, along with the drying gases and evaporated moisture, is vented as before. The first application was made to a mill used for pulverizing coal. The coal contained very little moisture, so very little drying effect was needed for the success of this experiment. Actually, some calculated efficiencies of over 100 per cent were obtained before due allowance was made for the conversion of mill power into useful heat. When moisture is low, the heat obtained from the conversion of milt power is an appreciable part of that required. That drying can be effected in the short time interval available in a grinding mill of this type is better understood by visual-

Jan 1, 1948

By I. C. R. Gilchrist

A principal requirement in the design of a backfill transportation pipeline is the accurate prediction of the values of key slurry flow parameters. Input variables, such as particle size and solids concentration, are usually determined by the load-bearing requirements of the placed fill. It is therefore necessary to identify the optimum conditions of transport for a particular combination of input variables. This paper critically reviews the methods available to predict the values of flow parameters in slurry pipeline systems in the light of experimental data generated at a test facility. Particular reference is made to rod milled waste and deslimed tailings transport systems and, in the case of the former, it is shown that both pumping and load-bearing requirements can be optimized.

Jan 1, 1988

By P. Henning

Rock fall accidents continue to be the main cause of fatal and serious injuries in the mining industry. Although there has been an improvement in the FOG (falls of ground) related fatal and serious injuries over the past few years, the rate is still too high. New technology has made it possible to reduce this rate significantly. FOG accidents imply failure of the support system or the design of the support system in use. In this paper we deal with the latest technology available to assist the mines to reduce the rates of FOG related injuries and fatalities.

Jan 1, 2010

RESEMBLING in some respects the enterprise at R Chuquicamata is that of the New Cornelia Copper Co. at Ajo, Arizona. Controlled from its inception by the Calumet & Arizona Mining Co., New Cornelia became a unit of the Phelps Dodge Corporation when the latter absorbed the Calumet & Arizona company in 1931. Both Ajo and Chuquicamata are situated in low mountains in extremely arid desert country. Both ore deposits consisted-the past tense is used advisedly because the oxidized ore has nearly all been mined at Ajo-of a horizon of leachable oxidized ores overlying sulphide or concentrating ore. The principal technical problem in each case was the development of a method for treating the oxidized ore. In both instances the capping of waste rock was comparatively thin and power-shovel mining in open cuts was the obvious method of exploitation. Nature had smiled on neither place in the matter of climate. Though Ajo does not have the persistent winds for which Chuquicamata is notorious, the comforts of living are largely artificial. And it is only fair to emphasize that at each place the mining companies have achieved remarkable success in providing these comforts. Of course, in point of size New Cornelia must be content with the inferior role. From May, 1917, when production was started, to the middle of June, 1930, the company leached 16,368,000 tons of oxidized ore, thereby virtually exhausting the ores of this class. Up to the same date, about 15,000,000 tons of sulphide ore had been mined, and a rough estimate gives 170,000,000 tons remaining in reserve. The total tonnage does not loom very large in comparison with Chuquicamata's billion tons; but it is exceeded by similar figures for only four of the Porphyry group.

Jan 1, 1933

Jan 1, 1932

By L. Whitby-Costescu

The Shebandowan mine and concentrator, with a capacity of 9072 t/week (10,000 st per week), was opened in 1972 after several years of environmental study and planning. The mine is located in a popular vacation cottage .area about 96 km (60 miles) west of Thunder Bay, Ontario. Undertaken before environmental assessment procedures had become an established requirement, Shebandowan is one of the first Canadian examples of effectively integrating environmental work into overall planning. Although it is an underground mine, many of the environmental considerations are common to open-pit operations, and hence, it provides a case history for the use of all mine planners. Computers can play a part in such studies by simulating the annual production of mine wastes, by on-line polling of environmental sensors and by processing sensor data. Early attention was focused on establishing base-line conditions in a two-year water quality monitoring program. Since production began, water quality monitoring has shown that mining has not caused any significant changes. However, mine drainage disposal arrangements had to be modified in the light of operating experience. There is no townsite associated with the mine, and workers commute from established municipalities. Harmonious relations have been achieved with local cottage residents after their initial fears were dispelled. Plans are ready to begin revegetating the tailings when a sufficient area of dry elevated tailings has been accumulated. Rehabilitation of the land disturbed during development was begun before start-up and was completed during initial operations.

Jan 1, 1979

By Masayuki Otagawa

THE mining methods adopted in Japanese mines are less known to the mining world than those of other countries, owing to the geographical remoteness, but they present many features of interest to mining engineers. The Ashio copper mine is at the village of Ashio, in Tochigiken, about 109 mi. (175 km.) from Tokyo by the Imperial Government railroad. The area of the concession owned by the Furukawa Mining Co. is 5000 acres (5023 ha.). The mine was formerly divided into three, viz.. Honzan, Kotaki, and Tsudo, which are distant 3 mi. (4.8 km.) from each other as shown in Fig. 1. However, in 1914, when the writer was general manager, the three mining plants were amalgamated into one with the result of simplifying the working projects and increasing the operat-ing efficiency. The mine was visited-by the Institute party in November, 1911, when Mr. K. Inouye was the general manager. Shortly afterward, the writer was appointed manager, and, in 1915, he was succeeded by Dr. K. Asano. At the end of 1917, Mr. K. Kibe became general manager and, after two years, was succeeded by the present manager, Mr. I. Sugimoto-. For some time Dr. J. Kojima was acting general manager. It is needless to say that the gentlemen mentioned are all members of the Institute, as is also Baron T. Furukawa, president of the company. Since the discovery of the enormous masses, called "kajika" or "bullhead" deposits; which the writer has described elsewhere,' the mining methods have been varied to meet the different conditions presented by these deposits.

Jan 2, 1922

By Robert H. Richards

(Read at the Amenia Meeting, October, 1877.) DURING the summer of 1870, I had an opportunity to visit the laboratory of the late David Forbes, Esq., in London, and was much interested in a labor-saving device which lie had attached to his agate mortar. The mortar was placed in the centre of a small table, and about ten inches to the right of the mortar, a post, perhaps two

Jan 1, 1878

By Davey CJ

An integrated computer software system has been developed by Fluor Australia for modelling, mine planning and production scheduling of single or multi-strata deposits. Modelling of both quantitative and qualitative variables (e.g., thickness, basal elevation, density, grade/ash content) is performed via continuous, minimum curvature surface representations. These may be developed from inverse distance weighting or geostatistical kriging interpolation techniques in conjunction with bicubic spline surface interpolation. The modelling process, rather than being purely mathematical, incorporates facilities for reflecting structural geology interpretations (e.g., faulting, pinchouts) and geological boundary conditions (e.g., subcrop and outcrop). Volumetric programs provide the facility for continuous integration over models (at any specified integration increment) for any number of specified irregular polygonal panel areas or regular rectilinearly defined block plan areas. Multi-variable integration is available, incorporating any or all operations of addition, subtraction, multiplication and division to provide multi-strata tonnages, etc., (e.g., Volume 1 * density 1 + Volume 2 density 2). These programs provide true three-dimensional integration with the facility for calculating results on an area by area, level by level, strata by strata as well as overall basis.

Jan 1, 1983

By Scott Horton

In no Confederation of Independent States (CIS) country does the mining sector play quite so commanding a role as in the mountainous Central Asian republic of Kyrgyzstan. Its president, Askar Akayev, says he would like to develop his country into "a Central Asian Switzerland." However, in Switzerland, the gold is largely under the street along Zurich's renowned Bahnhofstrasse. In Kyrgyzstan, though, the gold is still in the mountains. Accordingly, Kyrgyzstan's recent adoption of a comprehensive new mining law has attracted widespread attention. Akayev recently completed a tour of the United States. His visit drew large crowds from the business community at events staged in New York and Washington, DC. The success can be attributed as much to the attractiveness of Kyrgyzstan's min¬ing sector as to the personal charms of the nation's president. Led by a half-billion-dollar investment by Canada's uranium giant Cameco, the mining sector is leading Kyrgyzstan's economic revival. And Bishkek is beginning to show some of the same signs of a "boom town" that already mark its Kazak neighbor, Almaty. Mercedes and BMWs already outnumber the Volgas on Bishkek's major traffic arteries. After nearly two years of debates involving government advisors and a final round of deliberation in Parliament, Kyrgyzstan finally adopted its second mining law. This followed quickly in the wake of measures enacted in neighboring Kazakstan and proposed in Russia. Kyrgyzstan's new law replaces the 1992 law "On Mineral Resources." It was one of the first enactments of an independent Kyrgyzstan. The 1992 law made a major contribution to the introduction of free-market economics in the country by ending the state monopoly over mineral exploration and development. It also laid the groundwork for an independent mining industry. According to Kyrgyz government reports, Kyrgyzstan has about 270 commercially viable deposits of almost 40 types of mineral resources. These include oil, gas, coal, precious-, nonferrous-, ferrous-, rare- and basemetals. However, the nation's key draw is gold. Since enactment of the first law, the government has issued more than 100 licenses for geological prospecting for and development of established deposits of gold.

Jan 1, 1998

One major aim of the operation of Aberfoyle's matte fuming pilot plant was to assess refractories performance in the process working vessel, a SIROSMELT furnace. High wear rates were expected and whether these could be tolerated was an important question. During operations a range of refractory types was evaluated in both insulated and water cooled furnace designs. Furnace operations enabled lining wear measurements to be made. Results confirm that existing technology can solve any expected refractory problems. Design aspects that should be included in commercial furnaces are shown.

Jan 1, 1984

By C C. Souza, S C. Amarante, A C. Arauj

All major companies now producing iron ore in Brazil currently employ or are planning to employ some sort of concentration method to upgrade their fine sized products (sinter feed fines and pellet feed fines). The criteria for the selection of the most suitable concentration method for each application include a series of parameters; the most important among them being related to ore mineralogy. This paper reviews the concentration operations currently employed in Brazil and shows how ore mineralogy plays a fundamental role in the selection of a given method. Examples involving gravity concentration (jigs and spirals), magnetic separation (Rare Earth wet drums, ferrous wheel, and Jones type magnetic separators) and flotation (especially column flotation) are discussed. Fundamentally important for the selection of a concentration method is the mineralogy of iron-bearing minerals and gangue minerals presented to the feed to the concentration unit operation. For instance, the presence of liberated quartz as the major gangue mineral for a pellet feed size stream generally indicates that flotation will be the best selection for the concentration of that stream. However, even if magnetite is present only in the form of crystal relicts within hematite grains, it may lead to the selection of either magnetic separation alone or a combination of magnetic separation and flotation for concentrating such material. Presence of gibbsite and/or kaolinite as the major phase containing alumina will also affect the selection of the concentration method in all cases. Again for the case of pellet feed size streams, kaolinite does not interfere with flotation whereas gibbsite tends to contaminate flotation concentrate (pellet feed fines) as it is depressed together with iron oxides and hydroxides during the reverse cationic flotation process.

Jan 1, 2002

By M. Board, R. Pakalnis, S. Seldon, R. Brummer

The Kidd operation mines roughly 10 000 tpd of copper and zinc ore from underground and employs about 700 personnel. In the late summer and fall of 1997, major movements in the proximity of the east wall of the Kidd open pit had been observed. The Kidd pit had been mined from 1965 to 1977 to depths of 260 m (850 ft), followed by subsequent mining by underground methods to current depths in excess of 1830 m (6800 ft). The No. 1 Mine (2600 Level to pit bottom) is approximately 90% extracted, with most of the remaining reserve between 2000 and 2500 levels. Although initial concerns centred on the stability of the pit walls, the potential ramifications of the movements to the underground were recognized and measures were implemented to ensure the safety of personnel as well as to minimize the potential damage to existing operations. In late October, damage to underground operations was observed as deep as the 2000 Level. In the final assessment, the global mobilized mass was a “wedge”-shaped hangingwall block extending from surface to about 610 m (2000 ft) below surface. The “wedge” boundaries are defined by a fault and shear zone which have continuous lengths in excess of 1070 m (3500 ft); its base by shearing through the host rock mass; and its eastern face by deep tension cracks (Fig. 1). Mined and filled excavations at the western face of the “wedge” allowed mobilization of a mass in excess of 30 M tonnes. Movements in excess of 75 cm (30 in.) per day were recorded on the surface stations that preceded the complete shutdown of operations on October 29, 1997. The evaluations in the days following the shut-down were centred on mitigation of immediate and future possible negative impacts on infrastructure and mining of the remaining No. 1 Mine reserves while ensuring the safety of the operators. This effort required the rapid assessment of the potential damage of the block movement on underground infrastructure, stope access, drainage and ventilation, as well as airblast potential due to the falling of a large block into the pit near the ventilation intake at the base of the pit. These assessments were presented in a previous paper (Seldon et al., 1999). In total, the lateral movement of the block on the ground surface was over 4.3 m (14 ft). This paper describes the detailed studies performed to understand the failure mechanism and to assess the impacts on future mining in No. 1 Mine.

Jan 1, 2000

By Robert E. Kirby

The potash mine and refinery of AMAX Chemical Corporation is located 26 miles east of Carlsbad on the north end of the potash producing areas. This property was developed between 1949-1952 and production of muriate began in August 1952. Two shafts, located 400 feet apart, were sunk to the ore zone 900 feet below the surface. A 15-foot diameter ore hoisting shaft equipped with a 1000-ton ore pocket, crusher and automatic skip loading equipment plus a 20-foot diameter service shaft with a curtain wall for exhausting air were provided. Sylvinite production to date has been from the first ore zone, stratigraphically, the lowest of ten identified mineralized zones in the salt section. This first ore zone has been the top producer of Muriate of Potash over the years. In our area we do have some very marginal mineralization in the third ore zone some 35 feet above the first. We have not mined from this seam as yet but we are investigating the economics very closely with access slopes and an experimental mining panel. Our ore seam is generally flat lying but does have local gentle rolling with grades from 2% to 5%. The mine was designed and developed from the beginning with conventional mining practices and rail haulage from the mining section to the ore pocket. As mining progressed in the early years, production increased from less than one million tons annually to a maximum of 3,935,000 tons in 1964. Efficiency also kept pace with the increasing production and continually improved until the efficiency of the production crews reached an average of 100.3 tons per manshift in 1964.

Jan 1, 1974

By Eleanor Haslem, Carl A. Zapffe

THE present investigation has three principal purposes: I. To develop a method for measuring hydrogen embrittlement that avoids certain errors complicating previously used methods. 2. To explore in a preliminary way some of the fundamental factors controlling embrittlement. 3. To apply the measurements specifically to stainless grades of steel, whose sensitivity or insensitivity to hydrogen embrittlement has never been clearly defined. SELECTING A TEST METHOD Hydrogen embrittlement is a condition of low ductility resulting from excessive absorption of hydrogen by the steel at some period in its history, and may be revealed by almost any test involving cold deformation. Tensile tests, impact tests, and bend tests are commonly used for detecting hydrogen embrittlement; although numerous other tests, such as cold-upsetting or extrusion, could be used. The simplicity of using a wire specimen for cathodic and acid pickling tests, and the especial adaptability of the bend test to wire specimens, led to selection of the bend test for the present research. Bend tests usually have been of three general types: (I) fatigue, (2) reversed bend, and (3) single bend. The first two types, though widely used, are inaccurate because their measurements are counteracted by an aging effect; and the third type has not been developed to its fullest utility. The effect of cold-working in removing hydrogen is well known. During a bend test hydrogen can be observed escaping from the convex side simply by coating the specimen with oil; and any technique employing flexing or reverse bending may therefore have its measurements masked because the mechanical action reduces the embrittlement during the test. This factor is so important that even elastic flexing of a hydrogen-charged wire specimen should be avoided before testing if accurate measurements are desired. DEVELOPING A MACHINE FOR MEASURING HYDROGEN EMBRITTLEMENT As for the single-bend test, its theory is good, and recovery caused by mechanical work is kept to a minimum. With a given degree of hydrogen embrittlement, a specimen will break at a given angle if the bending is applied unidirectionally and uniformly. The radius of the bend must be kept constant; and the rate of bending should be a constant, slow enough to allow reading of the breaking angle, and yet rapid enough to prevent significant recovery from the escape of hydrogen during bending. A simple apparatus (Fig. I) was constructed using a type H Foxboro drive unit A, which is a synchronous motor with a gear-reduction ratio of 600 to I. A cord, fixed at one end to a wooden pulley B on the drive shaft of the motor, follows the pe-

Jan 1, 1946

By Robert K. Waits

The resistivity, temperature coefficient of resistance, stability at 200°C, and structure of annealed Si-Cr films have been studied as a function of film cowposition. Colorimetric analyses of the films indicated 20 to 60 wt pct Cr over the resistivity range of 1 to 10-4 ohm-cm. Resistors were delinealed by photolithographic etching, contacted with aluminum, and mounted in TO-5 packages. The average temperature coefjicient of resistance, a, between 50" and 150°C is related empirically to the film resistivity, p, by a (ppm per °C) = -2900 - 1140 log p (ohm-cm). Near a resistivity of 3.5 x 10-3 ohm-ctn (30 at. pct Cr) , the temperature coefficient changes from positive to negative. The temperature at which the resistance maximum occurs increases with decreasing resistivity. The conduction mechanism between —196" and +360 C appears to have two components—an activated conduction dominating a1 lord chromium content and higher temperatures and a linearly temperature-dependent conduction dominating at high chromium content and low temperatures. Electron diffraction patterns showed the films to consist of Cr3Si and amorphous silicon. Resistors thicker than 150A typically drifted less than 0.6 pct after 5000 hr stor-age at 200°C. Resistors as thin as approxinzately 80°A that had a silicon overlayer showed similar stability. Such films had sheet resistivities of 10 kilohm per square and temperature coefficients near —600 ppm per "c. THIN-FILM resistors having sheet resistivities greater than 1 kilohm per square are required in many integrated circuit applications. A process for producing such resistors should be compatible with present circuit fabrication and packaging, including exposure to maximum temperatures of 300° to 570°C during contact alloying, die-attach, and ceramic package sealing. In addition, the resistors should have a low temperature coefficient of resistance and change less than 1 pct during storage at temperatures up to 200°C for 1000 hr or longer. Reports of work beginning in 1949 at Battelle Memorial Institute indicated that, among other materials, the silicides of chromium, Cr-Co, molybdenum, tantalum, and tungsten were worthy of further study as high-stability resistive films. Later work on Si-Cr films vacuum-deposited by flash evaporation from a tungsten boat1 led to the conclusion that Si-Cr had "excellent properties for development into high- stability ... resistive elements". More recently, Si-Cr films have been vacuum-deposited from a tantalum boat,2 from a BeO-coated tantalum boat,3 by flash evaporation from an electron-beam-heated tungsten Source,4 and by triode sputtering.5 This paper describes a study of Si-Cr films vacuum-deposited by electron-beam heating of Si-Cr alloys. Resistivity, temperature coefficient of resistance, stability at 200°C, and structure were studied as a function of film composition. EXPERIMENTAL PROCEDURE Film Deposition. Si-Cr sources were prepared by vacuum-melting pressed pellets of mixtures of powdered chromium (99.9 pct, -100 mesh, Lunex Co., Pleasant Valley, Iowa) and silicon (99.99 pct, -325 mesh, United Mineral and Chemical Co., New York). A permanent-magnet focused electron-beam gun having a water-cooled copper crucible was used to heat the pellet until complete melting occurred. The mixture was then kept molten for an additional 5 min. The alloy thus formed was used as the source material for Si-Cr film depositions using the same electron-beam gun. Pieces of semiconductor-grade silicon ingots were used as sources for silicon depositions. The films were deposited within a standard vacuum evaporator having an 18-in.-diam Pyrex bell jar, a liquid-nitrogen-cooled trap, and a Consolidated Vacuum Corp. 6-in. diffusion pump containing Dow Corning DC-705 pump fluid. The substrates were thermally oxidized* silicon wafers 2 to 2.5 cm in diam. The

Jan 1, 1969

By Marvin S. Seppanen

Computer programers are often required to make modifications to unfamiliar Fortran routines. This Bureau of Mines report describes a computer program designed to aid the programer in such a situation by reorganizing: Fortran routines. This reorganization includes a sequential renumbering of the routine's statement numbers, a sequential renumbering, and relocation of format statements, an alphanumeric reordering of dimensioned and typed variables, a uniform pattern of text spacing, and a sequential numbering of the records in the final Fortran routine. The computer program has been extensively tested by the author and has proved to be a valuable tool for reorganizing Fortran routines developed under contract and later utilized by the Bureau, and for preparing routines for publication.

Jan 1, 1975

By E. Asselin, P. Sauve, H. Salomon-de-Friedberg

"Residues from the CESL Copper process were leached using the CESL Gold process. The kinetics of reaction were confirmed to be fast with over 50% of the gold consistently extracted within 2 minutes of leaching. Final gold extractions under the conditions tested here were between 80-90% obtained within 90 minutes. A model for the CESL Gold process was developed. This model is slightly different based on whether the copper concentrate leach residue was ground or unground prior to leaching: the gold leaching rate constant is approximately twice as fast for the ground samples. Unground residues had a P80 of 13 ?m whereas ground residues had a P80 of 10 ?m. The model predicts a 43% decrease in leaching time for ground residue versus unground residue to achieve the same extraction under otherwise identical conditions. The model was more accurate for the ground samples. INTRODUCTIONTeck's CESL Process (Jones, 1996) was developed to provide an economically attractive alternative to traditional copper smelting. It has since evolved to also deal with copper concentrates containing valuable impurities such as nickel and undesirable ones, such as arsenic, while reducing downstream costs. Effective recovery of gold and silver is essential to the economics of the CESL Process. The Brook Hunt 2007 copper concentrate assay data (Wood-Mackenzie, 2007) were examined to determine the ratio of gold plus silver value relative to the contained copper. As indicated in Figure 1 below, nearly two-thirds of concentrates have 10% or more of their total value tied up in their precious metals content.CESL has extensively examined the recovery of gold and silver at bench, pilot and demonstration scale, which has culminated in the present CESL Gold Process (Barr et al., 2007). The process targets the precious metal contained in the leach residues produced in the copper portion of the flowsheet. This residue has two primary components which influence downstream processing. The major iron-rich fraction (-75 wt%) of the residue is a mixture of various forms of iron oxide precipitates including hematite and jarosite (Sahu and Asselin, 2011). The iron fraction ranges from amorphous to crystalline in nature depending on the compound precipitated, upstream operating conditions and feed material composition. The lesser, but by no means less important, fraction is a mixture of elemental sulphur and unreacted sulphides (principally pyrite).Elemental sulphur formed during the oxidative leach of the copper concentrate reacts with cyanide to form thiocyanate (SCN-) making a traditional cyanide leach approach too costly. Since the SCN formation reaction strongly depends on contact time with the CESL residues (Po, 2007) short contact times were preferred. This factor guided the development work towards the eventual room temperature, pressure cyanidation process which results in near complete gold dissolution in 90 minutes. A recent CESL gold demonstration campaign treating 1.5 t/d of residue from the copper oxidation step achieved an average gold extraction of 91% (96% maximum)."

Jan 1, 2012

By D. François-Bongareon

"In previous publications by the author, significant advances and generalizations of the theory of sampling of broken ores, prompted and confirmed by experimental results, have been presented. This paper shows how to implement these theoretical results in practice. The models to be used are reviewed, with a description of the various sampling parameters involved. Then the methodology for the experimental assessment of those parameters, including segregation factors, is described in simple and practical terms, along with the use of handy sampling protocol charts. Finally, the particularly important case of processing data from a sample-tree experiment is detailed, and examples of applications covering all situations and types of experiments are provided.IntroductionThe reliability of a sample of crushed material , i.e. the expected (average) deviation between its true grade and the true grade of the lot it comes from, mostly depends upon six factors:• the size (mass) of the sample (and to a minor extent, for small lots, the size of the lot);• the size of the rock particles (degree of comminution)• the nature of the mineralization at microscopic scale (i.e., within even the smallest particles);• the geological nature of the mineralization and mineral of interest;• the grade of the lot in the mineral of interest; and• in some cases, the state of segregation of the lot.In practice, one should first disregard segregation problems, and assess sample variability in the segregation-free case. Once the numerical parameters governing this variability have been assessed, then segregation factors can be assessed in turn. In the case of laboratory standard sample preparation procedures, segregation can usually be disregarded altogether."

Jan 1, 1993

Environmental leaching methods such as the toxicity characteristic leaching procedure (TCLP) test or the ASLP test are used to assess wastes including mining and mineral processing by-products. These tests were initially designed as regulatory tests simulating conditions in a co-disposal municipal landfill waste facility and do not allow leaching performance to be predicted in site specific scenarios. Further, the assumptions underlying these tests tend to define worst-case scenario leaching conditions for by-products from the mineral industry disposed of outside a co-disposal scenario.   Recently some characterisation leaching test methods have been issued as ISO/TS standards (ISO/TS 21268-3:2007 and ISO/TS 21268-4:2007) that aim to define the leaching performance of soil type wastes and to predict the leaching performance of these wastes over long time frames. In this preliminary work these standards are trialled for by-products from the mineral industry and the quality of information that can be obtained from them compared to TCLP/ASLP tests.   Where clear comparisons were possible, the TCLP/ASLP tests were typically found to extract a similar, or often greater, amount of component than (i) tests conducted at the same pH but without the acetate buffer or (ii) for test conditions aimed at simulating long-term leaching in an eluate defined by the waste type. While further work is required, the preliminary results indicated that the tests may provide reliable long-term estimates of leaching performance of wastes in soil types of variable pH and provide a suitable basis for long-term planning for optimal waste handling, treatment or even re-use.

Jun 1, 2010